FEASIBILITY STUDY
AND UPDATED MINERAL RESOURCE ESTIMATE
OF THE
LOS RICOS SOUTH PROJECT,
JALISCO, MEXICO
LATITUDE 21° 02’ 45” N and LONGITUDE 103° 56’ 08” W
UTM WGS84 ZONE 13Q 610,600 m E and 2,327,600 m N
FOR GOGOLD RESOURCES INC.
NI 43-101 & 43-101F1 TECHNICAL REPORT
Andrew Bradfield, P.Eng.
Jarita Barry, P.Geo.
Fred Brown, P.Geo.
David Burga, P.Geo.
D. Grant Feasby, P.Eng.
Eugene Puritch, P.Eng., FEC, CET
D. Gregory Robinson, P.Eng.
William Stone, Ph.D., P.Geo.
David Salari, P.Eng., D.E.N.M. Engineering Ltd.
Tommaso Roberto Raponi, P.Eng., Ausenco Engineering Canada ULC
Scott Cameron Elfen, P.E., Ausenco Engineering Canada ULC
Jonathan Cooper, P.Eng., Ausenco Engineering Canada ULC
James Smith, P.Eng., WSP Canada Inc.
P&E Mining Consultants Inc.
Report 469
Effective Date: January 14, 2025
Report Date: February 28, 2025
TABLE OF CONTENTS
1.0
2.0
3.0
4.0
SUMMARY .........................................................................................................................1
1.1
Introduction ..............................................................................................................1
1.2
Location, Climate, Access and Infrastructure ..........................................................2
1.3
Mineral Tenure.........................................................................................................4
1.4
Surface Rights ..........................................................................................................4
1.5
Environmental ..........................................................................................................5
1.6
Permits .....................................................................................................................5
1.7
Geology and Mineralization ....................................................................................5
1.8
Exploration History..................................................................................................6
1.9
Drilling .....................................................................................................................6
1.10 Sampling and Assaying............................................................................................7
1.11 Data QA/QC and Verification .................................................................................7
1.12 Metallurgical Testing ...............................................................................................8
1.12.1
Process Plant Testing .............................................................................8
1.12.2
SART Testing ......................................................................................10
1.13 Mineral Resource Estimate ....................................................................................10
1.14 Mineral Reserve Estimate ......................................................................................14
1.15 Mining Methods .....................................................................................................17
1.16 Recovery Methods .................................................................................................18
1.17 Project Infrastructure .............................................................................................19
1.17.1
Dry Stack Tailings Facility ..................................................................20
1.17.2
Infrastructure ........................................................................................21
1.17.3
Water Management ..............................................................................21
1.18 Market Studies and Contracts ................................................................................22
1.19 Environmental Studies, Permits, and Social or Community Impacts ....................22
1.20 Capital and Operating Costs ..................................................................................23
1.21 Economic Analysis ................................................................................................24
1.22 Risks and Opportunities .........................................................................................26
1.23 Conclusions ............................................................................................................26
1.24 Recommendations ..................................................................................................27
1.24.1
Metallurgical Testwork ........................................................................27
1.24.2
DSTF Geotechnical..............................................................................27
1.24.3
Mine Geotechnical ...............................................................................27
INTRODUCTION .............................................................................................................28
2.1
Terms of Reference ................................................................................................28
2.2
Site Visits ...............................................................................................................28
2.3
Previous Technical Reports ...................................................................................29
2.4
Sources of Information ..........................................................................................29
2.5
Units and Currency ................................................................................................31
RELIANCE ON OTHER EXPERTS ................................................................................42
PROPERTY DESCRIPTION AND LOCATION .............................................................43
4.1
Option Agreement..................................................................................................47
4.2
The Concession Agreement ...................................................................................47
4.3
Los Ricos South Project .........................................................................................48
4.4
Author Comments on Section 4 .............................................................................49
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
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6.0
7.0
8.0
9.0
10.0
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ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND
PHYSIOGRAPHY .............................................................................................................50
5.1
Accessibility ...........................................................................................................50
5.2
Climate ...................................................................................................................51
5.3
Local Resources and Infrastructure .......................................................................51
5.4
Physiography, Flora and Fauna..............................................................................52
5.4.1
Encino Forest .......................................................................................52
5.4.2
Secondary Vegetation of Low Deciduous Forest ................................52
5.4.3
Induced Pasture ....................................................................................52
5.5
Surface Rights ........................................................................................................53
5.6
Author Comments on Section 5 .............................................................................53
HISTORY ..........................................................................................................................55
6.1
Early History ..........................................................................................................55
6.2
1860 to 1907 - The Martinez and Montero Mine ..................................................55
6.3
1908 to 1930 - The Cinco Minas Mining Company ..............................................56
6.4
James Watson Gerard Papers .................................................................................60
6.5
1931 to 1980 - Several Companies ........................................................................64
6.6
1990 to 2016 - Minera San Jorge, S.A. de C.V. ....................................................64
6.7
2002 to 2005 - TUMI Resources Inc. ....................................................................64
6.8
2005 to 2007 - Bandera Gold Ltd. .........................................................................65
6.9
2008 to 2016 - Litigation between Bandera and MSJ............................................66
6.10 Previous Mineral Resource Estimate .....................................................................67
GEOLOGICAL SETTING AND MINERALIZATION ...................................................70
7.1
Regional Geology ..................................................................................................70
7.2
Local Geology ........................................................................................................72
7.3
Structural Geology .................................................................................................76
7.4
Mineralization ........................................................................................................78
DEPOSIT TYPES ..............................................................................................................81
EXPLORATION................................................................................................................83
9.1
Digital Terrain Model ............................................................................................83
9.2
Geological Mapping and Prospecting ....................................................................85
9.3
Trenching and Channel Sampling..........................................................................90
9.4
Underground Mapping and Sampling ....................................................................97
9.5
Eagle Concession Exploration ...............................................................................99
9.6
Exploration Potential .............................................................................................99
DRILLING .......................................................................................................................101
10.1 Eagle Concession Drilling ...................................................................................101
10.2 Main (Abra) Deposit Drilling ..............................................................................107
SAMPLE PREPARATION, ANALYSES AND SECURITY ........................................113
11.1 Sample Preparation ..............................................................................................113
11.1.1
Channel Sampling ..............................................................................113
11.1.2
Drill Core Sampling ...........................................................................113
11.2 Bulk Density ........................................................................................................115
11.3 2019 Quality Assurance/Quality Control Review ...............................................116
11.3.1
Performance of Certified Reference Materials ..................................116
11.3.2
Performance of Blanks.......................................................................117
11.3.3
Performance of Field Duplicates .......................................................117
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11.4
12.0
13.0
14.0
2020 Quality Assurance/Quality Control Review ...............................................119
11.4.1
Performance of Certified Reference Materials ..................................119
11.4.2
Performance of Blanks.......................................................................120
11.4.3
Performance of Field Duplicates .......................................................121
11.5 2020 Umpire Sampling Program .........................................................................122
11.6 Jul 2020 - May 2023 Quality Assurance/Quality Control Review ......................126
11.6.1
Performance of Certified Reference Materials ..................................126
11.6.2
Performance of Blanks.......................................................................132
11.6.3
Performance of Field Duplicates .......................................................133
11.7 2023 Umpire Sampling Program .........................................................................135
11.8 Conclusions ..........................................................................................................138
DATA VERIFICATION .................................................................................................139
12.1 Database Verification...........................................................................................139
12.2 P&E Site Visit and Independent Sampling ..........................................................140
12.3 Conclusion ...........................................................................................................143
MINERAL PROCESSING AND METALLURGICAL TESTING ...............................144
13.1 Introduction ..........................................................................................................144
13.2 Metallurgical Testwork ........................................................................................144
13.2.1
GoGold Resources Testwork .............................................................145
13.2.1.1 Head Assays ...........................................................................146
13.2.1.2 Mineralogy .............................................................................152
13.2.1.3 Gold and Silver Deportment ..................................................154
13.2.1.4 Comminution Testwork .........................................................156
13.2.1.5 Gravity Concentration Testwork............................................160
13.2.1.6 Leach Testwork ......................................................................161
13.2.1.7 Merrill Crowe Testwork ........................................................177
13.2.1.8 SART Testwork: Three Campaigns .......................................182
13.2.1.9 Liquid-Solid Separation Testwork .........................................195
13.2.1.10 Cyanide Detoxification Testwork ..........................................209
13.3 Metallurgical Variability......................................................................................211
13.4 Deleterious Elements ...........................................................................................217
13.5 Recovery Estimates..............................................................................................217
MINERAL RESOURCES ESTIMATE ..........................................................................225
14.1 Data Supplied .......................................................................................................226
14.2 Economic Considerations ....................................................................................227
14.3 Mineralized Domains ...........................................................................................227
14.4 Exploratory Data Analysis ...................................................................................230
14.5 Compositing .........................................................................................................231
14.6 Treatment of Extreme Values ..............................................................................233
14.7 Continuity Analysis .............................................................................................236
14.8 Block Model.........................................................................................................239
14.9 Grade Estimation, Bulk Density and Mineral Resource Classification ...............240
14.10 Mineral Resource Estimate ..................................................................................240
14.11 Validation ............................................................................................................246
14.12 Cerro Colorado.....................................................................................................249
14.12.1
Data Supplied .....................................................................................249
14.12.2
Economic Considerations ..................................................................250
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15.0
16.0
14.12.3
Mineralization Domains.....................................................................251
14.12.4
Exploratory Data Analysis .................................................................251
14.12.5
Compositing .......................................................................................252
14.12.6
Treatment of Extreme Values ............................................................252
14.12.7
Continuity Analyses ...........................................................................254
14.12.8
Block Model.......................................................................................255
14.12.9
Grade Estimation, Bulk Density and Classification ..........................256
14.12.10 Mineral Resource Estimate ................................................................256
14.12.11 Validation...........................................................................................257
MINERAL RESERVE ESTIMATE ................................................................................260
15.1 Mineral Reserve Summary ..................................................................................260
15.2 Factors that May Affect the Mineral Reserve Estimate .......................................260
15.3 Underground Mineral Reserve Estimate ..............................................................263
15.3.1
Block Model Modifications from Mineral Resource Estimate to
Mineral Reserve Estimate ..................................................................263
15.3.2
Accounting for Historical Mine Workings Material ..........................263
15.3.2.1 Bulk Density ..........................................................................264
15.3.2.2 Grade ......................................................................................264
15.3.2.3 Treatment of HMW Mineral Resources ................................265
15.3.3
Cut-off Grade Calculations ................................................................266
15.3.3.1 Metal Prices, Recoveries and Equivalency Calculations .......266
15.3.3.2 Contractor Estimates ..............................................................267
15.3.3.3 First-Principles Costs Estimate ..............................................267
15.3.3.4 Additional Site Cost Estimates ..............................................267
15.3.3.5 Economic, Marginal and Break-Even Cut-off Grade
Estimates ................................................................................267
15.3.4
Dilution ..............................................................................................269
15.3.4.1 Internal Dilution .....................................................................269
15.3.4.2 External Dilution ....................................................................270
15.3.4.3 Total Dilution and Additional Considerations .......................270
15.3.5
Mining Loss .......................................................................................271
15.3.6
Underground Mineral Reserve Estimate ............................................271
15.3.7
Mining Plan Summary .......................................................................271
15.3.8
Underground Mineral Reserve Estimate ............................................272
15.4 Open Pit Mineral Reserve Estimate .....................................................................273
15.4.1
Open Pit Optimization .......................................................................274
15.4.1.1 Abra Pit Optimization Result .................................................275
15.4.1.2 Cerro Colorado Pit Optimization Result ................................276
15.4.2
Open Pit Designs................................................................................277
15.4.2.1 Open Pit Geotechnical Studies...............................................279
15.4.3
Hydrogeological Studies ....................................................................290
15.4.4
Open Pit Mining Dilution and Losses ................................................291
15.4.5
Open Pit Mineral Reserve Estimate ...................................................292
MINING METHODS ......................................................................................................293
16.1 Underground Mining ...........................................................................................293
16.1.1
Design Methodology..........................................................................299
16.1.1.1 Modified Hill of Value Analysis............................................299
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16.2
16.3
16.1.1.2 Contractor Versus Owner ......................................................300
16.1.1.3 Historical Mine Workings......................................................300
16.1.1.4 Stope Design ..........................................................................302
16.1.1.5 Cut-off Grades .......................................................................303
16.1.1.6 Production Rate......................................................................303
16.1.1.7 Fleet Sizing and Selection......................................................304
16.1.1.8 Backfill...................................................................................305
16.1.1.9 Water Inflows.........................................................................305
16.1.2
Geological and Geotechnical Considerations ....................................306
16.1.2.1 WSP Study and Recommendations .......................................306
16.1.2.2 Stope Sizing and Stability ......................................................321
16.1.2.3 Level Spacing.........................................................................322
16.1.2.4 Faults and Breccias ................................................................322
16.1.2.5 Historical Mine Workings......................................................323
16.1.2.6 Pillars .....................................................................................325
16.1.2.7 High-Grade Domains .............................................................327
16.1.2.8 Ground Support Standards .....................................................327
16.1.3
Cut-off Grades ...................................................................................327
16.1.4
Development ......................................................................................330
16.1.4.1 Lateral Development ..............................................................330
16.1.4.2 Vertical Development ............................................................331
16.1.5
Production ..........................................................................................331
16.1.5.1 Mining Methods .....................................................................331
16.1.5.2 Stope Sizing ...........................................................................336
16.1.5.3 Backfill...................................................................................337
16.1.6
Dilution, Mining Loss and Recovery .................................................339
16.1.7
Equipment ..........................................................................................339
16.1.8
Services and Infrastructure ................................................................340
16.1.8.1 Ventilation..............................................................................340
16.1.8.2 Dewatering .............................................................................347
16.1.8.3 Compressed Air .....................................................................349
16.1.8.4 Electrical ................................................................................349
16.1.8.5 Refuge Stations and Egress ....................................................351
16.1.8.6 Explosives Storage .................................................................353
16.1.8.7 Materials Handling.................................................................353
16.1.8.8 Maintenance Facilities ...........................................................353
16.1.8.9 Backfill Reticulation ..............................................................354
16.1.9
Personnel Requirements.....................................................................354
16.1.10
Underground Mining Schedule ..........................................................355
Open Pit Mining...................................................................................................365
16.2.1
Open Pit Mining Schedule .................................................................365
16.2.2
Open Pit Mining Practices .................................................................369
16.2.3
Contract Mining .................................................................................369
16.2.4
Owner’s Mining Team .......................................................................370
16.2.5
Waste Rock Storage Facilities ...........................................................371
16.2.6
Mine Support Facilities ......................................................................371
Life-of-Mine Production Schedule ......................................................................371
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17.0
18.0
RECOVERY METHODS ................................................................................................373
17.1 Overview ..............................................................................................................373
17.2 Process Flowsheet ................................................................................................373
17.3 Key Process Design Criteria ................................................................................375
17.4 Plant Description..................................................................................................376
17.4.1
Crushing Area ....................................................................................376
17.4.2
Grinding .............................................................................................378
17.4.3
Leaching.............................................................................................380
17.4.3.1 Pre-Leach Thickening ............................................................380
17.4.3.2 Leaching.................................................................................380
17.4.4
Counter-Current Decantation (CCD) Circuit .....................................382
17.4.5
Merril Crowe Circuit..........................................................................383
17.4.6
Refinery..............................................................................................384
17.4.7
SART Plant ........................................................................................385
17.4.7.1 SART Circuit .........................................................................385
17.4.8
Cyanide Detoxification ......................................................................390
17.4.9
Tailings Dewatering ...........................................................................390
17.5 Reagents Handling and Storage ...........................................................................392
17.5.1
Hydrated lime.....................................................................................393
17.5.2
Sodium Cyanide .................................................................................393
17.5.3
Flocculant...........................................................................................393
17.5.4
Coagulant ...........................................................................................393
17.5.5
Antiscalant .........................................................................................393
17.5.6
Fluxes – Silica, Borax, Sodium Carbonate, Nitre ..............................394
17.5.7
Sodium Metabisulphite ......................................................................394
17.5.8
Copper Sulphate (Pentahydrate) ........................................................394
17.5.9
Lead Nitrate .......................................................................................394
17.5.10
Zinc Powder .......................................................................................394
17.5.11
Diatomaceous Earth ...........................................................................395
17.5.12
Sodium Hydroxide .............................................................................395
17.5.13
Sodium Hydrosulphide ......................................................................395
17.5.14
Sulphuric Acid ...................................................................................395
17.6 Energy, Water, and Process Materials Requirements ..........................................395
17.6.1
Electrical Power .................................................................................395
17.6.2
Fresh Water ........................................................................................396
17.6.2.1 Fire Water ..............................................................................397
17.6.2.2 Gland Water ...........................................................................397
17.6.2.3 Safety Showers.......................................................................397
17.6.2.4 Process Water.........................................................................397
17.6.2.5 Barren Solution ......................................................................398
17.6.3
Air Services........................................................................................398
17.6.3.1 Plant and Instrument Air ........................................................398
17.6.3.2 High Pressure Air for Tailings Area ......................................398
17.6.3.3 Precipitate Filter Blow and Pressing Air ...............................398
17.6.3.4 Low Pressure Air for Leach and Detox Circuit .....................399
17.7 Process Materials .................................................................................................399
PROJECT INFRASTRUCTURE ....................................................................................400
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18.1
18.2
Introduction ..........................................................................................................400
Roads and Logistics .............................................................................................402
18.2.1
Site Preparation ..................................................................................402
18.2.2
Access to Site .....................................................................................402
18.2.3
Plant Site Roads .................................................................................404
18.2.4
Airports ..............................................................................................404
18.2.5
Security ..............................................................................................404
18.2.6
Shipping Logistics .............................................................................404
18.2.6.1 Gold-Silver Doré ....................................................................404
18.2.6.2 Copper Sulphide.....................................................................404
18.3 Power and Electrical ............................................................................................404
18.3.1
Electrical System Demand .................................................................404
18.3.2
Facility Power Supply ........................................................................405
18.3.3
Site Power Distribution ......................................................................407
18.3.4
Plant Power Distribution ....................................................................407
18.4 Support Buildings ................................................................................................408
18.5 Ore Stockpile .......................................................................................................409
18.6 Fuel ......................................................................................................................409
18.7 Fire Protection ......................................................................................................409
18.8 Sewage .................................................................................................................409
18.9 Hazardous Waste Storage Facility .......................................................................410
18.10 Mining Infrastructure ...........................................................................................410
18.10.1
Haul Roads .........................................................................................410
18.10.2
Explosives Facilities ..........................................................................410
18.10.3
Paste Plant ..........................................................................................410
18.10.4
Waste Rock Storage Facility ..............................................................411
18.11 Water Supply and Management ...........................................................................411
18.11.1
Water Supply .....................................................................................411
18.11.1.1 Off-Site Water Dams and Overland Water Pipelines ............411
18.11.1.2 Waste Water Pipeline.............................................................412
18.11.1.3 Contact Water Pipeline ..........................................................412
18.11.1.4 Reclaim Water Pipeline .........................................................412
18.11.1.5 Hydrometeorology .................................................................412
18.11.2
Water Management Structures ...........................................................414
18.11.3
Site-Wide Water Balance ...................................................................416
18.12 Dry Stack Tailings Facility ..................................................................................419
18.12.1
Site Description and Overview ..........................................................419
18.12.2
Tailings Characterization ...................................................................421
18.12.3
DSTF Foundation...............................................................................421
18.12.4
Non-Contact Surface Water Diversion Systems ................................422
18.12.5
Underdrain Non-Contact Water Collection System Installation .......422
18.12.6
Composite Liner System ....................................................................422
18.12.7
Overliner Collection System ..............................................................423
18.12.8
Contact Water Collection Ponds ........................................................423
18.12.9
Non-Contact Water Collection Ponds................................................423
18.12.10 DSTF Design Key Elements ..............................................................423
18.12.11 Geotechnical Analysis .......................................................................424
18.12.12 Stability Analysis ...............................................................................424
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20.0
18.12.13 Seepage Analysis ...............................................................................424
18.12.14 Starter Embankment and Stability Berms ..........................................425
18.12.15 Ultimate Embankment .......................................................................426
18.12.16 Filtered Tailings .................................................................................427
18.12.17 Coarse Graded Filtered Tailings Cover .............................................427
18.12.18 DSTF Construction ............................................................................428
18.12.18.1 Starter Phase...........................................................................428
18.12.18.2 Ultimate Phase .......................................................................429
MARKET STUDIES AND CONTRACTS .....................................................................430
19.1 Saleable Products .................................................................................................430
19.2 Contracts ..............................................................................................................430
ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY
IMPACT ..........................................................................................................................431
20.1 Background and Project Location........................................................................431
20.2 Project Local Features..........................................................................................432
20.3 Major Project Components ..................................................................................432
20.3.1
Waste Rock Storage ...........................................................................432
20.3.2
Dry Stack Tailings Facility (DSTF)...................................................433
20.3.3
Process Plant ......................................................................................433
20.4 Project Development............................................................................................433
20.4.1
Site Preparation and Construction .....................................................433
20.4.2
Operations ..........................................................................................434
20.4.3
Closure Planning ................................................................................434
20.5 Background Infrastructure and Environmental Conditions .................................435
20.5.1
Historical Mining Activities ..............................................................435
20.5.2
Historical Environmental Liabilities ..................................................437
20.5.3
Current Activities within the Environmental System ........................437
20.6 Applicable Project Environmental Laws and Regulations ..................................438
20.6.1
Studies, Procedures or Authorizations for the Project under Federal
Jurisdiction .........................................................................................438
20.6.2
Studies, Formalities or Authorizations for the Project Under
Municipal Jurisdiction .......................................................................439
20.6.3
Environmental, Mining Safety and Labour/Social Security
Regulations ........................................................................................440
20.7 Environmental Studies .........................................................................................440
20.7.1
Definition of the Environmental System ...........................................441
20.7.2
Physical Studies .................................................................................441
20.7.3
Water Quality .....................................................................................441
20.7.4
Climatology........................................................................................441
20.7.5
Biological Characterisation of the ES ................................................443
20.7.6
Disturbance Index ..............................................................................443
20.7.7
Cultural and Human Status ................................................................444
20.7.8
Modification of the Environment by Project Perimeter Noise,
Fugitive Dust and Use of Dangerous Substances ..............................444
20.7.9
Summary – Existing Environmental Aspects ....................................444
20.8 Social Aspects ......................................................................................................445
20.8.1
Population ..........................................................................................445
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20.8.2
20.8.3
20.8.4
20.8.5
20.8.6
21.0
22.0
Public Health Services .......................................................................446
Housing and Services.........................................................................446
Indigenous Ceremonial Centers .........................................................446
Local Perceptions about Mining and by Inference, the Project .........446
Environmental Impacts Resulting from the Development and
Operation of the Project .....................................................................447
CAPITAL AND OPERATING COSTS ..........................................................................448
21.1 Introduction ..........................................................................................................448
21.2 Capital Costs ........................................................................................................448
21.2.1
Introduction ........................................................................................448
21.2.2
Initial Capital Costs............................................................................451
21.2.2.1 WBS 1100 Mining Infrastructure ..........................................451
21.2.2.2 WBS 1300 Underground Mine Development........................452
21.2.2.3 WBS 2000 to 4000 Process Plant, Tailings Facility and OnSite Infrastructure Capital Costs ............................................454
21.2.2.4 WBS 2700 SART Plant Cost Estimate ..................................461
21.2.2.5 WBS 5000 Off-Site Infrastructure Capital Costs...................462
21.2.2.6 WBS 6000 Indirect Costs.......................................................463
21.2.2.7 WBS 7000 Project Delivery...................................................465
21.2.2.8 WBS 8000 Owner’s Costs .....................................................465
21.2.2.9 WBS 9000 Contingency ........................................................466
21.2.3
Sustaining Capital Costs ....................................................................467
21.3 Operating Costs....................................................................................................468
21.3.1
Introduction ........................................................................................468
21.3.2
Underground Mining Operating Costs...............................................468
21.3.2.1 Development ..........................................................................469
21.3.2.2 Production ..............................................................................469
21.3.2.3 Power and Diesel ...................................................................470
21.3.2.4 Other Operating Costs............................................................470
21.3.3
Paste Backfill Plant Operating Costs .................................................470
21.3.4
Open Pit Mining.................................................................................472
21.3.4.1 Contract Mining .....................................................................472
21.3.4.2 Supervision and Technical Services ......................................473
21.3.5
Process Plant Operating Costs ...........................................................473
21.3.5.1 Basis of Estimate....................................................................473
21.3.5.2 Process Operating Costs ........................................................474
21.3.6
SART Plant ........................................................................................479
21.3.7
Dry Stack Tailings Facility Operating Costs .....................................479
21.3.8
General and Administrative Costs .....................................................480
21.4 Mining Tax...........................................................................................................480
21.5 Closure Costs .......................................................................................................481
21.6 Cash Costs and All-in Sustaining Costs ..............................................................481
ECONOMIC ANALYSIS ...............................................................................................482
22.1 Summary ..............................................................................................................482
22.2 Basic Assumptions ...............................................................................................483
22.3 Cash Flow Summary ............................................................................................484
22.4 Sensitivities ..........................................................................................................486
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23.0
24.0
ADJACENT PROPERTIES ............................................................................................489
OTHER RELEVANT DATA AND INFORMATION ...................................................490
24.1 Project Execution Plan .........................................................................................490
24.2 Project Organization and Alignment Strategy .....................................................490
24.3 Health, Safety, Environment and Community Management Plan .......................492
24.4 Engineering Management Plan ............................................................................492
24.5 Procurement and Contracts Strategy and Management Plan ...............................493
24.6 Project Controls and Reporting Plan....................................................................493
24.7 Contractor Temporary Facilities ..........................................................................494
24.8 Commissioning Management Plan ......................................................................494
25.0 INTERPRETATION AND CONCLUSIONS .................................................................495
25.1 Risks.....................................................................................................................495
25.1.1
Risk Workshop...................................................................................495
25.1.2
Metallurgical Testing .........................................................................496
25.1.3
Recovery Methods .............................................................................497
25.1.4
Mining Methods .................................................................................498
25.2 Opportunities........................................................................................................499
25.2.1
Mineral Resources .............................................................................499
25.2.2
Mining Methods .................................................................................499
25.2.3
Dry Stack Tailings Facility ................................................................499
25.2.4
Recovery Methods .............................................................................499
25.3 Conclusions ..........................................................................................................500
25.3.1
Location, Mineral Tenure and Geology .............................................500
25.3.2
Drilling ...............................................................................................501
25.3.3
Data Verification................................................................................501
25.3.4
Metallurgical Testing .........................................................................501
25.3.5
Mineral Resources .............................................................................502
25.3.6
Mineral Reserve Estimate ..................................................................502
25.3.7
Mining Methods .................................................................................502
25.3.8
Recovery Methods .............................................................................503
25.3.9
Project Infrastructure .........................................................................504
25.3.10
Environmental ....................................................................................504
25.3.11
Capital and Operating Costs ..............................................................504
25.3.12
Economic Analysis ............................................................................505
25.3.13
Risks and Opportunities .....................................................................505
26.0 RECOMMENDATIONS .................................................................................................506
26.1 Metallurgical Testing ...........................................................................................506
26.2 DSTF Geotechnical Recommendations ...............................................................507
26.3 Underground Geotechnical Recommendations ...................................................507
26.4 Open Pit Geotechnical Recommendations...........................................................508
27.0 REFERENCES ................................................................................................................509
28.0 CERTIFICATES ..............................................................................................................513
APPENDIX A
SURFACE DRILL HOLE PLAN ......................................................526
APPENDIX B
3-D DOMAINS..................................................................................528
APPENDIX C
AGEQ BLOCK MODEL CROSS-SECTIONS AND PLANS .........530
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APPENDIX D
APPENDIX E
APPENDIX F
APPENDIX G
CLASSIFICATION BLOCK MODEL CROSS-SECTIONS AND
PLANS ...............................................................................................541
OPTIMIZED PIT SHELLS ...............................................................552
DRILL HOLE LOCATION DATA...................................................554
DRILL HOLE SIGNIFICANT INTERSECTIONS, VEINS AND
ASSAYS DATA ................................................................................562
LIST OF TABLES
Table 1.1 Los Ricos South Mineral Resource Estimate – Pit-Constrained and Out-of-Pit (1-9) ... 12
Table 1.2 Los Ricos South Mineral Reserve Estimate (1-8) .......................................................... 15
Table 1.3 Metal Prices and Exchange Rate ................................................................................. 22
Table 1.4 Capital Cost Summary ................................................................................................. 23
Table 1.5 Operating Cost Summary............................................................................................. 24
Table 1.6 Economic Evaluation Summary .................................................................................. 25
Table 1.7 Project Cash Flow Summary ....................................................................................... 25
Table 1.8 Silver and Gold Price Sensitivity NPV, IRR and Payback .......................................... 26
Table 1.9 Capital and Operating Cost Sensitivity of NPV and IRR ............................................ 26
Table 2.1 Report Authors and Co-Authors .................................................................................. 30
Table 2.2 Terminology and Abbreviations .................................................................................. 31
Table 2.3 Metric Conversions ...................................................................................................... 40
Table 2.4 Unit Measurement Abbreviations ................................................................................ 40
Table 4.1 Mining Concessions of the Los Ricos South Property (1-2) .......................................... 46
Table 5.1 Annual Temperature Ranges for Cinco Minas ............................................................ 51
Table 6.1 Historical Silver Production from 1901 to 1907 .......................................................... 56
Table 6.2 Production History of the CMMC Company 1914 to 1930......................................... 58
Table 6.3 Reports and Maps of the Cinco Minas Company 1908 to 1930 .................................. 61
Table 6.4 Channel Samples with Silver and Gold Assays in the Cinco Minas Mine 1908 to
1930 .................................................................................................................... 62
Table 6.5 Summary Statistics of the TUMI Drill Program .......................................................... 65
Table 6.6 2023 Mineral Resource Estimate – Pit-Constrained and Out-of-Pit (1-9) ..................... 68
Table 9.1 Assay Results El Troce Samples ................................................................................. 98
Table 10.1 Summary of Drill Programs on the Los Ricos South Property................................ 101
Table 11.1 Summary of Bulk Density Statistics ........................................................................ 116
Table 11.2 Summary of 2019 Au Certified Reference Materials Used at Los Ricos South...... 117
Table 11.3 Summary of 2020 Certified Reference Materials Used at Los Ricos South ........... 120
Table 11.4 Umpire Sampling Program Drill Holes ................................................................... 123
Table 11.5 Summary of Certified Reference Materials Used at Los Ricos South .................... 126
Table 11.6 Umpire Sampling Program Drill Holes ................................................................... 136
Table 13.1 LRS Metallurgical Testwork Program Summary .................................................... 145
Table 13.2 2020-2021 Abra Zone Head Assays ........................................................................ 146
Table 13.3 2021-2022 Abra Zone Composite and Variability Head Assays ............................. 146
Table 13.4 2023 Eagle Zone Hole Composite Head Assays ..................................................... 147
Table 13.5 2024 Abra Zone Samples Head Assays ................................................................... 149
Table 13.6 2024 Eagle Zone Samples Head Assays .................................................................. 150
Table 13.7 2024 CC Zone Samples Head Assays...................................................................... 151
Table 13.8 2021-2022 Abra Zone XRD Analysis Results......................................................... 152
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Table 13.9 2023 Eagle Zone Hole QEMSCAN and XRD Results ............................................ 153
Table 13.10 2024 Eagle, Abra, and CC Zone QEMSCAN and XRD Results .......................... 153
Table 13.11 2020-2021 Los Ricos Vein and Los Ricos Host Comminution Test Results ........ 157
Table 13.12 2021-2022 Abra Zone Comminution Test Results ................................................ 157
Table 13.13 2023 Eagle Zone Comminution Test Results ........................................................ 158
Table 13.14 2024 Eagle, Abra and CC Zone Comminution Test Results ................................. 159
Table 13.15 Eagle Zone High-Grade Sample Gravity Pre-Concentration and Gravity Tailings
Leach Test Results ............................................................................................ 160
Table 13.16 2020-2021 Abra Zone Bottle Roll Leach Test Results .......................................... 162
Table 13.17 2021-2022 Abra Zone Master Composite Leach Bottle Roll Test Results ........... 164
Table 13.18 2021-2022 Abra Zone Variability Sample Bottle Roll Leach Test Results .......... 166
Table 13.19 2023 Eagle Zone Hole Composite Sample Leach Test Results............................. 168
Table 13.20 2024 Eagle Zone Master Composite Bottle Roll Leach Test Results.................... 170
Table 13.21 2024 Eagle Zone Variability Sample Bottle Roll Leach Test Results ................... 170
Table 13.22 2024 Abra Zone Master Composite Bottle Roll Leach Test Results ..................... 173
Table 13.23 2024 Abra Zone Variability Sample Bottle Roll Leach Tests ............................... 173
Table 13.24 2024 CC Master Composite Leach Bottle Roll Test Results................................. 175
Table 13.25 2024 CC Variability Sample Bottle Roll Leach Test Results ................................ 175
Table 13.26 2021-2022 Abra Zone Merrill Crowe Test Results ............................................... 178
Table 13.27 2023 Eagle Zone Hole Composite Merrill Crowe Small Scale Test Results ........ 179
Table 13.28 2023 Eagle Zone Hole Composite Merrill Crowe Bulk Test Results.................... 180
Table 13.29 2024 Eagle, Abra, and CC Zone Merrill Crowe Small Scale and Bulk Testwork
Results .............................................................................................................. 181
Table 13.30 Hole 012 Composite SART Test Summary ........................................................... 185
Table 13.31 Hole 031 Composite SART Test Summary ........................................................... 185
Table 13.32 Hole 014/035 Composite SART Test Summary.................................................... 186
Table 13.33 SART Test Summary ............................................................................................. 187
Table 13.34 Abra Low-Grade Composite SART Test Summary .............................................. 189
Table 13.35 Abra Medium-Grade Composite SART Test Summary ........................................ 189
Table 13.36 Abra High-Grade Composite SART Test Summary ............................................. 191
Table 13.37 Eagle Low-Grade Composite SART Test Summary ............................................. 191
Table 13.38 Eagle Medium-Grade Composite SART Test Summary ....................................... 193
Table 13.39 Eagle High-Grade Composite SART Test Summary ............................................ 193
Table 13.40 Cerro Colorado Composite SART Test Summary ................................................ 194
Table 13.41 Summary of Los Ricos South SART Test Results ................................................ 195
Table 13.42 2023 Eagle Solid/Liquid Feed Sample Characterization ....................................... 196
Table 13.43 Summary of Thickening Results by Thickener Unit Area – BL-2 Pulp ................ 196
Table 13.44 2024 Eagle and Abra Zones Solid-Liquid Testwork Sample List ......................... 197
Table 13.45 2024 Metso Outotec Solid-Liquid Sample SG and PSD Results .......................... 198
Table 13.46 2024 Eagle Zone Pre-leach Dynamic Thickening Results .................................... 199
Table 13.47 2024 Abra Zone Pre-leach Tailings Dynamic Thickening Results ....................... 200
Table 13.48 2024 Eagle Zone Leach Tailings Dynamic Thickening Results............................ 201
Table 13.49 Abra Leach Tailings Dynamic Thickening Results ............................................... 201
Table 13.50 2024 Eagle Zone Detox Tailings Sample Pressure Filtration Test Results ........... 204
Table 13.51 2024 Eagle Zone (New Eagle) Detox Tailings Sample Pressure Filtration Test
Results .............................................................................................................. 204
Table 13.52 2024 Abra Zone Detox Tailings Sample Pressure Filtration Test Results ............ 204
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Table 13.53 2024 Abra Zone (New Abra) Detox Tailings Sample Pressure Filtration Test
Results .............................................................................................................. 204
Table 13.54 2024 Abra Zone (Old Abra) Detox Tailings Sample Pressure Filtration Test
Results .............................................................................................................. 205
Table 13.55 2024 Eagle Zone Detox Tailings Vacuum Filter Results ...................................... 206
Table 13.56 2024 Eagle Zone (New Eagle) Detox Tailings Vacuum Filter Results ................. 207
Table 13.57 2024 Abra Zone Detox Tailings Vacuum Filter Results ....................................... 207
Table 13.58 2024 Abra Zone (New Abra) Detox Tailings Vacuum Filter Results .................... 208
Table 13.59 2024 Abra Zone (Old Abra) Detox Tailings Vacuum Filter Results ..................... 208
Table 13.60 2024 Eagle Zone SO2/Air Cyanide Detoxification Results ................................... 210
Table 13.61 2024 Abra Zone SO2/Air Cyanide Detoxification Results .................................... 210
Table 13.62 Eagle and Abra Bottle Roll and Bulk Leach Tests Used in the Development of
the Recovery Model ......................................................................................... 219
Table 13.63 CC Recovery Estimate Sample Test IDs and Recoveries...................................... 223
Table 13.64 Nominal LOM Gold, Silver and Copper Recovery Estimates for Eagle and Abra
Zones ................................................................................................................ 224
Table 14.1 Drill Hole Summary................................................................................................. 226
Table 14.2 Economic Parameters .............................................................................................. 227
Table 14.3 Mineralized Veins .................................................................................................... 229
Table 14.4 Assay Summary Statistics........................................................................................ 230
Table 14.5 Summary of Bulk Density Statistics ........................................................................ 231
Table 14.6 Composite Summary Statistics ................................................................................ 232
Table 14.7 Silver Composite Capping Thresholds .................................................................... 233
Table 14.8 Gold Composite Capping Thresholds ...................................................................... 233
Table 14.9 Copper Composite Capping Thresholds .................................................................. 234
Table 14.10 Block Model Setup ................................................................................................ 239
Table 14.11 Los Ricos South Mineral Resource Estimate – Pit-Constrained and Out-of-Pit (19)
........................................................................................................................ 242
Table 14.12 AgEq Cut-off Sensitivities: Pit-Constrained Mineral Resource (1-2) ...................... 244
Table 14.13 AgEq Cut-off Sensitivities: Out-Of-Pit Mineral Resource (1-2) ............................. 245
Table 14.14 Comparison of ID3 and NN Average Block Grades .............................................. 246
Table 14.15 Volume Comparison .............................................................................................. 247
Table 14.16 Drill Hole Summary............................................................................................... 250
Table 14.17 Economic Parameters ............................................................................................ 250
Table 14.18 Mineralized Veins .................................................................................................. 251
Table 14.19 Assay Summary Statistics...................................................................................... 251
Table 14.20 Summary Composite Statistics .............................................................................. 252
Table 14.21 Silver Composite Capping Thresholds .................................................................. 253
Table 14.22 Gold Composite Capping Thresholds .................................................................... 253
Table 14.23 Block Model Setup ................................................................................................ 256
Table 14.24 Cerro Colorado Mineral Resource Estimate (1-8).................................................... 257
Table 14.25 Grade Comparison ................................................................................................. 257
Table 14.26 Volume Comparison .............................................................................................. 258
Table 15.1 Los Ricos South Mineral Reserve Estimate (1-8) ...................................................... 261
Table 15.2 Comparison of Original Mineral Resource Model and Revised Model .................. 263
Table 15.3 Old Workings Intercept Data and Bulk Density Calculations ................................. 264
Table 15.4 Historical Workings Intercept Data and Metal Grade Calculations ........................ 265
Table 15.5 Mineral Reserve Metal Prices and Process Recoveries ........................................... 266
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Table 15.6 Initial Cut-Off Grade Calculations .......................................................................... 267
Table 15.7 Underground Mine Plan Summary by Mineral Resource Class1 ............................ 272
Table 15.8 Underground Mineral Reserve Estimate (1-6) ........................................................... 272
Table 15.9 Underground Mineral Reserve by Grade Bin .......................................................... 273
Table 15.10 Pit Optimization Parameters .................................................................................. 274
Table 15.11 Summary of the Available Geotechnical Logging for Each Domain .................... 282
Table 15.12 Classification Results for Each Domain ................................................................ 282
Table 15.13 Hoek-Brown Strength Parameters for Intact Rock by Domain .............................. 283
Table 15.14 Summary of Selected Structural Sets..................................................................... 284
Table 15.15 Summary of the Kinematic Analysis and Simplified Design for Sector 1 – FW
Domain ............................................................................................................. 285
Table 15.16 Summary of the Slope Stability Input Parameters ................................................. 287
Table 15.17 Summary of the Recommended Design Configurations for the Proposed Pits ..... 289
Table 15.18 Abra Pit Net Dilution ............................................................................................. 291
Table 15.19 Cerro Colorado Net Dilution ................................................................................. 291
Table 15.20 Open Pit Mineral Reserve Estimate (1-6) ................................................................ 292
Table 16.1 LOM Totals by Mining Area ................................................................................... 299
Table 16.2 Modified HOV Analysis Parameters ....................................................................... 299
Table 16.3 LOM Totals by Mining Area ................................................................................... 302
Table 16.4 Comparable Equipment Types................................................................................. 305
Table 16.5 Summary of the Logged Lithologies Combined into their Respective Domains .... 309
Table 16.6 Summary of Selected Structural Sets....................................................................... 310
Table 16.7 Summary of Hoek-Brown Intact Rock Strength Envelopes for the Lithological
Units at Proposed Project Underground Mine.................................................. 311
Table 16.8 Summary of RMR76 Mean and Calculated Q' Parameters by Geotechnical Domain
.......................................................................................................................... 311
Table 16.9 Ground Support Selection ........................................................................................ 315
Table 16.10 Backfill Strength for Drifting Driving and Overhand Applications ...................... 319
Table 16.11 Backfill Strength for Underhand Applications (MPa) ........................................... 319
Table 16.12 Modified Stability Number Factors ....................................................................... 321
Table 16.13 WSP Recommended Stope Sizing ......................................................................... 321
Table 16.14 Portion of UG Mineral Reserves Extracted from Crown Pillar Areas .................. 325
Table 16.15 UG Stope Proportions by COG ............................................................................. 328
Table 16.16 Lateral Development Metres Summary ................................................................. 330
Table 16.17 Vertical Development Metres Summary ............................................................... 331
Table 16.18 Average Stope Thickness....................................................................................... 336
Table 16.19 WSP PF Strength Requirements (MPa) ................................................................. 337
Table 16.20 Pastefill Recipes ..................................................................................................... 338
Table 16.21 Estimated UG Fleet Requirements ........................................................................ 340
Table 16.22 Primary Ventilation Fans ....................................................................................... 341
Table 16.23 Underground Mining Schedule .............................................................................. 364
Table 16.24 Open Pit Mining Sequence .................................................................................... 366
Table 16.25 Open Pit Production Schedule ............................................................................... 368
Table 16.26 Contractor Open Pit Mining Fleet ......................................................................... 370
Table 16.27 Owner’s Mining Team ........................................................................................... 370
Table 16.28 Processing Schedule (Life of Mine) ...................................................................... 372
Table 17.1 Key Process Design Criteria .................................................................................... 375
Table 17.2 SART Major Process Design Criteria ...................................................................... 388
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Table 17.3 Reagent LOM Average Annual Consumptions ....................................................... 392
Table 17.4 Power Requirements and Energy Consumption ...................................................... 396
Table 17.5 Processing Fresh Water Estimated Consumption .................................................... 396
Table 17.6 LOM Annual Average Wear Consumable Rates ..................................................... 399
Table 18.1 Project Site Electrical Demand Electrical Demand ................................................. 405
Table 18.2 Description of On-Site Buildings ............................................................................ 408
Table 18.3 Details of Climate Stations Including Name, ID, Coordinates, and Distance to Site
.......................................................................................................................... 412
Table 18.4 Summary of Monthly Temperature, Precipitation, and Evaporation Data for
Hostotipaqullo Station ...................................................................................... 413
Table 18.5 Intensity-Duration-Frequency Values for Los Ricos South Site (mm) ................... 413
Table 18.6 Site Wide Water Balance for Los Ricos South Mine Site ....................................... 417
Table 19.1 Metal Prices and Exchange Rate ............................................................................. 430
Table 20.1 Surface Area of Project Components....................................................................... 432
Table 20.2 Permits and Procedures Required for the Approval of the Project .......................... 438
Table 20.3 Requirements by State-Municipal Authorities for the Project................................. 439
Table 20.4 ES Hydrological Balance ......................................................................................... 441
Table 20.5 Climate Types within ES ......................................................................................... 442
Table 20.6 Population in Age-Sex Relationship of the Municipalities of Hostotipaquillo,
Magdalena and Tequila, State of Jalisco .......................................................... 445
Table 21.1 Capital Cost Summary ............................................................................................. 449
Table 21.2 WBS Breakdown of Capital Costs........................................................................... 450
Table 21.3 Exchange Rates ........................................................................................................ 451
Table 21.4 Capital Cost Summary for Underground Portion of Project.................................... 452
Table 21.5 Initial Development Capital Costs ........................................................................... 453
Table 21.6 Capital Cost Estimate Summary - Process Plant, Tailings Management Facility
and On-Site Infrastructure ................................................................................ 455
Table 21.7 Cost Summary by Major Discipline ........................................................................ 455
Table 21.8 Mechanical Equipment Supply Price Basis ............................................................. 457
Table 21.9 Mechanical Equipment Package List ....................................................................... 457
Table 21.10 Electrical Equipment Supply Price Basis .............................................................. 458
Table 21.11 Electrical Equipment Supply Price Basis .............................................................. 458
Table 21.12 Material Commodity Codes ................................................................................... 459
Table 21.13 Construction Contract Packages ............................................................................ 459
Table 21.14 Capital Growth Allowance .................................................................................... 461
Table 21.15 SART Plant Capital Cost Estimate ........................................................................ 461
Table 21.16 Off-Site Infrastructure Capital Costs ..................................................................... 462
Table 21.17 Indirect Capital Costs............................................................................................. 463
Table 21.18 Owner’s Costs ........................................................................................................ 465
Table 21.19 Open Pit Sustaining Capital Costs ......................................................................... 467
Table 21.20 Operating Cost Summary....................................................................................... 468
Table 21.21 Underground Mining Operating Costs .................................................................. 469
Table 21.22 Underground Operating Development Costs ......................................................... 469
Table 21.23 Underground Operating Production Costs ............................................................. 470
Table 21.24 Paste Plant LOM Operating Cost Summary .......................................................... 471
Table 21.25 Open Pit Mining Operating Cost Summary ........................................................... 472
Table 21.26 Open Pit Contractor Unit Rates ............................................................................. 473
Table 21.27 LOM Processing Operating Cost Summary .......................................................... 475
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Table 21.28 Hourly Labour Requirements and Annual Salaries ............................................... 476
Table 21.29 SART Plant Operating Costs ................................................................................. 479
Table 21.30 General and Administration Costs ......................................................................... 480
Table 21.31 Closure Costs ......................................................................................................... 481
Table 22.1 Economic Evaluation Summary .............................................................................. 483
Table 22.2 Project Cash Flow Summary ................................................................................... 485
Table 22.3 Silver and Gold Price Sensitivity NPV, IRR and Payback ...................................... 486
Table 22.4 Capital and Operating Cost Sensitivity of NPV and IRR ........................................ 487
Table 22.5 Cash Flow Model Summary .................................................................................... 488
Table 25.1 Project Risks ............................................................................................................ 495
LIST OF FIGURES
Figure 1.1
Figure 1.2
Figure 1.3
Figure 4.1
Figure 4.2
Figure 5.1
Figure 5.2
Figure 6.1
Figure 6.2
Figure 6.3
Figure 6.4
Figure 7.1
Figure 7.2
Figure 7.3
Figure 7.4
Figure 7.5
Figure 7.6
Figure 7.7
Figure 8.1
Figure 9.1
Figure 9.2
Figure 9.3
Figure 9.4
Figure 9.5
Figure 9.6
Figure 9.7
Figure 9.8
Figure 9.9
Figure 9.10
Figure 9.11
Los Ricos South Property Location Map ............................................................. 3
Longitudinal Projection of Underground Mine Design...................................... 18
Surface Infrastructure at Los Ricos South Project ............................................. 20
Property Location Map, Jalisco, Mexico ............................................................ 43
Los Ricos South Property Concessions .............................................................. 45
Access to the Los Ricos South Project ............................................................... 50
Cinco Minas Ejido Lands ................................................................................... 54
Photograph of the Cinco Minas Patio by H. E. Crawford, 1907 ........................ 56
Composite Level Plan Map of the Cinco Minas Mining Company
Underground Levels ........................................................................................... 59
Location of Stope Assay Data Found in the Gerard Papers at the University
of Montana.......................................................................................................... 63
Photograph of the El Abra Workings ................................................................. 66
Geological Map of Mexico* ............................................................................... 71
The Main Cenozoic Volcanic Provinces of Mexico........................................... 72
Map Showing the Structures and Volcanic Centres in the Hostotipiquillo
District, Mexico .................................................................................................. 74
Property Geology................................................................................................ 75
Structural Geological Setting of the Los Ricos South Project Area ................... 77
Photograph of Drill Core Showing Sulphide Minerals ...................................... 79
Photographs of the Los Ricos Vein in Drill Core, Hole LRGG-19-079 ............ 80
Hydrothermal Fluid System Model Showing Epithermal and Porphyry
Related Mineral Deposit Types .......................................................................... 82
Area Covered with the 2019 1 m Digital Terrain Model by PhotoSat ............... 84
Local Geology Map ............................................................................................ 86
Northern Geology Map EL Agulila Area ........................................................... 87
Central Geology Map El Abra Geology ............................................................. 88
Southern Geology Map Cerro Colorado Area .................................................... 89
Trench Location Map of the Los Lamas Area.................................................... 91
Trench Number 10, Los Lamas .......................................................................... 92
Trench Number 11, Los Lamas .......................................................................... 92
Northern Trench Map El Aguila Area ................................................................ 93
Central Trench Map El Abra .............................................................................. 94
Southern Trench Map Las Lamas....................................................................... 95
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Figure 9.12
Figure 9.13
Figure 9.14
Figure 9.15
Figure 9.16
Figure 10.1
Figure 10.2
Figure 10.3
Figure 10.4
Figure 10.5
Figure 10.6
Figure 10.7
Figure 10.8
Figure 10.9
Figure 11.1
Figure 11.2
Figure 11.3
Figure 11.4
Figure 11.5
Figure 11.6
Figure 11.7
Figure 11.8
Figure 11.9
Figure 11.10
Figure 11.11
Figure 11.12
Figure 11.13
Figure 11.14
Figure 11.15
Figure 11.16
Figure 11.17
Figure 11.18
Figure 11.19
Figure 11.20
Figure 11.21
Figure 11.22
Figure 11.23
Figure 12.1
Southern Trench Map Cerro Colorado ............................................................... 96
Location of El Troce Workings .......................................................................... 97
Sampling El Troce Workings ............................................................................. 98
Eagle and Main Area IP Geophysical Response ................................................ 99
Longitudinal Projection of Exploration Target ................................................ 100
Eagle Longitudinal Projection .......................................................................... 103
Eagle Cross-Section Projection Drill Holes 11, 12, 13, 14 .............................. 104
Eagle Cross-Section Projection Drill Holes 30, 31 .......................................... 105
Eagle Cross-Section Projection Drill Holes 32, 35 .......................................... 106
Eagle Cross-Section Projection Drill Hole LRGAG-22-118 ........................... 107
Longitudinal Projection – Main Area ............................................................... 109
Longitudinal Projection – Los Ricos South Resource Area with Pierce Points
.......................................................................................................................... 110
Main Area Cross-Section Projection Drill Hole LRGG-22-209 ...................... 111
Main Area Cross-Section Projection Drill Hole LRGG-22-218 ...................... 112
Performance of Au Field Duplicates for 2019 Drilling at Los Ricos South .... 118
Performance of Ag Field Duplicates for 2019 Drilling at Los Ricos South .... 119
Performance of Au Field Duplicates for 2020 Drilling at Los Ricos South .... 121
Performance of Ag Field Duplicates for 2020 Drilling at Los Ricos South .... 122
2019 Umpire Sampling Assays for Au ............................................................. 124
2019 Umpire Sampling Assays for Ag ............................................................. 124
2020 Umpire Sampling Assays for Au ............................................................. 125
2020 Umpire Sampling Assays for Ag ............................................................. 125
Performance of Au AR09003X CRM for Jul 2020 – May 2023 Drilling at
Project ............................................................................................................... 127
Performance of Au OxB130 CRM for Jul 2020 – May 2023 Drilling at
Project ............................................................................................................... 128
Performance of Au Oxi164 CRM for Jul 2020 – May 2023 Drilling at Project
.......................................................................................................................... 128
Performance of Au SN104 CRM for Jul 2020 – May 2023 Drilling at Project
.......................................................................................................................... 129
Performance of Ag SN104 CRM for Jul 2020 – May 2023 Drilling at Project
.......................................................................................................................... 129
Performance of Au SN118 CRM for Jul 2020 – May 2023 Drilling at Project
.......................................................................................................................... 130
Performance of Ag SN118 CRM for Jul 2020 – May 2023 Drilling at Project
.......................................................................................................................... 130
Performance of Au SQ88 CRM for Jul 2020 – May 2023 Drilling at Project . 131
Performance of Ag SQ88 CRM for Jul 2020 – May 2023 Drilling at Project . 131
Performance of Au Blank for Jul 2020 – May 2023 Drilling at Project .......... 132
Performance of Ag Blank for Jul 2020 – May 2023 Drilling at Project .......... 133
Performance of Au Field Duplicates for Jul 2020 – May 2023 Drilling at
Project ............................................................................................................... 134
Performance of Ag Field Duplicates for Jul 2020 – May 2023 Drilling at
Project ............................................................................................................... 135
2023 Umpire Sampling Assays for Au ............................................................. 137
2023 Umpire Sampling Assays for Ag ............................................................. 137
Results of the August 2019 Au Verification Samples (ALS)........................... 141
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Figure 12.2
Figure 12.3
Figure 12.4
Figure 13.1
Figure 13.2
Figure 13.3
Figure 13.4
Figure 13.5
Figure 13.6
Figure 13.7
Figure 13.8
Figure 13.9
Figure 14.1
Figure 14.2
Figure 14.3
Figure 14.4
Figure 14.5
Figure 14.6
Figure 14.7
Figure 14.8
Figure 14.9
Figure 14.10
Figure 14.11
Figure 14.12
Figure 14.13
Figure 15.1
Figure 15.2
Figure 15.3
Figure 15.4
Figure 15.5
Figure 15.6
Figure 15.7
Figure 15.8
Figure 16.1
Figure 16.2
Figure 16.3
Figure 16.4
Figure 16.5
Figure 16.6
Results of the August 2019 Ag Verification Samples (ALS)........................... 141
Results of the May 2023 Au Verification Samples (Actlabs) .......................... 142
Results of the May 2023 Ag Verification Samples (Actlabs) .......................... 142
Elemental Silver Deportment of the Eagle, Abra and CC Zone Samples ........ 155
Silver Mineral Deportation of the Eagle, Abra, and CC Zone Samples........... 156
Eagle and Abra Zone Pre-leach and Post-leach Underflow Yield Stress
Versus Underflow Solids Density .................................................................... 202
2024 Abra and Eagle Met Composite and Variability and Comminution
Sample Locations ............................................................................................. 212
CC Master Composite, Variability and Comminution Sample Locations ....... 213
Eagle, Abra and CC Testwork and Mine Plan Head Grade Box Plots............. 216
Eagle and Abra Zones Leach Residue Gold Grade Versus Calculated Head
Grade Plot ......................................................................................................... 221
Eagle and Abra Zone Leach Residue Silver Grade Versus Calculated Head
Grade ................................................................................................................ 222
Eagle and Abra Zone Leach Residue Copper Grade Versus Calculated Head
Grade Plot ......................................................................................................... 222
Drill Hole Plan View ........................................................................................ 228
Mineralized Veins ............................................................................................ 229
Sample Lengths ................................................................................................ 231
Log-Probability Plots for Abra Domain Composites ....................................... 235
Log-Probability Plots for Eagle Domain Composites ...................................... 236
Semi-Variograms for Ag Composites .............................................................. 237
Semi-Variograms for Au Composites .............................................................. 238
Longitudinal Projection Showing the Abra and Eagle Veins with Historical
Workings and Search Ellipsoid Orientations ................................................... 239
Swath Plots -Abra ............................................................................................. 247
Swath Plots – Eagle .......................................................................................... 248
Log-Probability Plots for Composites .............................................................. 254
Cerro Colorado Semi-Variograms .................................................................... 255
Swath Plots ....................................................................................................... 259
Abra Pit Optimization Results .......................................................................... 276
Cerro Colorado Pit Optimization Results ......................................................... 277
Abra Pit Design ................................................................................................ 278
Cerro Colorado Pit Design ............................................................................... 279
Plan View of the Los Ricos South Open Pits ................................................... 280
Abra Open Pit Showing Geotechnical Drill Hole Traces ................................. 281
Abra Orebody Trending Northwest-Southeast with an Average Dip Direction
of 230° and Showing Design Sectors ............................................................... 285
Summary of the Recommended Design Configurations for the Proposed Pit
Geometries ........................................................................................................ 289
Underground Mine Isometric View.................................................................. 294
Three Underground Mining Zones ................................................................... 296
Longitudinal Projection Showing Stope Types ................................................ 298
Longitudinal Projection showing Historical Mine Workings .......................... 301
Farmyard Portal Water Retention Dam ............................................................ 306
View Looking Northeast of Planned Underground Development Showing
Areas of the Mine that have no Data from the Geotechnical Investigation ..... 308
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Figure 16.7
Figure 16.8
Figure 16.9
Figure 16.10
Figure 16.11
Figure 16.12
Figure 16.13
Figure 16.14
Figure 16.15
Figure 16.16
Figure 16.17
Figure 16.18
Figure 16.19
Figure 16.20
Figure 16.21
Figure 16.22
Figure 16.23
Figure 16.24
Figure 16.25
Figure 16.26
Figure 17.1
Figure 17.2
Figure 17.3
Figure 17.4
Figure 17.5
Figure 17.6
Figure 17.7
Figure 17.8
Figure 17.9
Figure 18.1
Figure 18.2
Figure 18.3
Figure 18.4
Figure 18.5
Figure 18.6
Figure 18.7
Figure 18.8
Figure 20.1
Figure 20.2
Figure 20.3
Figure 20.4
Figure 20.5
Figure 22.1
Figure 24.1
Longitudinal Projection Showing Pillars.......................................................... 326
Longitudinal Projection of Stopes by Designed AgEq Cut-off Grade ............. 329
Extracting Stopes Adjacent to Historical Mine Workings ............................... 335
Longitudinal Projection of Ventilation System Year -1 ................................... 343
Longitudinal Projection of Ventilation System Year 4 .................................... 344
Longitudinal Projection of Ventilation System Year 8 .................................... 345
Longitudinal Projection of Ventilation System Year 12 .................................. 346
Longitudinal Projection of Mine Dewatering System ...................................... 348
Longitudinal Projection of Electrical and Compressed Air Systems ............... 350
Longitudinal Projection of Refuge Stations and Egress ................................... 352
Longitudinal Projection of Mine Plan Year -1 ................................................. 356
Longitudinal Projection of Mine Plan Year 1 .................................................. 357
Longitudinal Projection of Mine Plan Year 2 .................................................. 358
Longitudinal Projection of Mine Plan Year 3 .................................................. 359
Longitudinal Projection of Mine Plan Year 5 .................................................. 360
Longitudinal Projection of Mine Plan Year 8 .................................................. 361
Longitudinal Projection of Mine Plan Year 10 ................................................ 362
Longitudinal Projection of Mine Plan Year 15 ................................................ 363
Abra Phase 1 and Phase 2 Open Pits ................................................................ 366
Cerro Colorado Open Pits ................................................................................ 367
LRS Overall Process Flowsheet ....................................................................... 374
Overall Process Plant Layout ........................................................................... 376
Crushing Area Layout ...................................................................................... 377
Crushing and Grinding Layout ......................................................................... 379
Leach Circuit Layout ........................................................................................ 381
CCD Circuit Layout ......................................................................................... 382
SART Process Flow Diagram .......................................................................... 387
SART Plant General Layout ............................................................................. 389
Tailings Filtration Layout ................................................................................. 391
Overall Site Layout........................................................................................... 401
Access Roads to the Project Site ...................................................................... 403
Overhead Power Line Route to Project Site ..................................................... 406
Water Management Proposed Structure for the Los Ricos South Mine Site ... 415
DSTF Location and Footprint .......................................................................... 420
Most Critical Cross-Section of the DSTF Final Configuration Used in the
Stability and Seepage Analyses ........................................................................ 424
Stater DSTF Configuration............................................................................... 426
DSTF Closure Concept..................................................................................... 428
Environmental System of the Los Ricos South Project.................................... 431
Regional Historical Mine Locations ................................................................. 436
Close Regional Historical Mining Works ........................................................ 437
Average Monthly Precipitation, Station 14068 Hostotipaquillo ...................... 442
Maximum, Average and Minimum Temperatures ........................................... 443
Annual Cash Flow Profile ................................................................................ 486
Project Organization Chart ............................................................................... 491
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Important Notice
This Report was prepared as a National Instrument 43-101 Technical Report for GoGold
Resources Inc. (“GoGold” or “the Company”) by P&E Mining Consultants Inc. (“P&E”), with
input from Ausenco Engineering Canada ULC (“Ausenco”), D.E.N.M. Engineering Ltd.
(“D.E.N.M.”), BQE Water Inc. (“BQE Water”), WSP Canada Inc. (“WSP”) and Consultores
Interdisciplinario en Medio Ambiente S.C. (“CIMA”), collectively the Report Authors. The quality
of information, conclusions, and estimates contained herein is consistent with the level of effort
involved in the Report Authors’ services, based on i) information available at the time of
preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and
qualifications set forth in this Report.
To the extent that any statements of GoGold or opinions of an expert who is not a Qualified Person
have been relied upon for legal, political, environmental, or tax matters relevant to the Technical
Report by a Qualified Person, the specifics of that source and level or reliance have been cited
within the relevant portions of the body of this Report.
This Report is intended for use by GoGold subject to terms and conditions of its contracts with
each of the Report Authors. Except for the purposes legislated under Canadian provincial and
territorial securities law, any other uses of this Report by any third party are at that party’s sole
risk.
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Page xx
1.0
SUMMARY
1.1
INTRODUCTION
This National Instrument (“NI”) 43-101 Technical Report was prepared by P&E Mining
Consultants Inc. (“P&E”) for GoGold Resources Inc. (“GoGold” or “the Company”) to provide a
Feasibility Study (“FS”) and updated Mineral Resource Estimate for the Los Ricos South Project
(the “Project”), located in the State of Jalisco, Mexico. The Los Ricos South Property is 100%
owned by GoGold.
Input to this FS was also provided by Ausenco Engineering Canada ULC (“Ausenco”), D.E.N.M.
Engineering Ltd. (“D.E.N.M.”), BQE Water Inc. (“BQE Water”), WSP Canada Inc. (“WSP”) and
Consultores Interdisciplinario en Medio Ambiente S.C. (“CIMA”). This Technical Report has an
effective date of January 14, 2025.
GoGold is a corporation trading on the Toronto Stock Exchange (“TSX”) under the symbol
“GGD”. The Los Ricos South Property (the “Property”) is held by GoGold’s wholly-owned
Mexican subsidiary Minera CM Jalisco, S.A. de C.V. (“MCMJ”), previously known as Minera
Durango Dorado S.A. de C.V. (“MDD”). The Property comprises 16 concessions covering
11,994 ha in the Hostotipaquillo Region of Jalisco State, Mexico. GoGold executed an Option
Agreement for 29 concessions, which include the Project’s 11 concessions (and 18 concessions in
the Los Ricos North area) on March 26, 2019. Subsequently, on August 22, 2019, GoGold acquired
100% ownership of the 29 concessions, including the 11 concessions of the Project, and purchased
a 2% NSR royalty through a Concession Agreement. Since then, GoGold acquired five more
concessions: Siete Pozos, Cinco Minas (Eagle), El Cofre, San Pedro and San Pedro II.
Between 1914 and 1930, the Cinco Minas Mining Company produced 33.3 million ounces of silver
(33.3 Moz Ag), 233.5 thousand ounces of gold (233.5 koz Au) from 2.45 million short tons (2.45
MT) from underground workings along the Los Ricos Vein. The vein system strikes northwest–
southeast for a distance of 3,500 m on the Property, dips approximately 65° to the west, and varies
from 5 m to 30 m in width. Historical mining operations on the vein extended for a length of 450
m along strike and from surface down-dip for a distance of 850 m. The operators worked only the
high-grade sections of the vein in stopes 1 m to 5 m wide. Stopes were backfilled with development
waste rock consisting of “low-grade quartz vein material”.
GoGold initiated a due diligence program at the Project site in January 2019, which included
geological mapping, channel and trench sampling, diamond drilling to twin three, legacy, reverse
circulation (“RC”) drill holes, bulk density measurements, and analyzing for precious metals (gold
and silver).
Through the remainder of 2019 and to 2023, GoGold carried out an exploration program including:
purchasing the 1 m topographic Digital Terrain Model (“DTM”) for the Los Ricos South Property;
completing 515 diamond drill holes for 87,736 m; collecting and assaying drill core, surface grab
and underground channel samples; compiling the legacy mining data recovered from the
University of Montana; geological mapping of the Property; and initiating an environmental study
and social and community impact studies.
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P&E prepared a Technical Report without a Mineral Resource Estimate on the Los Ricos South
Property for GoGold with an effective date of August 20, 2019. P&E completed an Initial Mineral
Resource Estimate on the Los Ricos South Property for GoGold with an effective date of July 28,
2020, which formed the basis for an initial Preliminary Economic Assessment (“PEA”) with an
effective date of January 20, 2021. GoGold continued its exploration program to May 2023 and
evaluated the potential for near surface and underground gold, silver and copper mineralization.
An updated Technical Report on the Los Ricos South Mineral Resource Estimate and PEA was
completed by P&E with an effective date of September 12, 2023. Subsequently, site work has
consisted of geotechnical drilling, surveying of underground historical workings and preparation
of items such as electrical power and water supply.
1.2
LOCATION, CLIMATE, ACCESS AND INFRASTRUCTURE
The Los Ricos South Project is situated near the village limits of Cinco Minas, in the State of
Jalisco, Mexico (Figure 1.1). The village has a population of approximately 300 and can be easily
accessed on a well-maintained paved highway from the City of Guadalajara by travelling 85 km
northward on MEX 16D to Cuauhtémoc, then southward for 20 km on MEX 24. This is
approximately a two-hour drive. An international airport is located in Guadalajara with regularly
scheduled flights to the US and to Mexico City and other Mexican destinations.
Historical underground workings at the Los Ricos Vein are located at approximately latitude 21°
02’ 45” N and longitude 103° 56’ 08” W. The UTM coordinates (WGS84 Zone 13Q) are 610,600
m E and 2,327,600 m N.
The Village of Cinco Minas is situated at an elevation of approximately 1,520 masl and has an
altitude-moderated semi-temporal climate, with rainfall limited to heavy thunderstorms during the
hot summer months. The dry season extends from October to May, the days range from mild to
hot and nights from chilly to mild. Frost is common though not persistent in the winter.
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FIGURE 1.1
LOS RICOS SOUTH PROPERTY LOCATION MAP
Source: Modified by P&E (November 2024) from GoGold (2019)
Note: “A” in the main map and Los Ricos in the inset map represent the location of the Los Ricos South Property.
The warmest months are typically July to September and can be humid. Annual precipitation
ranges from 800 mm to 1,200 mm, with much of it associated with thunderstorms during July to
September.
The Hostotipaquillo District has a very long tradition of mining. There is an ample supply of skilled
personnel, equipment suppliers and contractors sufficient for the Project. An electrical power line
from the local grid crosses the Property, and water is available at a cost from the local water
commission. Telephone and cell phone coverage are excellent, as is access to high-speed internet.
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To date, exploration crews stay in Cinco Minas and make the short trip to site as required.
1.3
MINERAL TENURE
On August 22, 2019, GoGold announced it had entered into various agreements
(“the Concession Agreements”) to accelerate the acquisition of the 29 concessions that comprised
the Los Ricos area, including the Project’s 11 concessions, from a private Mexican owner.
With the signing of the Concession Agreements, GoGold was required to make payments as
follows:
•
US$500,000 upon signing the agreements;
•
US$3,220,000 was paid in equal monthly cash payments over 24 months from the
signing of the agreements; and
•
9,046,968 GoGold common shares were delivered over 24 months from the signing of
the agreements.
The Company also acquired a 2% NSR royalty for the Los Ricos South Property for payments as
follows:
•
US$1M in cash; paid in equal installments over 36 months; and
•
4,875,012 GoGold common shares delivered over 18 months.
There are now no royalties on the Property concessions.
In October 2022, the Company acquired the Eagle Concession, which connects the southern and
northern parts of Los Ricos South into a single, contiguous property and overlies the northern
continuation of the structure that hosts the main silver-gold mineralization. The terms of the
acquisition agreement involve payment of US$2.1M over four years to the vendor, Minera CM
Jalisco, S.A. de C.V. (“MCMJ”), the registered owner of the Eagle (Cinco Minas) Concession.
The Siete Pozos and El Cofre Concessions were acquired in 2020 and 2023, respectively.
Regarding San Pedro and San Pedro II, MCMJ is in full control of these concessions according to
the Exploration and Exploitation Agreement, with Assignment of Rights Option executed on
September 20, 2020. GoGold, through its subsidiary MCMJ, legally holds 100% of 13 of the 16
mining concessions that make-up the Los Ricos South Property, and is in full control of all 16
Titles. The 16 concessions comprise an area of 11,994 ha.
1.4
SURFACE RIGHTS
The Ejido of Cinco Minas owns the surface rights over the majority of the Project’s concessions,
and area covering the current Mineral Resource Estimate. On August 9, 2020, the Ejido of Cinco
Minas signed an Agreement with the Company for a period of 12 years with an additional 12-year
renewal period. The Agreement gives access to enter and carry out exploration and exploitation
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on 1,280 hectares for an annual fee of 1,000,000 Mexican pesos. When the Project is constructed,
the Company will pay a further annual fee of 500,000 Mexican pesos for use of up to 71 ha with
an additional annual fee of 7,000 Mexican pesos per hectare in excess of 71 ha. For the years the
Project is in production, there will be an additional annual fee of US$100,000.
1.5
ENVIRONMENTAL
According to the Concession Agreement, GoGold has not inherited any environmental liabilities
from the historical mining operations. All historical disturbances and environmental liabilities rest
with the State of Jalisco.
GoGold has retained Mexican environmental firm CIMA to complete baseline environmental and
socioeconomic studies of the Los Ricos South Project. CIMA carried out a baseline environmental
study of water quality, dust, noise, soil sampling, vegetation and other environmental issues. This
environmental assessment was required for submission to the authorities for permitting of a
commercial development, such as a mine and process plant. The Mexican Federal government
department responsible for environmental matters and permitting is the Secretary of the
Environment, Natural Resources and Fisheries (“SEMARNET”).
1.6
PERMITS
According to Mexican mining law, a series of permits are required to support and approve
exploration and mining activities. GoGold had a yearly permit from SEMARNET to allow drilling
and exploration activities valid through to March 19, 2023. The permit will not be renewed until a
future exploration program is planned.
A thorough examination of the permits and appropriate regulations is required to determine how
best to accommodate any Project development schedule.
1.7
GEOLOGY AND MINERALIZATION
The Cinco Minas Vein system at Los Ricos is a classic epithermal precious metal deposit, which
exhibits at least three phases of quartz and sulphide mineralization and deformation. The vein,
which is up to 30 m in width and has been traced over 3.5 km on the concessions, occurs along
a northwesterly trending structure that roughly marks the boundary between two calc-alkaline
magmatic arcs: 1) the older Sierra Madre Occidental volcanic province to the north;
and 2) the younger Trans-Mexican volcanic arc to the south. The age of these volcanics ranges
from Cretaceous (100 Ma) to Neogene (18 Ma) (previously Tertiary).
Mineralization consists of pyrite and chalcopyrite as the most abundant sulphides, with local
concentrations of galena, sphalerite, argentite, native silver, miargyrite (AgSbS2), and other silver
sulphosalts. Banded, milky, amethystine and brecciated quartz exhibit several periods of quartz
emplacement and intra-mineral deformation.
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1.8
EXPLORATION HISTORY
Silver and gold mining in the Hostotipaquillo area dates back to Spanish colonial times.
Small family-owned mining operations during the 1800s produced high-grade silver and gold
material from narrow underground workings at several locations on the Los Ricos South Property.
From 1908 to 1930, the Cinco Minas Mining Company (“CMMC”) produced 33,333,369 ounces
of silver and 233,495 ounces of gold from 2,446,040 short tons of mineralized material (Gerard,
1951). CMMC ceased operations in 1930. Exploration work was carried out by several groups in
the 1970s and 1990s, by TUMI Resources during 2002 to 2005, and Bandera Gold in 2006 and
2007. The Property was inactive from 2007 to late 2018 due to a protracted legal dispute.
GoGold commenced its exploration program in January 2019. The work has been undertaken
by GoGold staff and reputable Mexican consultants and contractors. Activities on the Project
include topographical surveying, satellite topographical mapping, geological mapping, sampling,
trenching, diamond drilling and assaying, structural mapping, and compilation of the historical
mining records of CMMC obtained from the Mansfield Library at the University of Montana,
Missoula, Montana.
Following acquisition of the Eagle Concession in October, 2022, GoGold completed preliminary
geological mapping, sampling, geophysical IP surveying and commenced a drill program. The
mapping and geophysical survey results indicated that the mineralized Eagle structure extends
>2.5 km to the northwest. Significant geophysical anomalies were identified as targets for
subsequent drill testing.
In addition to the exploration work completed by GoGold, the Authors established that the Los
Ricos South Abra and Eagle mineral deposits contain an additional Exploration Target of 1.8 to
2.2 Mt at 375 to 425 g/t AgEq for 22 to 30 Moz AgEq. The Exploration Target is based on the
estimated strike length, depth and thickness of the known mineralization. The potential quantities
and grades of the Exploration Target are conceptual in nature. There has been insufficient
work done by a Qualified Person to define this estimate as a Mineral Resource. The Company
is not treating this estimate as Mineral Resources, and readers should not place undue reliance
on this estimate. Even with additional work, there is no certainty that this estimate will be
classified as Mineral Resources or that it will ever prove to be economically recoverable.
1.9
DRILLING
GoGold’s drilling program on the Los Ricos South Project began in February 2019. As of the
effective date of this Technical Report, 524 drill holes totalling 89,380 m of HQ-size diamond drill
core have been completed by GoGold. As many as four drills operated on the programs. An
additional drill hole, LRGG-20-207, was a metallurgical test hole drilled in the Abra Vein. The
core from this drill hole was used for grinding testwork at SGS Lakefield and is not included in
the Mineral Resource Estimate.
The drilling has mainly concentrated on the Abra and Eagle Veins in the area of the historical
Cinco Minas underground mine. Drilling cross-sections are spaced at 50 m intervals along strike
between 0 N and 1,150 N. The drill holes are inclined at dips between -45° to -65° to test the veins
at 50 m intervals down-dip. All drill hole collars are surveyed with differential GPS and marked
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with concrete monuments. Downhole survey data are collected every 40 m to 50 m in order to
measure deviations in the drill holes. Drill core was transferred from the drill rig to a secure drill
core logging and storage facility. The drill core was logged using standard procedures and the
information captured and recorded in digital GVMapper™ software. Standardized logging forms
and geological legends have been developed for the Project. All drill core has been photographed
with digital cameras. All field data are forwarded daily via satellite to the Servicios y Proyectos
Mineros de México, S.A. de C.V. (“SPM”) server in Hermosillo, where it is checked, verified, and
incorporated into the main database. SPM is a geological consulting/contracting service company
that undertakes all field work for GoGold.
Drill core recoveries are excellent, although some drill core is lost around the historical
underground workings. Special attention is made to identify the backfill material placed in the
historical stopes and to model the mined-out and backfilled voids created by the stopes.
1.10
SAMPLING AND ASSAYING
Samples of the drill core typically average 1.0 m in length and were cut using a diamond saw.
QA/QC protocols are followed, including insertion of certified reference materials (“CRMs”),
blanks and duplicate pairs. Gold and silver assays are determined for high-grade samples with
gravimetric methods and use normal fire assay/atomic absorption methods. All samples are
prepared using the four acid digestion procedures to ensure sufficient digestion and accurate
reporting of silver values.
Multi-element data is collected using the ICP procedure. The Authors are satisfied with GoGold’s
sampling and assaying protocols on the Los Ricos South Project.
1.11
DATA QA/QC AND VERIFICATION
The Authors completed data verification by undertaking a site visit to the Los Ricos South Project
during GoGold’s mapping, trenching and drilling programs. Drill core handling, logging and
sampling procedures, QA/QC protocols, drill core recovery, and RQD and bulk density
measurements were observed and reviewed.
The Project data are stored in a Microsoft Access™ database. All geological data are entered
electronically in the field and forwarded via satellite communications to the SPM office in
Hermosillo, Sonora. Assays are received electronically from the laboratory and imported into the
database. Drill hole collar locations are manually entered into the database. Checks are routinely
performed on the survey, drill hole collar coordinates and assay data. Digital and paper records
are kept for all assay and QA/QC data.
Mr. Fred Brown, P.Geo., of P&E, a Qualified Person under the terms of NI 43-101, completed a
site visit to the Los Ricos South Project on August 15 and 16, 2019. A data verification drill core
sampling program was conducted as part of the on-site review. Mr. David Burga, P.Geo., of P&E,
a Qualified Person under the terms of NI 43-101, completed a site visit to the Los Ricos South
Property on May 16 and 17, 2023, and conducted a data verification drill core sampling program
as part of the on-site review.
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The Authors consider the due diligence results to be acceptable and results are suitable for
verification use in the Mineral Resource Estimate.
1.12
METALLURGICAL TESTING
1.12.1
Process Plant Testing
Four metallurgical test programs were conducted on samples from the Los Ricos South Deposit
between 2021 and 2024. SGS-Lakefield Research Ltd. (“SGS”), with support from Metso,
conducted the metallurgical testing programs in 2020-2021, 2021-2022, 2023 and in 2024.
The SGS test programs in 2020-2021, 2021-2022 and 2023 focused on determining comminution
characteristics for the Abra and Eagle Zones and on flowsheet development and selection (gravity
pre-concentration, flotation and whole ore leach). Initial investigations into optimizing leach and
zinc precipitation and SART performance were conducted. Both Metso and SGS performed solidliquid separation trials as part of these programs.
The test program completed by SGS in 2024 for this FS included Eagle, Abra and Cerro Colorado
(“CC”) Zone master composite, comminution and variability samples representing a range of
grade, depth, and zone parameters from the target mining areas. The testwork included head assay
analysis, master composite mineralogy and gold and silver deportment studies, comminution tests,
gravity concentration, whole ore cyanidation, Merrill Crowe precipitation, SART, cyanide
detoxification and solid-liquid separation tests.
The 2024 testwork samples provide representative spatial coverage of the Zones, particularly in
the early years of Eagle and Abra operation. They also cover the range of feed grades expected in
the mill feed according to the mine production schedule.
The major Eagle and Abra mineral is quartz (83.2% to 87.4%), with minor minerals being chlorite
(2.6% to 4.1%) and orthoclase (4.8% to 7.4%). The Eagle composites contained higher levels of
hematite (2.1% to 2.3%). The CC master composite predominant minerals were quartz (76.3%),
chlorite (6.1%), orthoclase (9.0%) and calcite (3.4%). Overall, there is no indication of smectites,
which are poor filtering clays.
Comminution testwork indicated the Eagle, Abra and CC Zones exhibit moderate competence for
breakage in a SAG mill with SMC Axb values ranging from 36.3 to 70.8. The SMC results indicate
Eagle is the most competent material, followed by Abra and then CC. The Bond ball mill work
index (“BWi”) results for the three Zones ranged from 15.2 kWh/t to 19.3 kWh/t, with the Abra
and Eagle samples displaying similar hardness. Abrasion testing indicated the Eagle is the most
abrasive deposit with results placing it within the 72nd to 93rd abrasiveness percentile according to
the SGS database.
Gravity concentration and gravity tailings leach testwork was completed on the Eagle high-grade
composite sample to assess the amenability of the ore to gravity pre-concentration as earlier testing
demonstrated potential of improved recoveries. Overall recovery for the tested gravity
concentration and gravity tailings leach flowsheet was 93% for gold and 89% for silver, which
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was equivalent to the recoveries in the corresponding whole ore leach test. The sample tested was
not amenable to gravity concentration, with a maximum gravity recoverable gold content of 5.1%.
The Eagle, Abra and CC samples tested responded well to cyanidation. Gold and silver extractions
at an 80% passing size (k80) of 75 µm (microns) and a 96-hour residence time ranged from 77% to
99% for gold and 74% to 95% for silver. Copper extractions ranged from 44% to 88% for Eagle
and Abra and 29% to 52% for CC. Leach sodium cyanide (NaCN) consumption ranged from 0.8
to 23.5 kg/t and lime consumption ranged from 0.1 kg/t to 1.0 kg/t. NaCN consumption increased
with increased copper head grades. The variability samples indicated that the Eagle gold recoveries
were negatively impacted by excluding pre-aeration prior to leaching, but the Abra and CC sample
gold recoveries were positively impacted by excluding pre-aeration.
One Eagle variability sample (CN-50) displayed uncharacteristic gold (76% to 79%) and minimal
silver extraction (2% to 3%) and requires further investigation to understand the reason for poor
metallurgical performance. Elemental head assay analysis of this variability sample indicated total
carbon and copper to be 2.25 times and 3.24 times the average of the Eagle Zone samples, which
is likely to be related to the relatively poor performance.
Merrill Crowe precipitation results demonstrated low gold and silver barren solution grades at zinc
addition stoichiometric values equal to 5:1 Zn:Au+Ag. The bench scale Eagle and Abra sample
extractions ranged from 65.0% to 100.0% for gold and 89.0% to 99.9% for silver. Approximately
56.0% of the gold and 80.0% of the silver precipitated in the single CC sample bench test and more
testwork is required to characterize the metallurgical performance of this Deposit in future
programs. A higher free cyanide solution concentration or stoichiometric ratio, which was tested
in previous campaigns, was not attempted in the FS campaign. Copper precipitation measured
between 0% to 13.9% in the bulk tests. One sample, characterized by the name “Abra low
composite” due to its lower Au and Ag content, was the only sample with a measured copper
extraction higher than 13.9% Cu, at 45.7% Cu. Merrill Crowe precipitation was included in the
test program primarily to generate barren solution for SART testing and does not represent plant
scale precipitation efficiencies.
SO2/air cyanide destruction tests indicate that for the Eagle Zone, the addition of 5:1 SO2:CNWAD
(weak acid dissociable) of sulphur dioxide (SO2) in combination with 1.9 g/g CaO:CNWAD of lime
was effective in reducing the cyanide concentration below the targeted threshold of 3 mg/L to 5
mg/L. For the Abra Zone, the test results showed for plant design, the addition of up to 8.1:1
SO2:CNWAD of SO2, 3.3 g/g CaO:CNWAD of lime and 0.07 g/g CuSO4:CNWAD of copper sulphate
was effective in reducing the cyanide concentration below the targeted threshold.
Dynamic settling tests were performed on Eagle and Abra pre-leach and post-leach tailings
samples for thickener design and sizing. Underflow densities greater than 55% w/w solids were
achieved at solids flux loading rates of 0.8 t/h⋅m2. Flocculant dosage rates of 10 g/t and 70 g/t were
used to achieve acceptable overflow clarities during the testwork. The overflow clarity of the postleach Abra tailings samples were double that of the Eagle sample and further optimization testwork
is recommended. Underflow yield stresses were generally less than 40 Pa to 50 Pa, however, some
high values were observed in the post-leach Abra sample, attributed to underflow ratholing affects
from the high flocculant addition.
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Pressure filtration testwork was performed on five detoxified tailings samples: two Eagle and three
Abra. Filter cake moistures of 14.8% to 16.1% were achieved at a 40 mm cake thickness, without
membrane squeezing. A filter cycle time of 11 to 12 minutes was estimated for production. The
Old Abra Zone sample displayed the most problematic dewatering characteristics.
All 2024 samples contained <20 g/t antimony, which did not affect leaching or recovery. Arsenic
levels were below detection levels in the 2021, 2022 and 2023 testwork but reached up to 112 g/t
and 182 g/t on two Abra samples from the 2024 program. The arsenic levels showed no adverse
effects on leaching or recovery. Mercury levels ranged from <0.3 g/t to 2.9 g/t, with no adverse
recovery effects. A mercury retort is recommended in the refinery area for mercury removal. To
manage copper concentrations in doré, the SART process and Merrill Crowe precipitate leach
circuit are recommended. No other deleterious elements were found at levels impacting gold-silver
payability or process performance.
Robust metallurgical projection models were derived from bottle roll and bulk leach tests for gold,
silver and copper, and from METSIM simulations. Based on the selected flowsheet for the plant
operation, over the range of grades (variability samples) tested, the nominal Eagle and Abra
recoveries for gold, silver and copper were estimated to be 92.6%, 85.5% and 65.4%, respectively.
For the CC Zone, which is introduced to the mill in production year 10, a fixed gold, silver and
copper recovery of 92.5%, 87.6% and 42.6% has been estimated.
1.12.2
SART Testing
For SART testing by SGS, the copper mineralization in the seven metallurgical composites was
very reactive with cyanide, leading to high cyanide consumptions and high concentrations of
copper in the Merrill Crowe barren solutions. Treatment of the Merrill Crowe barren by the SART
process will liberate the cyanide associated with the copper cyanide complex as free cyanide, for
recycling to leach, and will also produce a high-grade copper sulphide product (Cu2S) as a byproduct for sale. Testing of the SART process on Merrill Crowe barren solutions was very
successful. Free cyanide recoveries of 80% to 157% were achieved with 125% of the
stoichiometric addition of sodium hydrosulphide at pH 4 and 88% to 161% free cyanide recoveries
were achieved at 200% of the stoichiometric addition of sodium hydrosulphide at pH 4.
1.13
MINERAL RESOURCE ESTIMATE
The Mineral Resource Estimate database contains 615 drill holes of which 554 drill holes totalling
90,095 m were used for Mineral Resource modelling. The 554 drill holes consist of 518 diamond
drill holes (totalling 87,241 m) completed by GoGold and 36 historical reverse circulation holes
(totalling 2,853 m). Mineralization models were developed by GoGold and reviewed and modified
by the Authors. A total of eight individual mineralized domains were identified through drilling
and surface sampling. The modelled mineralization domains are constrained by individual
wireframes based on a nominal 0.30 g/t AuEq cut-off for low-grade domains and a nominal 3.0 g/t
AuEq cut-off for high-grade sub-domains. The average true thickness of the modelled Los Ricos
South Domains is 7.5 m. The average true thickness of the modelled Eagle Main Domain is 16.1
m with the high-grade sub domains, Eagle Plum FW and Eagle Plum HW, having an average true
thickness of 5.4 and 6.1 m, respectively. Mineralization wireframes were used as hard constraining
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boundaries for the purposes for block coding, statistical analysis, compositing limits and grade
estimation.
A three-dimensional sub-blocked model, with 3 m x 3 m x 3 m parent blocks and 1 m x 1 m x 1 m
sub-blocks, was used for the Mineral Resource Estimate. The block model consists of estimated
Au, Ag and Cu grades, bulk density, domain codes and classification criteria. AuEq and AgEq
block grades were subsequently calculated from the estimated Au, Ag and Cu grades, and
considered metal prices and metallurgical recoveries.
Assay samples were composited to a 1.0 m standard length. Au and Ag grades were estimated
using Inverse Distance Cubed interpolation from between one and 12 composites, with a maximum
of two composites per drill hole. Composites were capped by mineralization domain prior to grade
estimation. Composite samples were selected within a search ellipse oriented down the plunge of
identified high-grade trends. Individual bulk density values were applied to mineralized domains
separately and were determined using 4,515 measurements taken from drill holes.
Classification criteria were determined from observed grade, geological continuity and
variography. All blocks within 30 m of three or more drill holes were classified as Measured,
blocks within 60 m of two or more drill holes were classified as Indicated, and all additional
estimated blocks were classified as Inferred. For the purposes of classification, historical
underground channel sampling was treated as short drill holes for domain interpretation, however,
underground channel sampling was not used for grade estimation.
Pit-constrained Mineral Resources reported have been reported within an optimized pit shell and
include Measured, Indicated and Inferred Mineral classifications. Historical mining has been
depleted from the Abra Main Vein. Pit-constrained Mineral Resources are reported using a cut-off
of 40 g/t AgEq.
Out-of-pit Mineral Resources have been reported beneath the pit shell and are restricted to the
Eagle and Abra mineralized veins, which exhibit historical mineralized continuity and reasonable
potential for extraction by cut and fill, and longhole mining methods. Out-of-pit Mineral Resources
are reported using a cut-off of 130 g/t AgEq.
The Mineral Resource has an effective date of January 14, 2025, and is summarized in Table 1.1.
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TABLE 1.1
LOS RICOS SOUTH MINERAL RESOURCE ESTIMATE – PIT-CONSTRAINED AND OUT-OF-PIT (1-9)
Mining
Method
PitConstrained5
Pit - Cerro Co.6
Out-of-Pit7,8
Total
Classification
Tonnes
(M)
Measured
Indicated
Meas & Ind
Inferred
Indicated
Inferred
Measured
Indicated
Meas & Ind
Inferred
Measured
Indicated
Meas & Ind
Inferred
2.9
2.0
4.9
0.7
0.6
0.3
2.4
3.1
5.5
1.1
5.3
5.6
11.0
2.2
Au
(g/t)
1.13
0.74
0.97
0.66
0.90
0.87
2.00
2.25
2.14
1.48
1.53
1.59
1.56
1.11
Average Grade
Ag
Cu
AuEq
(g/t) (%)
(g/t)
150
0.03
2.96
107
0.03
2.07
133
0.03
2.60
108
0.03
2.00
43
0.01
1.43
34
0.01
1.28
218
0.14
4.79
187
0.26
4.80
201
0.20
4.79
127
0.51
3.61
181
0.08
3.79
144
0.15
3.50
162
0.12
3.64
106
0.27
2.70
AgEq
(g/t)
250
174
219
168
121
109
405
407
406
306
320
296
308
229
Au
(koz)
106
47
153
16
17
10
156
223
379
53
262
287
549
79
Contained Metal
Ag
Cu
AuEq
(koz) (Mlb) (koz)
14,065
1.7
278
6,747
1.4
130
20,812
3.1
408
2,552
0.5
47
787
0.1
26
377
0.1
14
17,025
7.5
373
18,525 17.4
476
35,550 24.9
849
4,544
12.4
130
31,090
9.2
651
26,059 18.9
632
57,150 28.1
1284
7,473
13.0
191
AgEq
(koz)
23,446
10,974
34,420
3,979
2,243
1,217
31,567
40,331
71,898
10,995
55,013
53,549
108,562
16,191
Notes:
1.
Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may
be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.
2.
The Inferred Mineral Resource in this estimate has a lower level of confidence than that applied to an Indicated Mineral Resource and must
not be converted to a Mineral Reserve. It is reasonably expected that the majority of the Inferred Mineral Resource could be upgraded to an
Indicated Mineral Resource with continued exploration.
3.
The Mineral Resources were estimated in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM
Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions
and adopted by the CIM Council.
4.
Historically mined areas were depleted from the Mineral Resource model.
5.
The pit constrained cut-off grade of 40 g/t AgEq was derived from US$1,850/oz Au price, US$23.75/oz Ag price, 86% Ag and 95% Au process
recovery, US$28/tonne process and G&A cost. The constraining pit optimization parameters were $2.10/t mineralized material and waste
mining cost, and 45-degree pit slopes.
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6.
7.
8.
9.
Cerro Colorado Resource constrained to open pit mining methods only; Out-of-pit Mineral Resources are restricted to the Eagle and Abra
mineralized veins, which exhibit historical continuity and reasonable potential for extraction by cut and fill and longhole mining methods.
The out-of-pit cut-off grade of 130 g/t AgEq was derived from US$1,850/oz Au price, US$23.75/oz Ag price, 86% Ag and 95% Au process
recovery, US$30/tonne process and G&A cost, and a $60/tonne mining cost. The out-of-pit Mineral Resource grade blocks were quantified
above the 130 g/t AgEq cut-off, below the constraining pit shell and within the constraining mineralized wireframes. Out–of-Pit Mineral
Resources are restricted to the Eagle and Abra Veins, which exhibit historical continuity and reasonable potential for extraction by cut and
fill and longhole mining methods
AgEq and AuEq were calculated at an Ag/Au ratio of 86:1 for pit constrained and out-of-pit Mineral Resources.
Totals may not sum due to rounding.
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1.14
MINERAL RESERVE ESTIMATE
The Los Ricos South Project will consist of both underground and open pit mining operations.
The Mineral Reserve for the Los Ricos South Project is summarized in Table 1.2 and has an
effective date of January 14, 2025. Approximately 44% of the Mineral Reserve is classified as
Proven.
Information that will affect the AgEq cut-off value used for estimating the Mineral Reserve
includes metal prices, the US Dollar:Mexican Peso exchange rate, dilution, overall mine and
process variable and fixed costs, process recoveries, copper concentrate quality targets, and
management of the operation. The Authors consider that the underground mining and open pit
methodologies, design criteria and parameters used are appropriate.
External factors, such as permitting, social, environmental governance (“ESG”) and political
issues could affect the viability of the Project. Therefore, these factors could materially impact the
Mineral Reserve estimate for the Project.
Total dilution in the underground mine plan is estimated at 32.1%. Overall average underground
mining loss for the Deposit is estimated at 8.8% by mass.
For open pit dilution, the Mineral Resource block model was regularized from a volume percent
model to a whole block selective mining unit (“SMU”) model. Regularizing resulted in waste and
ore portions of a block being combined into a single block value. Combined dilution and ore loss
during mining was estimated at 10% on tonnes and 9% on grade at the Abra Deposit. At the
narrower vein Cerro Colorado Deposit, combined dilution and ore loss was estimated at 30% on
tonnes and 20% on grade.
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TABLE 1.2
LOS RICOS SOUTH MINERAL RESERVE ESTIMATE (1-8)
Classification
Tonnes
(k)
Proven
Probable
Subtotal Underground
3,902
3,611
7,512
Proven
Probable
Subtotal Open Pit
Proven
Probable
Total Proven & Probable
580
2,140
2,720
4,482
5,751
10,233
187
152
170
Average Grade
Au
Cu
AgEq
(g/t) (%)
(g/t)
Underground
1.61 0.09
334
1.70 0.18
318
1.65 0.13
326
23,472
17,647
41,119
202
197
399
7.5
14.6
22.1
41,939
36,895
78,834
95
72
77
Open Pit
0.72 0.02
0.64 0.03
0.66 0.02
159
130
136
1,768
4,961
6,728
13
44
57
0.3
1.4
1.7
2,965
8,924
11,889
175
122
145
Total
1.49 0.08
1.31 0.13
1.39 0.10
312
248
276
25,240
22,607
47,847
215
241
457
7.7
16.1
23.8
44,904
45,819
90,723
Ag
(g/t)
Ag
(koz)
Contained Metal
Au
Cu
AgEq
(koz) (Mlb)
(koz)
Notes:
1.
Mineral Reserves are based on Measured and Indicated Mineral Resource Classifications only.
2.
Mineral Reserves are reported using the 2014 CIM Definition Standards and 2019 Best Practices Guidelines and have an effective date of
January 14, 2025.
3.
Mineral Reserves are defined within mine plans and incorporate mining dilution and ore losses.
4.
Open Pit Mineral Reserves are based on metal prices of US$23.75/oz Ag, US$1,850/oz Au and US$4.00/lb Cu, and are constrained within
optimized pit shells and designs that use 45–48º overall wall slopes, and process recoveries of 86% Ag, 95% Au and 51% Cu.
5.
An Open Pit cut-off grade of 46.4 g/t AgEq is estimated to differentiate ore from waste and is based on cost assumptions of US$26.22/t
processing, US$4.11/t site general and administrative, and 0.5% government mining tax on net revenue. Mining costs are estimated at
US$2.10/t of ore and waste rock.
6.
Underground Mineral Reserves are based on metal prices of US$23.75/oz Ag, US$1,850/oz Au and US$4.00/lb Cu, and are constrained
within a mine design, and use process recoveries of 86% Ag, 95% Au and 77% Cu.
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7.
8.
An Underground marginal cut-off grade of 150 g/t AgEq is estimated to differentiate ore from waste, and is based on cost assumptions of
US$34.93/t processing, US$4.46/t site general and administrative, and mining costs of US$41.93/t. An Underground economic cut-off grade
of 210 g/t AgEq is estimated to account for sustaining capital development costs of US$53.86/t in addition to those used to calculate the
marginal cut-off grade.
Totals may not sum due to rounding.
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1.15
MINING METHODS
The Los Ricos South Project will consist of both underground and open pit mining operations.
Production will commence with underground mining. Open pit mining will be initiated in YR 10
and there will be an underground / open pit overlap period of four years. The duration of all mining
activity will be 15 years after the start of production. The life-of-mine ore processing rate is 2,000
tpd or 720,000 tonnes per year (tpa).
Underground extraction of ore will be through longhole sublevel mining with cemented paste
backfill. The majority (~70%) of tonnes are extracted from previously unmined areas on the Eagle
and Abra Veins, with a minority of tonnes (~30%) coming from areas adjacent to historical mine
workings. The Cerro Colorado Vein was not evaluated for underground mining for the purposes
of this Technical Report since there were insufficient tonnage of Measured and Indicated Mineral
Resource above cut-off grade to justify underground mining. All underground mining will be
completed by contractor. Sublevels are spaced at a nominal interval of 20 m. Stopes are a
maximum of 15 m along strike, with width limited by vein thickness. Stope widths at Eagle average
13.6 m and at Abra average 9.3 m. Haulage truck capacity will typically be 30 t, matched with 10 t
load-haul-dump (“LHD”) units. Steady-state production of 2,000 tpd will be reached at the end of
YR 1 after a ramp-up period, and full production will be maintained until midway through YR 9,
after which underground production declines as open pit production begins. Over the life-of-mine
(“LOM”), a total of 7.51 Mt of ore will be recovered from the underground mine at average grades
of 1.65 g/t Au, 170 g/t Ag and 0.13% Cu, equivalent to 326 g/t AgEq and a total metal content of
78.8 Moz AgEq.
Figure 1.2 presents a longitudinal projection of the underground mine workings, both existing and
planned.
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FIGURE 1.2
LONGITUDINAL PROJECTION OF UNDERGROUND MINE DESIGN
Source: P&E (2025)
The Los Ricos South Deposits are near surface and lend themselves to conventional open pit
mining methods. A contractor will be engaged to mine the Abra, Cerro Colorado North and Cerro
Colorado South open pits. The contractor will undertake all drill and blast, loading, hauling, and
mine site maintenance activities. The Owner will provide overall mine management and technical
services. Mining will typically be done on 5 m high benches, using conventional equipment such
as hydraulic excavators, front-end loaders and 60 t haulage trucks. The overall strip ratio for open
pit mining is 7.6:1. Over the LOM a total of 2.7 Mt of ore will be mined from open pits, at average
grades of 0.66 g/t Au, 77 g/t Ag, 0.02% Cu equivalent to 136 g/t AgEq or 11.9 Moz AgEq. Waste
rock will be transported to nearby storage facilities, and ore will be hauled to the primary crusher
or placed on temporary stockpiles.
1.16
RECOVERY METHODS
A flexible and robust flowsheet has been conceived, which is designed to treat a variety of grades.
The flowsheet is based on metallurgical testwork, and Ausenco’s expertise and experience with
similar plants and local operations. The flowsheet is based on well-proven unit operations in the
industry and there are no unique or novel processing methods required for gold and silver recovery.
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The process plant is designed to produce doré as final product and a copper sulphide concentrate
secondary product. The process plant will be located at the mine site and will receive run of mine
(“ROM”) ore from the underground and open pit mines. The key Project design criteria for the
process plant are:
•
Major equipment, which has been designed for a nominal throughput of 2,000 tpd;
•
A crushing circuit with an availability of 70%, supported using a surge bin, dedicated
feeder and emergency stockpile to provide continuous feed to the grinding circuit and
balance of the process plant;
•
Milling and tailings filter circuits with availabilities of 92% and 85%, respectively;
•
Axb via SMC ore hardness characteristics of 37.8 and Bond ball mill work index of
19.0 kWh/t;
•
Design head grades of 2.98 g/t Au, 335 g/t Ag and 0.25% Cu; and
•
Overall process recoveries of 92.7% Au and 85.6% Ag for LOM average grades.
The process flowsheet includes semi-autogenous grinding (“SAG”), pre-leach thickening, cyanide
leaching, a five-stage counter-current decantation (“CCD”) circuit for pregnant solution recovery,
a Merrill Crowe circuit, precipitate acid digestion, smelting, a sulphidization, acidification, recycle
and thickening (“SART” process) circuit, tailings cyanide detoxification, thickening and filtration.
The total anticipated electrical demand for the process plant will be 6.1 MW. Fresh water for the
process facilities will be supplied from the La Labor (Martin) Dam, Los Huajes Dam and the site
contact water pond. Wherever possible in the plant, process water or barren solution will be used
to minimize fresh water consumption. Plant consumables will include hydrated lime, sodium
cyanide, lead nitrate, flocculant, coagulant, sodium hydroxide, sulphuric acid, sodium
hydrosulphide, diatomaceous earth, zinc powder, copper sulphate, sodium metabisulphite,
antiscalant and flux.
1.17
PROJECT INFRASTRUCTURE
Figure 1.3 presents a plan view of the surface infrastructure planned for the Project.
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FIGURE 1.3
SURFACE INFRASTRUCTURE AT LOS RICOS SOUTH PROJECT
Source: Ausenco (2025)
1.17.1
Dry Stack Tailings Facility
A dry stack tailings facility (“DSTF”) was chosen as the lowest risk tailings storage method for
the Project location. A DSTF was the preferred solution, considering the mine plan, the limited
availability of construction materials, local climactic conditions, and restricted space for
infrastructure. The tailings will be thickened and filtered to generate a filter cake with a moisture
content of approximately 14 to 18% before being transported by truck to the DSTF site.
The DSTF will develop in two main phases, starter and ultimate, with a total storage capacity of
approximately 5.3 Mt of tailings. Both phases will have bench slopes at a 2.5:1 (H:V) ratio. The
ultimate DSTF is designed to reach a maximum height of approximately 100 m, measured from
the lowest point of the ultimate buttress to the highest elevation of the dry stack.
Key components of the DSTF include:
•
Water Management Systems:
o Contact and non-contact water channels.
o A non-contact water underdrain collection system featuring perforated pipes.
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o A geosynthetic clay liner (“GCL”) will be combined with a 2.0 mm linear lowdensity polyethylene (“LLDPE”) double-sided textured (“DST”) geomembrane
liner. The contact water collection system will be positioned above the liner to
ensure effective drainage containment.
o Contact and non-contact water storage ponds.
•
Surface Water Diversion:
o Non-contact water diversion channels will be constructed to reduce the creation of
surface contact water within the DSTF area.
•
Infrastructure:
o Access roads to facilitate operations and maintenance.
This design approach ensures efficient storage while prioritizing environmental protections and
operational efficiency sustainability.
1.17.2
Infrastructure
The Project will begin with site preparation that involves clearing scrub brush, removing topsoil,
and developing essential infrastructure such as roads, stockpile areas, and platforms for mining
and processing facilities. These roads will be designed with appropriate drainage and access to key
areas of the plant, while site access will be controlled through security gatehouses. The Project’s
electrical power needs will be met through two 34.5 kV circuits, supported by emergency
generators to ensure continuous operation. Electrical systems will be distributed across the site,
with strategically located transformer stations and prefabricated modular electrical rooms to
optimize efficiency.
Additionally, the site will include a range of support buildings to house necessary operations,
including fuel stations, power generation, and water treatment facilities, with much of the process
plant designed for open-air use.
1.17.3
Water Management
The Project relies on two primary water sources, the Huajes and La Labor Dams, to supply raw
water to the site. Each dam has a permitted extraction rate and annual allocation, with infrastructure
in place to transport water through overland pipelines. A key component of the water management
plan involves handling both contact and non-contact water through diversion channels, collection
channels, and collection ponds, which help minimize runoff and allow for water reuse in the
process.
The Project site experiences a temperate climate with significant rainfall during the summer
months, and the water balance analysis indicates that the site will generally face a water deficit
throughout the year, requiring supplemental water from external sources. Water management
structures have been designed to manage surface runoff and ensure water availability for the
process plant while maintaining a zero-discharge system, where surplus water will be stored.
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1.18
MARKET STUDIES AND CONTRACTS
Metal prices are presented in Table 1.3 and are based on an average of approximate December 31,
2024, three-year monthly trailing averages, and Consensus Economics Inc. long-term projections
from various financial institutions. The Mexican Peso:US Dollar exchange rate is the approximate
average of the past four years. The metal prices and exchange rate are subject to spot market
conditions. There are no metals streaming or hedging agreements in place.
The Project will produce two saleable products:
•
•
Silver and gold doré bars; and
Copper precipitate (Cu2S).
TABLE 1.3
METAL PRICES AND EXCHANGE RATE
Item
Gold (US$/oz)
Silver (US$/oz)
Copper (US$/lb)
Exchange Rate (MXN:US$)
Price
2,330
26.80
4.00
20
Currently there is one contract in place that is material to the Project. GoGold has entered an
agreement with CFE, the national power company. A dedicated two circuit 41 km overhead power
line will be installed from the La Yesca Dam where CFE has a power generation plant and
electrical substation. The routing of the line will be from the La Yesca Substation through
Hostotipaquillo Municipally and then to the Project site. Two overhead power lines will connect
34.5 kV, three phase and 60 Hz via a sub-station located near the process plant. The installation
contract amounts to $8.0M of which $0.74M has been paid by GoGold, with $7.26M remaining.
1.19
ENVIRONMENTAL STUDIES, PERMITS, AND SOCIAL OR COMMUNITY
IMPACTS
Exploration activities have been conducted under annual permits from the Mexican government
Secretary of the Environment and Natural Resources (“SEMARNAT”). An extensive list of
Federal and State permits will be required before mining can commence, along with environmental
impact studies. GoGold engaged the Mexican firm CIMA to conduct an environmental baseline
study in 2019. CIMA (2024) has completed a socio-economic baseline study that considers
economic, cultural, social, demographic, and geographical aspects of the local communities. No
specific environmental liability caused by historical mining activity and related to the Project has
been identified by on-ground surveys. The mining Project is expected to represent minimal risk to
the processes and structure of the local ecosystem, and to the roads, livestock and agricultural
activities. The baseline study concluded that there are no archaeological zones within the
environmental system.
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Page 22 of 577
1.20
CAPITAL AND OPERATING COSTS
All costs are in Q4 2024 US dollars. No provision has been included in the cost estimates to offset
future escalation. An average contingency of approximately 10% has been added to capital costs
(“CAPEX”). No contingency has been added to operating costs (“OPEX”).
The initial capital cost estimate addresses the engineering, procurement, construction and start-up
of the Los Ricos South Project, with a construction period to develop an underground mine, build
a process plant capable of treating 2,000 tpd, prepare a dry stack tailings management area, and
install associated ancillary surface facilities. The initial estimated cost to design, procure, construct
and start-up the underground mine and process plant facilities described in this Technical Report
is $227M. The estimated cost includes a contingency allowance of $21M.
Sustaining capital represents expenses for additional costs that are not included in the normal
operating costs, equipment purchases that will be necessary during the operating life of the Project,
tailings storage capacity increases, all capital costs associated with underground mining, and open
pit mine pre-stripping and development costs. Life-of-mine (“LOM”) sustaining capital is
estimated at $100M.
Total capital costs over the LOM are estimated at $326M and are presented in Table 1.4.
TABLE 1.4
CAPITAL COST SUMMARY
Item
Process Plant Directs & Tailings
Underground Mine Development
Open Pit Pre-Stripping
Infrastructure
EPCM
Project Indirects
Subtotal
Contingencies (~10%)
Total
1
Initial
($M)
83.54
51.05
0.00
38.56
17.97
14.58
205.70
20.99
226.69
Sustaining
($M)
10.22
62.76
17.66
0.00
0.00
0.00
90.65
9.07
99.71
Total
($M)1
93.77
113.82
17.66
38.56
17.97
14.58
296.35
30.05
326.40
Totals may not sum due to rounding.
Total OPEX over the life-of-mine (“LOM”) are estimated at $915M which average $89.43/t
processed. The LOM average operating costs for the Project during production years is
summarized in Table 1.5. Labour positions and rates for the Los Rico South process plant are based
on employee rates at the GoGold Parral operation in Chuhuahua, Mexico.
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Page 23 of 577
TABLE 1.5
OPERATING COST SUMMARY
Area
Open Pit Mining
Underground Mining
Paste Backfill Plant
Subtotal Average LOM Mining
Processing
SART Plant
Dry Stack Tailings Facility
G&A
Total cost per tonne processed
Cost
($/t processed)
18.46
44.04
5.89
42.92
33.36
4.13
2.13
6.88
89.43
Cost
($/t mined)
2.64
A weighted electrical rate cost supplied by Commission Federal de Electricity (“CFE”) is
estimated at $0.129/kWh after considering demand and usage (distribution) costs.
The diesel price has been quoted by a Mexican fuel distribution firm to be $1.11/L.
The average operating cash cost over the LOM is estimated at US$11.59/oz AgEq. The all-in
sustaining cost (“AISC”) is estimated to average US$12.78/oz AgEq over the LOM including
closure costs.
1.21
ECONOMIC ANALYSIS
Cautionary Statement - The results of the economic analyses discussed in this section represent
forward-looking information as defined under Canadian securities law. The results depend on
inputs that are subject to a number of known and unknown risks, uncertainties and other factors
that may cause actual results to differ materially from those presented here.
The Los Ricos South Project economic evaluation conclusions are summarized in Table 1.6. At
base case metal prices of US$2,330/oz Au, US$26.80/oz Ag and US$4.00/lb Cu, the Project has
an estimated US$355M after-tax net present value (“NPV”) at a 5% discount rate (“NPV5%”),
and an after-tax internal rate of return (“IRR”) of 28%. The payback period is estimated to be 2.6
years from start of production from underground mining.
P&E Mining Consultants Inc.
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Page 24 of 577
TABLE 1.6
ECONOMIC EVALUATION SUMMARY
Item
NPV0% ($M)
NPV5% ($M)
NPV7% ($M)
IRR (%)
Payback period (years)
Pre-Tax
851
553
468
39
1.8
After-Tax
575
355
293
28
2.6
Table 1.7 provides further details on the Project cash flow.
TABLE 1.7
PROJECT CASH FLOW SUMMARY
Item
UG Ore Mined
OP Ore Mined
Unit
kt
kt
Value
7,512
2,721
Process Plant Ore
kt
10,233
Silver Grade
g/t
145
Gold Grade
g/t
1.39
Copper Grade
AgEq Grade
%
g/t
0.10
276
Item
Net Revenue
Initial Capital Cost
Sustaining Capital
Cost
Mining Operating
Cost
Process Operating
(incl. SART &
DSTF) Cost
G&A Operating Cost
LOM Operating Cost
Silver Recovery
%
86
Operating Cash Cost
Gold Recovery
%
93
Copper Recovery
%
68
Silver Price
US$/oz
26.80
Gold Price
Copper Price
Payable Silver Metal
Payable Gold Metal
US$/oz
US$/lb
Moz
koz
2,330
4.00
41.1
423.6
Payable Copper
Mlb
11.2
Payable AgEq
Moz
79.9
All-in Sustaining
Cost
Mine Life
Average Process
Plant Rate
Pre-Tax NPV (5%)
After-Tax NPV (5%)
Pre-Tax IRR
After-Tax IRR
Pre-Tax Payback
Period
After-Tax Payback
Period
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Unit
US$M
US$M
Value
2,099
227
US$M
100
$/t ore
42.92
$/t ore
39.63
$/t ore
$/t ore
US$/oz
AgEq
US$/oz
AgEq
Yrs
6.88
89.43
tpd
2,000
US$M
US$M
%
%
553
355
39
28
Years
1.8
Years
2.6
11.59
12.78
15
Page 25 of 577
Table 1.8 presents a metal price sensitivity analysis and Table 1.9 presents an OPEX and CAPEX
sensitivity analysis.
TABLE 1.8
SILVER AND GOLD PRICE SENSITIVITY NPV, IRR AND PAYBACK
Sensitivity
-25%
-18%
-10%
Silver Price (US$/oz)
Gold Price (US$/oz)
After-Tax NPV (5%) (US$M)
After-Tax IRR (%)
After-Tax Payback (years)
20
1,739
110
13.6
4.7
22
1,913
184
18.3
3.8
24
2,087
255
22.5
3.2
Base
Case
26.8
2,330
355
28.0
2.6
+12%
+23%
+34%
30
2,608
469
33.7
2.0
33
2,869
575
38.8
1.8
36
3,130
681
43.7
1.7
TABLE 1.9
CAPITAL AND OPERATING COST SENSITIVITY OF NPV AND IRR
Sensitivity
-20%
-10%
Operating Costs – NPV (5%) (US$M)
Operating Costs – IRR (%)
Capital Costs – NPV (5%) (US$M)
Capital Costs – IRR (%)
440
32.6
394
35.1
399
30.6
373
30.9
Base
Case
355
28.0
355
28.0
+10%
+20%
317
26.3
331
24.5
275
24.1
310
22.0
The after-tax base case NPV and IRR are most sensitive to metal prices followed by operating
costs and capital costs.
1.22
RISKS AND OPPORTUNITIES
A risk assessment workshop was carried out on January 17, 2025, with personnel from P&E,
Ausenco and GoGold in attendance. Overall Project risks are perceived by the Authors as
moderate. Eight high risks items were identified, consisting of issues on Project constructability,
open pit mining permitting, dry stack tailings storage, and historical underground mine workings.
The main opportunity is that the Deposits remain open along strike to the northwest and down dip,
and there is potential to increase the current Mineral Resource Estimate.
1.23
CONCLUSIONS
The base case underground and open pit mine plan is scheduled over a production period of 15
years. The Proven and Probable Mineral Reserve Estimate for the Los Ricos South Project totals
10.2 Mt at average grades of 145 g/t Ag, 1.39 g/t Au and 0.10% Cu, or 276 g/t AgEq for 90.7 Moz
AgEq. Approximately 44% of the Mineral Reserve is classified as Proven.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 26 of 577
Based on the work undertaken to date, as summarized in this Technical Report, and the conclusions
listed in Section 25, the base case operating plan is a viable and attractive development
opportunity. Overall Project risks are perceived as moderate.
Current base case economic analysis indicates that the Project is forecast to be profitable. It is
recommended that GoGold continue the Project development and permitting plans, and implement
the recommendations set out in Section 26 of this Technical Report.
1.24
1.24.1
RECOMMENDATIONS
Metallurgical Testwork
Further metallurgical testwork is recommended to optimize the current flowsheet and design
criteria in a detailed design bridging phase. An estimated cost for the program is $0.25M.
1.24.2
DSTF Geotechnical
It is recommended to pursue further data collection of the subsurface in the DSTF area. The
objective of this program is to investigate variability in soil horizons and confirm foundation
loading, to characterize and verify geotechnical-geological features, distribution, and preliminary
characterization of the geotechnical units, foundation materials, and groundwater level. Data
collected will investigate variability in soil horizons and confirm foundation loading. A
geotechnical investigation program is recommended, including seven drill holes and seven test
pits. It is recommended that vehicular road access be constructed to reach the circumference of the
DSTF area to facilitate the preparation of drilling platforms. An estimated cost for the program is
$1.05M for field investigations and $0.07M for laboratory testing, totalling $1.12M.
1.24.3
Mine Geotechnical
The completed underground mine geotechnical and hydrogeological program has known data gaps
due to the increased mine footprint during the advancement of the Project. To mitigate data
uncertainty and confirm geotechnical design assumptions further geotechnical drilling is required
on the order of a minimum of 6,000 m. An estimated cost for the program is $1.5M for field
investigations and engineering study.
There is limited geotechnical information available regarding the Cerro Colorado Open Pits. The
current recommendations are based on extrapolating the data which have introduced uncertainties
to the study. It is recommended to collect more data in these regions by drilling a minimum of four
oriented boreholes (two per pit) with full geotechnical logging and laboratory strength testing. An
estimated cost for the program is $0.2M for field investigations and engineering study.
P&E Mining Consultants Inc.
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Page 27 of 577
2.0
INTRODUCTION
2.1
TERMS OF REFERENCE
This NI 43-101 Feasibility Study Technical Report of the Los Ricos South Project (the “Deposit”
or the “Property”), located in the State of Jalisco, Mexico, was prepared by P&E Mining
Consultants Inc. (“P&E”) at the request of GoGold Resources Inc. (“GoGold” or the “Company”).
Input to the Feasibility Study (“FS”) was also provided by Ausenco Engineering Canada ULC
(“Ausenco”), D.E.N.M. Engineering Ltd. (“D.E.N.M.”), BQE Water Inc. (“BQE Water”), WSP
Canada Inc. (“WSP”) and Consultores Interdisciplinario en Medio Ambiente S.C. (“CIMA”).
The Property is held by Minera CM Jalisco, S.A. de C.V. (“MCMJ”), a Mexican subsidiary
company that is wholly-owned by GoGold. MCMJ has rights to 100% of the minerals ownership
of the Property through a Concession Acquisition Agreement with a private Mexican vendor.
GoGold is a public, TSX-listed, mining company trading under the symbol “GGD”, with its head
office located at:
Suite 1301, Cogswell Tower,
2000 Barrington Street,
Halifax, Nova Scotia
Canada
B3J 3K1
The Mineral Resource Estimate reported herein is based on current drilling results and appropriate
metal pricing, and is conformable to the “CIM Standards on Mineral Resources and Reserves –
Definitions (2014) and Best Practices Guidelines (2019)”, as referred to in National Instrument
(“NI”) 43-101 and Form 43-101F, Standards of Disclosure for Mineral Projects.
The authors (“Authors”) of this Technical Report (“Technical Report”) understand that the Report
will support the public disclosure requirements of GoGold and will be filed on SEDAR+ as
required under NI 43-101 disclosure regulations.
This Technical Report has an effective date of January 14, 2025.
2.2
SITE VISITS
Mr. Fred Brown, P.Geo., of P&E, a Qualified Person under the terms of NI 43-101, completed a
site visit to the Los Ricos South Project on August 15 and 16, 2019. A data verification drill core
sampling program was conducted as part of the on-site review.
Mr. César Villalobos, Eng., of CIMA, conducted site visits to the Los Ricos South Project on
September 2-8, 2019, November 21-28, 2019, September 19-23, 2020, and August 8-11, 2020. A
baseline environmental data program was reviewed when on site. Mr. Villalobos also visited the
Property on January 11-14 and August 4-6 in 2023 during times when CIMA conducted
environmental data collection.
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Page 28 of 577
Mr. David Salari, P.Eng., of D.E.N.M., a Qualified Person under the terms of NI 43-101,
completed a site visit to the Los Ricos South Property on May 15 and 16, 2021. The purpose of
the visit was to assess engineering aspects of the Project, including potential process plant
locations.
Mr. David Burga, P.Geo., of P&E, a Qualified Person under the terms of NI 43-101, completed a
site visit to the Los Ricos South Property on May 16 and 17, 2023, and conducted a data
verification drill core sampling program as part of the on-site review.
Mr. D. Gregory Robinson, P.Eng., of P&E, a Qualified Person under the terms of NI 43-101,
completed a site visit to the Los Ricos South Property on November 28 and 29, 2023. The purpose
of the visit was to assess engineering aspects of the Project, including potential underground mine
access locations.
Mr. Scott Cameron Elfen, P.E., of Ausenco, a Qualified Person under the terms of NI 43-101,
completed a site visit to the Los Ricos South Property on August 13, 2024. The purpose of the visit
was to assess engineering aspects of the Project mainly related to the process plant and surface
infrastructure.
2.3
PREVIOUS TECHNICAL REPORTS
P&E prepared a Standards of Disclosure for Mineral Projects Technical Report of the Los Ricos
South Project with an effective date of August 20, 2019. An Initial Mineral Resource Estimate on
the Project was prepared by P&E with an effective date of July 28, 2020. P&E prepared a Technical
Report and Preliminary Economic Assessment (“PEA”) on the Project with an effective date of
January 20, 2021. The 2021 PEA was based on the 2020 Initial Mineral Resource Estimate. P&E
also prepared an Updated Mineral Resource Estimate and PEA on the Project with an effective
date of September 12, 2023. The reports are referenced in the Reference section (Section 27) of
this Technical Report.
In addition, there are four other prior Technical Reports on the Los Ricos South Property; Nebocat
2002, 2004(a) and 2004(b) and Munroe 2006. All are listed in the Reference section of this
Technical Report.
2.4
SOURCES OF INFORMATION
The Authors carried out a review of all relevant parts of the available literature and documented
results concerning the Project and held discussions with technical personnel from the Company
regarding all pertinent aspects of the Los Ricos South Project. This Technical Report is based, in
part, on internal company Technical Reports, and maps, published government reports, Company
letters, memoranda, public disclosure and public information. The reader is referred to the sources
of data, citations for which are compiled in the Reference section of this Technical Report, for
further detail on the Project.
The most recent previous NI 43-101 Technical Report on the Los Ricos South Project was
completed by P&E with an effective date of September 12, 2023. This Report has been relied upon
for the historical, geological, exploration and drilling sections of the current Technical Report.
P&E Mining Consultants Inc.
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Page 29 of 577
Considerable historical exploration and mining activities were carried out in the area of the
Los Ricos South Property, starting with the Spaniards in the early seventeenth century, various
private landowners in the nineteenth and early twentieth centuries, the Cinco Minas Mining
Company from 1908 to 1930, the Mexican Mining Co-Operative from 1931 to 1953, the Minera
San Jorge (“MSJ”) from 1990 to 2016, TUMI Resources Inc. (in Joint Venture with MSJ)
from 2002 to 2005, and the Bandera Gold Ltd. (in Joint Venture with MSJ) from 2005 to 2007.
Table 2.1 presents the Authors and Co-authors of each section of the Technical Report,
who acting as Qualified Persons as defined by NI 43-101, take responsibility for those sections of
the Technical Report as outlined in Section 28 Certificates of Author. The Authors acknowledge
the helpful cooperation of GoGold’s management and consultants, who quickly addressed all data
and material requests, and responded openly and cooperatively to all questions.
TABLE 2.1
REPORT AUTHORS AND CO-AUTHORS
Qualified Person
Mr. Andrew
P.Eng.
Bradfield,
Employer
Sections of Technical Report
P&E Mining Consultants Inc.
Ms. Jarita Barry, P.Geo.
P&E Mining Consultants Inc.
Mr. Fred Brown, P.Geo.
P&E Mining Consultants Inc.
Mr. David Burga, P.Geo.
P&E Mining Consultants Inc.
Mr. D. Grant Feasby, P.Eng. P&E Mining Consultants Inc.
Mr. Eugene Puritch, P.Eng. P&E Mining Consultants Inc.
Mr. D. Gregory Robinson,
P&E Mining Consultants Inc.
P.Eng.
Dr. William Stone, P.Geo.
P&E Mining Consultants Inc.
Mr. David Salari, P.Eng.
D.E.N.M. Engineering Ltd.
Mr. Tommaso
Raponi, P.Eng.
Roberto Ausenco
ULC
Engineering
Mr. Scott Cameron Elfen, Ausenco
P.E.
ULC
Engineering
1.1, 1.4, 1.5, 1.8, 1.20-1.24, 2, 3,
15.1, 15.2, 15.4.1, 15.4.2, 15.4.315.4.5, 16.2, 16.3, 18.3.2, 18.10.2,
18.10.4, 19, 21.1, 21.2.1, 21.2.2.5,
21.2.2.8, 21.2.2.9, 21.2.3, 21.3.1,
21.3.4, 21.3.8, 21.4-21.6, 22, 25.1.1,
25.3.6, 25.3.7, 25.3.11-25.3.13
1.10, 1.11, 11, 12, 25.3.3
1.13, 14.1, 14.3-14.11, 14.12.1,
14.12.3-14.12.11, 25.1.4, 25.3.5
1.9, 10, 25.3.2
1.19, 20, 25.3.10
14.2, 14.12.2
1.14, 1.15, 15.3, 16.1, 16.1.1,
16.1.2.2-16.1.2.8,
16.1.3-16.1.10,
18.10.1, 21.2.2.2, 21.3.2, 25.1.4,
25.2.2
1.2-1.8, 4, 5, 6, 7, 8, 9, 23, 25.3.1
1.12.2, 13.2.1.8, 17.4.7, 21.2.2.4,
21.3.6
1.12.1, 1.16, 1.17.2, 1.24.1, 13.1,
13.2.1.1-13.2.1.7,
13.2.1.9,
13.2.1.10, 13.3-13.5, 17.1-17.3,
Canada 17.4.1-17.4.6, 17.4.8, 17.4.9, 17.517.7, 18.1, 18.2, 18.3.1, 18.3.3,
18.3.4, 18.4-18.9, 18.10.3, 21.2.2.1,
21.2.2.3, 21.2.2.6, 21.2.2.7, 21.3.3,
21.3.5, 24, 25.1.2, 25.1.3, 25.2.4,
25.3.4, 25.3.8, 25.3.9, 26.1
Canada 1.17.1, 1.24.2, 18.12, 21.3.7, 25.2.3,
26.2
P&E Mining Consultants Inc.
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Page 30 of 577
TABLE 2.1
REPORT AUTHORS AND CO-AUTHORS
Qualified Person
Mr. Jonathan Cooper, P.Eng.
Mr. James Smith, P.Eng.
2.5
Employer
Engineering
Ausenco
ULC
WSP Canada Inc.
Sections of Technical Report
Canada 1.17.3, 18.11
1.24.3, 15.4.2.1, 16.1.2.1, 26.3, 26.4
UNITS AND CURRENCY
US$ are used throughout this Technical Report unless stated otherwise. Terminology and
abbreviations used in this Technical Report are summarized in Table 2.2, metric conversions are
listed in Table 2.3 and units are in Table 2.4.
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
$
$M
°
°C
%
% w/w
µm
>
<
σci
2-D
3-D
AACE
AAS
ABA
ACSR
Actlabs
Ag
AgEq
AI or Ai
AISC
ALS
ANFO
ANP
AP
APF
Meaning
dollar(s)
dollars, millions
degree(s)
degrees Celsius
percent
percent weight per weight (for a solution, weight percent
concentration)
microns
greater than
less than
uniaxial compressive strength of the intact rock
two-dimensional
three-dimensional
Advancement of Cost Engineering (International)
atomic absorption spectroscopy
acid base accounting
aluminum conductor steel-reinforced (cable)
Activation Laboratories Ltd.
silver
silver equivalent
abrasion index
all-in sustaining cost
ALS Chemex part of ALS Global, ALS Limited
ammonium nitrate fuel oil (mixture)
Area for the Protection of Natural Resources
acid potential (generation)
Federal Public Administration
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Page 31 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
Au
AuEq
AuEq75
Ausenco
Author
Bandera
BFA
BGC
BLS
BQE Water
BWI or BWi
C
CaO
CaCO3
calc
CAPEX
CC
CC Pit
CCD
CCTV
CENAPRED
CFE
CIM
CIMA
CITES
Cl
CMMC
CMS
CN
CND
CNfree or CNfree
CNO
CNT or CNT
CNS
CNWAD or CNWAD
COG(s)
Company, the
CONAGUA
CONANP
Meaning
gold
gold equivalent
gold equivalent ratio at 75:1 Ag/Au ratio @ 100% process recovery
Ausenco Engineering Canada ULC
author of section of this Technical Report who is a Qualified Person
Bandera Gold Ltd.
bench face angle
BGC Engineering Inc.
barren leach solutions
BQE Water Inc.
bond work index, bond ball mill work index
carbon
calcium oxide
calcium carbonate
calculated
capital costs
Cerro Colorado
Cerro Colorado Open Pit
counter-current decantation
closed circuit television
National Center for Disaster Prevention
Commission Federal de Electricity (Federal Electricity Commission)
Canadian Institute of Mining, Metallurgy, and Petroleum
Consultores Interdisciplinario en Medio Ambiente S.C.
Convention on International Trade in Endangered Species of Wild
Fauna and Flora
chlorine
Cinco Minas Mining Company
cavity monitoring system
cyanide
cyanide destruction test
free cyanide
cyanate
total cyanide
thiocyanate
weak acid dissociable cyanide
cut-off grade(s)
GoGold Resources Inc.
National Water Commission
Comisión Nacional de Áreas Naturales Protegidas – the National
Commission of Natural Protected Areas
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 32 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
Concession Agreements
CoV
CPF
CRM(s)
CSR
CSS
Cu
Cu2S
CUSTF
CWI or CWi
D&F
DCOG
D.E.N.M.
Deposit, the
DSO
DST
DSTF
DTM
DWT
E
E-GRG
ECOG
EGM
EL
ELOS
EMP
Eng.
EPC
EPCM
ER
ES
ESG
ETJ
eq m
FA
FAR(s)
Fe
FEL
FID
FoS
Meaning
agreements made to acquire 29 concessions comprising the Los Ricos
area
coefficient of variation
chamber filter
certified reference material control samples
cyclic stress ratios
closed side setting
copper
copper sulphide
Change of Use of Soils in Forest Land, change in forest land use
crushing work index
drift and fill
development cut-off grade
D.E.N.M. Engineering Ltd.
Los Ricos South Project
Deswik Stope Optimizer
double-sided textured
dry stack tailings facility
digital terrain model
drop-weight test
east
extended gravity recoverable gold
economic cut-off grade
engineering geology model
elevation
equivalent linear overbreak/slough
Engineering Management Plan
engineer
engineer, procure, construct
Engineering, Procurement and Construction Management
environmental risk study
environmental system
environmental, social, and governance
Technical Justification Study
equivalent metres
fire assay
fresh air raise(s)
iron
front-end loader
final investment decision
factor of safety
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Page 33 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
FS
ft
FW
FWD
g
g/t
G&A
GCL
GIS
GoGold
GPS
GRG
H
H
H2SO4
H:V
h
ha
HCl
HCN
HDPE
HMW
HOV
HR
hr
HSEC
HW
HV
HVAC
IBC
ICP
ID
ID2
ID3
in
INAH
INEGI
IP
IRA
IRR
ISO
Meaning
Feasibility Study
foot, feet
footwall
footwall drift/drive
gram
grams per tonne
general and administration
geosynthetic clay liner
geographic information systems
GoGold Resources Inc.
global positioning system
gravity recoverable gold
height
Shannon diversity index
sulphuric acid
horizontal to vertical ratio
hour(s)
hectare(s)
hydrogen chloride
hydrogen cyanide
high density polyethylene
Historical Mine Workings
Hill of Value
hydraulic radius
hour(s)
Health, Safety, Environment, Community
hanging wall
high voltage
heating, ventilation, and air conditioning
intermediate bulk containers
inductively coupled plasma (test)
identification
Inverse Distance Squared
Inverse Distance Cubed
inch(es)
Instituto Nacional de Antropología e Historia
National Institute of Statistics and Geography
induced polarization
inter ramp angle
internal rate of return
International Organization for Standardization (ISO)
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Page 34 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
ISO/IEC
ITH
JB
k
k80
klb
kg
kgD.S/m2.h
km
kN/m3
koz
kPa
kt
ktpa
kW
kWh
kV
kVA or kVa
L
L/s
LBA
LE
level
LGDFS
LGEEPA
LGPGIR
LH
LHD
LLDPE
LOM
LR
LR Pit
LRS
m
M
m2
m3
Meaning
International Organization for Standardization (ISO) / International
Electrotechnical Commission (IEC)
in-the-hole
Jalisco Block
thousand(s)
80% passing size
thousands of pounds
kilograms(s)
kg of dry solids per square metre per hour
kilometre(s)
kilonewtons per cubic metre
thousand(s) ounces
kilopascal
kilotonnes
thousand tonnes per annum / year
kilowatt(s)
kilowatt hour
kilovolts
1,000 volt amps
litre(s)
litres per second
environmental baseline
limit equilibrium
mine working level referring to the nominal elevation (m RL), e.g.
4285 level (mine workings at 4285 m RL)
Ley General de Silvicultura Sostenible (General Law of Sustainable
Forest Development)
Ley General Del Equilibrio Ecológico y la Protección al Ambiente
(General Law of Ecological Balance and Environmental Protection)
(Ley General para la Prevención y Gestión Integral de los Residuos)
General Law for the Prevention and Integral Management of Waste
longhole
load-haul-dump
linear low-density polyethylene
life of mine
Los Ricos
Los Ricos Pit
Los Ricos South
metre(s)
million(s)
square metres
cubic metres
P&E Mining Consultants Inc.
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Page 35 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
m3/s
Ma
masl or m asl
MBBR
MC
MCC
MCMJ
MCOG
MDD
Metso Outotec
MFP
mg
mg/L
mH or m H
MIA
MIBC
min(s)
ML
Mlb
MLC
mL
mm
Mm3
Moz
MPa
MPR
MRE
MSJ
Mt
MT
Mtpa
MVB
mW or m W
MW
MWh
N
N’
N/A
NaCN
NAD83
NAG
Meaning
cubic metres per second
millions of annum
metres above sea level
moving-bed-biofilm-reactor
Merrill Crowe
motor control centre
Minera CM Jalisco, S.A. de C.V.
marginal cut-off grade
Minera Durango Dorado S.A. de C.V.
Metso Outotec Canada
membrane filter press
milligram(s)
milligram(s) per litre
metre(s) high
Mexican environmental impact statement
methyl isobutyl carbinol
minute(s)
metal leaching
millions of pounds
Cia. Minera Las Cuevas, S.A.
metre(s) long, length
millimetre
millions of cubic metres
million ounces
megapascal or one million pascals
Mexico Mining Public Registry
Mineral Resource Estimate
Minera San Jorge, S. A. de C. V.
mega tonne or million tonnes
million short tons
million tonnes per annum or year
Mexican Volcanic Belt
metre(s) wide
megawatt
megawatt hour
north
modified stability number
not applicable
sodium cyanide
North American Datum of 1983
net acid generation
P&E Mining Consultants Inc.
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Page 36 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
NaHS
NaOH
Net NP
NI
NN
NOM
NP
NPV
NSR
NW
OEGT
Old Workings
OP
OPEX
OW
OZ
oz
oz/T
P50
P80 or P80
P&E
Pa
PAG
PAX
Pb
PDC
PEA
PEMB
P.Eng.
PEP
PF
PFD
PFS
P.Geo.
PLS
PLT
PPA
ppb
ppm
Property, the
Meaning
sodium hydrosulphide
sodium hydroxide
net neutralization potential
National Instrument
Nearest Neighbour
Norma Oficial Mexicana (Official Mexican Standards)
neutralization potential
net present value
net smelter return
northwest
ecological ordering of the territory
historical workings backfilled with lower-grade mineralized material
open pit
operating costs
Old Workings on the Abra Vein
ore zone
ounce
ounces per ton
50% percent passing
80% percent passing
P&E Mining Consultants Inc.
Pascal
potentially acid generating
potassium amyl xanthate
lead
process design criteria
preliminary economic assessment
pre-engineered metal building
Professional Engineer
Project Execution Plan
paste fill, paste backfill
process flow diagram
pre-feasibility study
Professional Geoscientist
pregnant leach solution
point load testing
accident prevention plan
parts per billion
parts per million
the Los Ricos South Property that is the subject of this Technical
Report
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Page 37 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
Project, the
PSD
psi
Q’
Q1, Q2, Q3, Q4
QA
QA/QC
QC
R² or R-squared
RAR(s)
RC
ROM
RQD
RWI or RWi
S
S
S=
ST
SA
SAG
SART
scfm
SD
SDS
SEDAR
SEM
SEMARNET
SG
SGS or SGS Lakefield
SIORE
SM
SMBS
SMC
SMO
SMU
SO2
SO4S
SPM
ST or ST
Meaning
the Los Ricos South Project that is the subject of this Technical
Report
particle size distribution
pounds per square inch
rock mass quality
first quarter, second quarter, third quarter, fourth quarter
quality assurance
quality assurance/quality control
quality control
coefficient of determination
return air raise(s)
reverse circulation
run of mine
rock quality designation
rod mill work index, bond rod mill work index
south
sulphur
sulphide sulphur
total sulphur
Environmental System
semi-autogenous grinding (mill)
sulphidation, acidification, recycling, and thickening
standard cubic feet per minute
standard deviation
safety data sheet
System for Electronic Document Analysis and Retrieval
scanning electron microscope
Secretary of the Environment, Natural Resources and Fisheries
specific gravity
part of SGS Canada Inc., SGS Mineral Services Laboratory at
Lakefield, Ontario.
Information Subsystem on Ecological Planning
screened metallics
sodium metabisulphite
SAG mill comminution
Sierra Madre Occidental
selective mining unit
sulphur dioxide
sulphate
Servicios y Proyectos Mineros de México, S.A. de C.V.
total sulphur
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Page 38 of 577
TABLE 2.2
TERMINOLOGY AND ABBREVIATIONS
Abbreviation
standards or CRMs
std dev
SWCC
t
T
t/m2.h
t/m3
Technical Report
TIMA
TMF
TMVA
tpa
tpd
Tpd
TSX
TUMI
UAB
UG
UGA
UPS
US$
USCS
UTM
V or v
v/v
VFD
VOD
VWP
W
W
w:o
WAD
WBS
WGS84
WOL
WRSF
wt %
XRD
WSP
YR or yr
Zn
Meaning
certified reference material control samples
standard deviation
soil water characteristic curve
metric tonne(s)
short ton(s)
hourly tonnes per cubic metre
tonnes per cubic metre
this NI 43-101 Technical Report
TESCAN Integrated Mineral Analyzer
tailings management facility
Trans-Mexico Volcanic Arc
tonnes per annum or year
tonnes per day
short tons per day
Toronto Stock Exchange
TUMI Resources Inc.
Biophysical Environmental Units
underground
Environmental Management Unit
uninterrupted power supply
United States dollar(s)
Unified Soil Classification System
Universal Transverse Mercator grid system
volt(s)
volume/volume
variable frequency drive
ventilation-on-demand
vibrating wire piezometer
west
width
waste to ore ratio
weak acid dissociable
work breakdown structure
World Geodetic System 1984
whole ore leach
waste rock storage facility
weight percent
X-ray diffraction
WSP Canada Inc.
year
zinc
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 39 of 577
TABLE 2.3
METRIC CONVERSIONS
To Convert From
feet
metres
miles
kilometres
acres
hectares
grams
ounce (Troy)
tonnes (t)
short tons (T)
grams per tonne
ounces (Troy) per ton
To
metres
feet
kilometres
miles
hectares
acres
ounces (Troy)
grams
short tons
tonnes
ounces (Troy) per ton
grams per tonne
Multiply By
0.3048
3.281
1.609
0.621
0.405
2.471
0.032
31.103
1.102
0.907
0.029
34.285
TABLE 2.4
UNIT MEASUREMENT ABBREVIATIONS
Abbreviation
μm
$
$/t
%
% w/w
¢/kWh
°
°C
cm
d
ft
GWh
g/t
h
ha
hp
k
kg
kg/t
Meaning
microns, micrometre
dollar
dollar per metric tonne
percent sign
percent solid by weight
cent per kilowatt hour
degree
degree Celsius
centimetre
day
feet
Gigawatt hours
grams per tonne
hour
hectare
horsepower
kilo, thousands
kilogram
kilogram per metric tonne
Abbreviation
m3/s
m3/y
mØ
m/h
m/s
Mt
Mtpy
min
min/h
mL
mm
MV
MVA
MW
oz
Pa
pH
ppb
ppm
P&E Mining Consultants Inc.
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Meaning
cubic metre per second
cubic metre per year
metre diameter
metre per hour
metre per second
million tonnes
million tonnes per year
minute
minute per hour
millilitre
millimetre
medium voltage
mega volt-ampere
megawatts
ounce (troy)
Pascal
Measure of acidity
part per billion
part per million
Page 40 of 577
TABLE 2.4
UNIT MEASUREMENT ABBREVIATIONS
Abbreviation
km
kPa
kV
kW
Meaning
kilometre
kilopascal
kilovolt
kilowatt
Abbreviation
s
t or tonne
tpd
t/h
kWh
kilowatt-hour
t/h/m
kWh/t
L
L/s
lb
M
m
m2
m3
m3/d
m3/h
kilowatt-hour per metric
t/h/m2
tonne
litre
t/m
litres per second
t/m2
pound(s)
t/m3
million
T
metre
tpy
square metre
V
cubic metre
W
cubic metre per day
wt%
cubic metre per hour
yr
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Meaning
second
metric tonne
metric tonne per day
metric tonne per hour
metric tonne per hour per
metre
metric tonne per hour per
square metre
metric tonne per month
metric tonne per square metre
metric tonne per cubic metre
short ton
metric tonnes per year
volt
Watt
weight percent
year
Page 41 of 577
3.0
RELIANCE ON OTHER EXPERTS
The Authors have assumed, and relied on the fact, that all the information and existing technical
documents listed in the References section of this Technical Report are accurate and complete in
all material aspects. Whereas the Authors carefully reviewed all the available information
presented, the Authors cannot guarantee its accuracy and completeness. The Authors reserve the
right, however, will not be obligated, to revise the Technical Report and Conclusions, if additional
information becomes known to the Authors subsequent to the effective date of this Technical
Report.
Copies of the land tenure documents, operating licenses, permits, and work contracts were not
reviewed. Information on land tenure was obtained from GoGold and included a legal due
diligence title opinion dated January 14, 2025 supplied by GoGold’s Mexican legal counsel,
Mr. Pablo Méndez Alvídrez, of the firm EC Legal Rubio Villegas, located in Chihuahua, Mexico.
The Authors have relied on tenure information from GoGold and has not undertaken an
independent detailed legal verification of title and ownership of the Los Ricos South Project. The
Authors have not verified the legality of any underlying agreement(s) that may exist concerning
the licenses, GoGold’s Mexican subsidiary, or other agreement(s) between third parties, however,
has relied on and considers it has a reasonable basis to rely upon GoGold to have conducted the
proper legal due diligence.
Select technical data, as noted in the Technical Report, were provided by GoGold and the Authors
have relied on the integrity of such data. A draft copy of the Technical Report has been reviewed
for factual errors by the GoGold, and the Authors have relied on GoGold’s knowledge of the
Property in this regard. All statements and opinions expressed in this document are given in good
faith and in the belief that such statements and opinions are not false and misleading at the effective
date of this Technical Report.
P&E Mining Consultants Inc.
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Page 42 of 577
4.0
PROPERTY DESCRIPTION AND LOCATION
The Los Ricos South Property is located in the State of Jalisco, Mexico (Figure 4.1), approximately
75 km northwest of the City of Guadalajara and 25 km east-southeast of the Company’s Los Ricos
North Property. Historical underground workings at the Los Ricos Vein are located on the Property
at latitude 21° 02’ 45” N and longitude 103° 56’ 08” W (UTM coordinates in WGS84 Zone 13Q
projection are 610,600 m E and 2,327,600 m N).
FIGURE 4.1
PROPERTY LOCATION MAP, JALISCO, MEXICO
Source: Modified by P&E (November 2024) from GoGold (2019)
Note: “A” in the main map and Los Ricos in the inset map represent the location of the Los Ricos South Property.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 43 of 577
The 11,994 ha Property consists of 16 concessions as shown in Figure 4.2 and listed in Table 4.1.
The El Rico, Lucerito and Cinco Minas concessions cover the Mineral Resources described in
Section 14 of this Technical Report. All the concessions are in good standing as of the effective
date of this Technical Report.
P&E Mining Consultants Inc.
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Page 44 of 577
FIGURE 4.2
LOS RICOS SOUTH PROPERTY CONCESSIONS
Source: GoGold (2024)
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Page 45 of 577
TABLE 4.1
MINING CONCESSIONS OF THE LOS RICOS SOUTH PROPERTY (1-2)
No.
Lot
Title
Start
End
1
2
3
4
5
6
7
8
9
10
11
12
13
14
15
16
Total
Jamaica II
Jamaica III
Jamaica IV Fraccion I
Los Ricos
El Rico
Milagros
San Jorge
San Jorge V
Reduccion San Jorge VI
Consuelo
Lucerito
Siete Pozos
Cinco Minas
El Cofre
San Pedro
San Pedro II
244971
244972
244973
244974
244975
244976
244977
244978
244979
244991
244992
220479
221732
224982
226132
227415
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
08-Jul-16
12-Aug-03
19-Mar-04
05-Jul-05
18-Nov-05
16-Jun-06
05-Apr-54
05-Apr-54
05-Apr-54
19-May-55
06-Jun-55
06-Jun-55
18-Mar-54
12-Feb-54
18-Mar-54
18-Mar-54
18-Mar-54
11-Aug-53
19-Mar-54
04-Jul-55
17-Nov-55
15-Jun-56
1
Surface
(ha)
509.86
850.27
3,265.61
62.15
34.07
177.26
3,291.02
100.00
1,280.34
45.56
26.92
80.00
1,106.99
27.93
1,011.96
124.09
11,994.05
Titleholder
(100%)
Location
Minera CM Jalisco, S.A. de C.V.
Hostotipaquillo, Jalisco
José Jaimes Achutigui
Minera CM Jalisco, S.A. de C.V.
Tequila, Jalisco
Tequila, Jalisco
Ricardo Michel Alvarez
Ricardo Michel Alvarez
Concessions information effective as of the January 14, 2025 date of the Title Legal Opinion for the Los Ricos South Property by El Rubio.
P&E Mining Consultants Inc.
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Page 46 of 424
Liens Status
free
free
free
free
free
free
free
free
free
free
free
free
free
free
free
free
active
active
active
active
active
active
active
active
active
active
active
active
active
active
active
active
4.1
OPTION AGREEMENT
GoGold began discussions with the private Mexican owner of the 29 concessions that were
collectively named Los Ricos (A in Figure 4.1) and the parties subsequently entered into a 60-day
due diligence agreement in January 2019. GoGold mobilized a field team and diamond drill to the
properties and began confirmation drill holes in early February 2019.
On March 26, 2019, GoGold signed an option agreement to acquire the 29 Los Ricos area
concessions from the private Mexican owner. As part of the option agreement, the Company was
obligated to make the following payments:
•
•
•
•
•
•
•
Initial upfront payment of $70,000;
Monthly payments of $12,000 for the first 12 months;
Monthly payments of $20,000 for months 13 to 24;
Monthly payments of $30,000 for months 25 to 36;
Monthly payments of $31,500 for months 37 to 60;
2% net smelter return royalty on five of the concessions; and
If the Company elects to exercise the option, a lump sum payment not to exceed $11M
can be made at any time within the option period.
During the five-year period, the Company would have exclusive exploration rights to the Los Ricos
area concessions. The Option Agreement was terminated when GoGold entered into the
Concession Agreements.
4.2
THE CONCESSION AGREEMENT
On August 22, 2019, GoGold announced it had entered into various agreements
(“the Concession Agreements”) to accelerate the acquisition of the 29 concessions that comprise
the Los Ricos area properties in Jalisco, Mexico from a private Mexican owner.
With the signing of the Concession Agreements, GoGold was required to make payments as
follows:
•
•
•
$500,000 in cash upon signing;
$3,220,000 in cash paid in installments over 24 months; and
9,046,968 GoGold common shares to be delivered in equal numbers over 24 months.
In conjunction with the signing of the Concession Agreements, the option agreement previously
entered into by the Company to acquire the 29 concessions in the Los Ricos area was terminated
(refer to press release dated March 26, 2019). As of the date of this Technical Report, all payments
have been completed in accordance with the Concession Agreements, and the Company holds title
to the 29 concessions included in the Concession Agreements through a wholly-owned subsidiary.
In addition to the Concession Agreements, the Company entered into an agreement to acquire the
existing 2% NSR royalty for the Los Ricos properties for payments as follows:
•
$1M in cash; paid in equal installments over 36 months; and
P&E Mining Consultants Inc.
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Page 47 of 424
•
4,875,012 GoGold common shares to be delivered in equal numbers over an
18-month period.
As of the date of this Technical Report, all payments related to the NSR royalty were completed,
and there are now no royalties on the Property concessions.
Since the previous Technical Report, five more concessions have been added to the Los Ricos
South Property: 1) Siete Pozos; 2) Cinco Minas (Eagle); 3) El Cofre; 4) San Pedro; and 5) San
Pedro II . The Siete Pozos Concession as added according to the Exploration and Exploitation
Agreement with Assignment of Rights option by Minera CM Jalisco, S.A. de C.V. (“MCMJ”), a
wholly-owned subsidiary of GoGold, previously known as Minera Durango Dorado S.A. de C.V.
(“MDD”), that was executed on November 4, 2020. The Agreement stipulates that payments
totalling $2.8M pesos (plus VAT = 16%) be made by MCMJ to the concession owner (Mr. José
Jaimes Achutigi) over a period of 24 months, as follows: $800,000 pesos on the effective date of
the Agreement; and $83,333.50 pesos per month from the effective date for 24 months. As of the
effective date of this Technical Report, all the payments have been made in full; however, MCMJ
has yet to file the request to transfer ownership.
The Eagle Concession was acquired by MCMJ on October 18, 2022 (listed as Cinco Minas in
Table 4.1). The rights to the Eagle Concession were acquired through an agreement to pay US$2.1
million over four years to the previous titleholders. Eagle connects the Company’s concessions in
the south portion of the Los Ricos Project to its northern concessions. As a result, all the Companycontrolled concessions in Los Ricos South area are now contiguous (see Figure 4.2).
The El Cofre Concession was acquired by MCMJ via an Assignment Agreement with the previous
titleholders on February 9, 2023. Payment of $2.4M pesos was made on that date for 100% of the
rights of ownership protected by the Title.
Regarding San Pedro and San Pedro II, MCMJ is in full control of these concessions according to
the Exploration and Exploitation Agreement, with Assignment of Rights Option executed on
September 20, 2020.
The Authors reviewed a legal title opinion on all GoGold’s Los Ricos South properties provided
by Mr. Pablo Méndez Alvídrez of the firm EC Legal Rubio Villegas located in Chihuahua, Mexico,
for GoGold dated January 14, 2025. The title opinion indicates that GoGold, through its subsidiary
MCMJ, legally holds 100% of 13 of the 16 mining concessions that make-up the Los Ricos South
Property, and is in full control of all 16 Titles.
4.3
LOS RICOS SOUTH PROJECT
The Los Ricos South Project, which is the subject of this Technical Report, was launched by
GoGold in March 2019. GoGold has been exploring and drilling primarily in the Main Deposit
area and on the adjacent Eagle Concession. According to Mexican mining law, permits are required
to support and approve exploration and mining activities. GoGold had a permit from the Secretary
of the Environment and Natural Resources (“SEMARNET”) that allowed for drilling and
exploration activities, which was valid through to March 19, 2023. The permit was not renewed,
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 48 of 577
because that phase of exploration was complete. New SEMARNET permits will be required to
cover any further drilling and exploration activities.
4.4
AUTHOR COMMENTS ON SECTION 4
To the extent known and described above, the Author is not aware of any other significant factors
or risks that may affect access, title or right or ability to perform work on the Los Ricos South
Property.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 49 of 577
5.0
ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND
PHYSIOGRAPHY
5.1
ACCESSIBILITY
The Los Ricos South Project is located near the Village of Cinco Minas, in the State of Jalisco,
Mexico. The Project is located within the Municipality of Hostotipaquillo of the Valles Region of
the State of Jalisco, approximately 85 km northwest of the City of Guadalajara. The nearest village,
for provision of goods and services, is the Town of Hostotipaquillo, head of the Municipality,
located approximately 15 km west of the Project (Figure 5.1).
Access to the Project site is by the Federal Road No. 15, from Guadalajara to Tepic. At km 88 a
road leads north to the Town of Hostotipaquillo. From Hostotipaquillo, it is approximately 20 km
southeast by a gravel road to the Village of Cinco Minas. It is an approximate two-hour drive from
Guadalajara to the Village of Cinco Minas and the Los Ricos South Project. There is an
international airport in Guadalajara with daily scheduled flights to the US and to Mexico City and
other Mexican destinations.
FIGURE 5.1
ACCESS TO THE LOS RICOS SOUTH PROJECT
Source: Modified by P&E (November 2024) from Google Earth
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Page 50 of 577
5.2
CLIMATE
Cinco Minas is situated at an elevation of approximately 1,200 masl and has an altitude-moderated
temporal climate with rainfall limited to heavy thunderstorms during the hot summer months. The
dry season extends from October to May, day temperatures range from mild to hot, and night
temperatures from chilly to mild. Annual precipitation ranges from 800 to 1,200 mm, much of it
associated with thunderstorms during the warm months of June to August.
Temperature characteristics, maximum, average, minimum and precipitation, were taken from the
National Weather Service's network of stations. The closest weather station to the Project is the
number 14068 Hostotipaquillo, located in the Municipality of Hostotipaquillo, approximately 15
km from the Project at an elevation of 1,300 masl. The temperature fluctuates in a range of 10°C
to 35°C, where the coldest month is January and the warmest is June (Table 5.1). It is expected
that any future mining operations will be conducted year-round.
TABLE 5.1
ANNUAL TEMPERATURE RANGES FOR CINCO MINAS
Month
January
February
March
April
May
June
July
August
September
October
November
December
5.3
Minimum
Temperature
(°C)
11.2
10.9
11.1
12.0
13.0
14.3
13.9
14.4
14.2
13.4
12.3
11.1
Maximum
Temperature
(°C)
29.7
30.5
31.5
32.5
33.2
33.0
31.0
30.5
30.8
31.1
30.2
29.7
Average
Temperature
(°C)
20.5
20.7
21.3
22.2
23.1
23.6
22.4
22.4
22.5
22.2
21.3
20.4
LOCAL RESOURCES AND INFRASTRUCTURE
The Village of Cinco Minas has a population of approximately 300. There is an adequate labour
source in the Village of Cinco Minas and in nearby Hostotipaquillo. The exploration and drilling
crews stay in the Village of Cinco Minas and make the short trip to site as required. Telephone and
cell phone coverage is good, as is access to high-speed internet.
The closest service centre is the Town of Magdalena, approximately a one-hour drive from Cinco
Minas. Magdalena lies 78 km northwest of Guadalajara. The municipality covers an area of
445 km². It borders the State of Nayarit to the west and the Town of Tequila to the east. As of
2018, the municipality had a total population of 16,214.
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The City of Guadalajara has a population of 7.5 million people and is the second largest city in
Mexico. The area has a long tradition of underground mining and there is an ample supply of
skilled personnel, equipment, suppliers and contractors sufficient for the Los Ricos South Project.
Electrical power is available from the local grid (Commission Federal de Electricity). A 220 kV
transmission line crosses the Property just to the south of Cinco Minas. There is water flowing
from the main adit of the Cinco Minas Mine and some streams flow with water all year. The Rio
Santiago (Santiago River) is located a few kilometres to the north at an elevation of 1,000 masl.
5.4
PHYSIOGRAPHY, FLORA AND FAUNA
The vegetation is variable across the Property and related to the topography and elevation.
Larger trees tend to be found along the valley floors, where the roots can access ground water.
Thorn bushes, cacti, grasses and various shrubs and vines are found on the hillsides.
5.4.1 Encino Forest
The Oak Forest is the type of vegetation in the middle and lower parts of the locality, at altitudes
between 1,800 and 1,500 masl, and is an important component of the environmental system.
Various species of oak of the genus Quercus are located here, and in a smaller percentage, pines
of the genus Pinus. Given the ecotonal conditions, there are also pines of higher altitudes and
shrubs from the low jungle present in the area.
5.4.2 Secondary Vegetation of Low Deciduous Forest
The area is in a state of ecological succession. There are indications that the original vegetation
was removed or disturbed heavily and is in a state of recovery.
5.4.3
Induced Pasture
Pasture vegetation is composed of grasses that arise when the original plant cover is removed.
Grassland usually appears as a result of clearings in any type of vegetation. It can also be
established in abandoned agricultural areas. This grassland is present in the upper parts of the sierra
and some lower areas, occupying areas of original forest and vegetation. This condition is
generally artificially maintained, and prevents the natural succession of vegetation that originally
occupied those places.
The animal sustaining capacity of these areas is low, due to the lack of management practices,
such as: the establishment of the medium seed pasture, weed control, diseases and pests,
fertilization, scheduled cuts or grazing, planting distance, irrigation, renewal of paddocks, etc.
In addition, the areas that support induced grassland have varying degrees of deterioration, due to
the management of the grass that replaced the original vegetation, in addition to the uncontrolled
grazing practices that led to soil loss (INEGI, 2003).
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5.5
SURFACE RIGHTS
The Ejido of Cinco Minas owns the surface rights over the majority of the Project’s concessions,
and covers the Mineral Resources. On August 9, 2020, the Ejido of Cinco Minas signed an
Agreement with the Company for a period of 12 years with an additional 12-year renewal period.
The Agreement gives access to enter and carry out the exploration and exploitation on 1,280 ha
for an annual fee of 1.0M Mexican pesos (Figure 5.2). When the Project is constructed, the
Company will pay a further annual fee of 0.5M Mexican pesos for use of up to 71 ha with an
additional annual fee of 7,000 Mexican pesos per hectare beyond 71 ha. For the years the Project
is in production, there will be an additional annual fee of US$100,000.
5.6
AUTHOR COMMENTS ON SECTION 5
The Author is of the opinion that there are no obvious impediments to building a mine, processing
or tailings facility within the area of the concessions.
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FIGURE 5.2
CINCO MINAS EJIDO LANDS
Source: GoGold (2023)
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6.0
HISTORY
Mineral exploration and mining for precious metals in the Hostotipaquillo area started in the early
seventeenth century. The Town of Hostotipaquillo became the regional centre for the mining
camps at Monte del Favor, San Pedro Analco and Cinco Minas.
6.1
EARLY HISTORY
Crawford (1908) states the Cinco Minas Property was known as an “Antigua” or ancient mine,
having been opened originally by the early Spaniards, and operated successfully by them until
forced to leave the country. The workings on the Destajos, Famosa and Trinidad levels of the
Cinco Minas Vein are associated with the early workers. The next documented record of
exploitation in the area was in 1824, when a Coronel Schiaffino held the Property (Nebocat, 2004).
6.2
1860 TO 1907 - THE MARTINEZ AND MONTERO MINE
From approximately 1860 to 1907, the Cinco Minas Property was owned by private Mexican
owners Mr. Luis Martinez and Mr. Montero of Guadalajara. The Property was noted for having
produced several bonanzas in its past production. Crawford (1908) reported the early production
records of the Mine was not kept with sufficient accuracy to use as data; however, from 1901 to
1907 complete reports were obtainable, as summarized in Table 6.1. Production was from narrow
underground stopes on the Destajos, Famosa, Trinidad and Tunnel levels of the San Dimas and
San Nicolas Chutes.
From the Cinco Minas and San Dimas Chute, Crawford (1908) reports 7,559 tons (“T”) of
mineralized material were treated on the “patio” (Figure 6.1) between 1901 and 1907, containing
an average of approximately 48.5 ounces (“oz”) of silver (“Ag”) (returns from bullion and
concentrates). Crawford (1908) also reports 1,070 tons of direct shipping mineralized material
were mined between 1901 and 1907 and graded an average of approximately 166.5 oz/T Ag.
Crawford (1908) also reports 5,942 tons of mineralized material averaging 57.8 oz/T Ag
(returns from bullion and concentrates) from the San Nicolas bonanza was processed on the patio
between 1901 and 1907. He also reports the owners produced 1,696 tons of direct shipping
material from the San Nicolas Bonanza averaging approximately 174.5 oz/T Ag.
Crawford (1908) states “of the 13,401 tons of ore treated on the patio, approximately 40% of the
gold was recovered”. The gold value obtained from shipping mineralized material, bullion and
concentrates was $73,000. At a gold price of US$20/oz Au, a total of 3,650 oz Au were produced
at a head grade of 0.272 oz/T Au.
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FIGURE 6.1
PHOTOGRAPH OF THE CINCO MINAS PATIO BY H. E. CRAWFORD, 1907
Source: GoGold (2019)
TABLE 6.1
HISTORICAL SILVER PRODUCTION FROM 1901 TO 1907
Mine
Method
Cinco Minas
Cinco Minas
San Nicolas
San Nicolas
Total Silver Ounces
Patio
Direct
Patio
Direct
Tons
(T)
7,559
1,979
5,842
1,696
Ag Grade
(oz/Ton)
48.5
166.5
51.8
174.5
Ounces
(oz)
366,611.50
329,503.50
302,615.60
295,952.00
1,294,682.60
Source: Crawford (1908)
6.3
1908 TO 1930 - THE CINCO MINAS MINING COMPANY
The Cinco Minas Mining Company (“CMMC”) was a private firm formed in 1905 by three
members of the Marcus Daly family; James Watson Gerard served as President, Marcus Daly Jr.
served as Secretary and Gerard’s wife, the former Mary Daly, who was the daughter of copper
magnate Marcus Daly, head of the Anaconda Copper Mining Company that developed the mines
of Butte, Montana.
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The firm hired a young mining engineer, Henry E. Crawford of Los Angeles, California to
“find a property” for the company. In 1907, Crawford began his review and due diligence of the
Martinez and Montero silver and gold mine at Cinco Minas in Jalisco, Mexico.
On Crawford’s recommendation, the CMMC purchased the Property from the Martinez and
Montero families in 1908 and began a six-year program to modernize the mine, sink a shaft, outline
mineralization, and build a modern flotation and cyanide process plant. In January 1914,
CMMC began processing mineralized material from the mine at a rate of 300 Tpd (tons per day).
The mine operated until March 1930, when it closed due to low silver prices related to the
Great Depression and civil unrest in Mexico. The production data are summarized in Table 6.2.
The mine and process plant operated at a daily production rate between 300 Tpd to 400 Tpd during
the First World War. However, shortages of manpower, equipment, power and supplies curtailed
operations at times throughout this period. In 1918, CMMC purchased new grinding mills for the
process plant and the production was raised to 550 Tpd.
During the life of the mine, CMMC developed and mined Cinco Minas Vein at the El Abra
mineralized shoot on nineteen levels (see Figure 6.2) down-dip for a distance of 840 m and
horizontally over 450 m near surface and over 120 m horizontally at the deepest level.
From 1914 to 1930, CMMC produced 33,700,000 oz Ag and 234,500 oz Au from 2,446,000 tons
of mineralized material (Gerard, 1951).
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TABLE 6.2
PRODUCTION HISTORY OF THE CMMC COMPANY 1914 TO 1930
Year
Tonnes
(t)
1914
1915
1916
1917
1918
1919
1920
1921
1922
1923
1924
1925
1926
1927
1928
1929
1930
Total
58,690
56,190
96,225
131,131
137,941
192,811
158,484
153,467
167,553
153,407
161,383
148,755
129,468
180,045
158,601
114,856
23,455
2,222,462
Tons
Processed
(T)
64,695
61,939
106,070
144,547
152,054
212,538
174,699
169,168
184,695
169,102
177,894
163,974
142,714
198,466
175,600
126,607
21,278
2,446,040
Silver
(g/t)
Silver
(oz/T)
Gold
(g/t)
Gold
(oz/T)
515.0
464.9
447.1
502.6
474.9
347.7
352.0
479.1
466.9
519.9
606.2
518.9
469.7
462.5
476.0
412.8
484.8
469.0
15.02
13.56
13.04
14.66
13.85
10.14
10.27
13.97
13.62
15.16
17.68
15.14
13.70
13.49
14.24
14.04
16.49
13.78
4.01
3.46
3.43
4.05
4.11
2.85
2.51
3.26
3.02
3.23
3.57
3.09
3.09
3.09
3.17
3.14
3.52
3.27
0.117
0.101
0.100
0.118
0.120
0.083
0.073
0.095
0.088
0.094
0.104
0.090
0.090
0.090
0.091
0.107
0.120
0.096
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Contained
Silver
(kg)
30,224.1
26,123.8
43,021.2
65,910.7
65,502.9
67,032.8
55,788.8
73,523.6
78,237.2
79,751.9
97,826.4
77,191.6
60,813.5
83,274.3
75,494.1
52,261.0
10,315.0
1,042,292.7
Contained
Gold
(kg)
235.4
194.6
329.9
530.5
567.5
549.4
398.3
499.9
505.5
495.5
575.4
459.0
399.5
555.6
502.6
397.0
75.0
7,270.7
Page 58 of 424
Contained
Silver
(oz)
971,718.9
839,892.8
1,383,152.8
2,119,059.0
2,105,947.9
2,155,135.3
1,793,634.6
2,363,818.3
2,515,361.2
2,564,059.8
3,145,165.9
2,481,746.5
1,955,181.8
2,677,306.3
2,500,000.0
1,778,023.7
350,936.9
33,700,141.9
Contained
Gold
(oz)
7,569.3
6,255.8
10,607.0
17,056.5
18,246.5
17,661.9
12,805.4
16,071.0
16,253.2
15,929.4
18,501.0
14,757.7
12,844.3
17,861.9
16,000.0
13,506.7
2,551.7
234,479.3
FIGURE 6.2
COMPOSITE LEVEL PLAN MAP OF THE CINCO MINAS MINING COMPANY
UNDERGROUND LEVELS
Source: GoGold (2019)
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6.4
JAMES WATSON GERARD PAPERS
The James Watson Gerard Papers is the collection of personal and professional papers,
photographic materials, and scrapbooks generated and/or collected by James W. Gerard.
These materials present a substantial documentation of Gerard’s activities as Ambassador to
Germany leading up to World War I, American Democratic Party activist, New York City
philanthropist, and international mining industry investor. Gerard married Mary Daly, daughter of
Marcus Daly, in 1901. During his lifetime, he maintained an interest in the Montana properties
and investments of the Daly family, and had a ranch of his own north of Hamilton, Montana.
After he died in 1951, his family donated his papers to the Archives and Special Collections of the
Maureen and Mike Mansfield Library at the University of Montana located in Missoula, MT.
The collection is divided into eleven series, and Series VIII is the Cinquo (sic) Minas Mining
Company papers, 1897 to 1942, 13.0 linear feet and seven oversize volumes.
The series contains extensive records of mining operation at the Cinquo (sic) Minas site in Jalisco,
Mexico. James Gerard was a major investor in the mining company, along with his
brother-in-law, Marcus Daly Jr. Most of the material was created or collected by
Henry E. Crawford, Cinquo (sic) Mina General Manager, to address Gerard’s questions and
concerns. Generally, these records exceed the standard range of investor reports and reflect the
importance mine management placed on Gerard’s investment.
The scope of the materials includes incoming and outgoing correspondence, auditor and tax
records, production and profit reports, property maps and structural blueprints, legal depositions
and real estate descriptions, capital inventories and financial volumes. Of particular note,
two folders in Box 470 includes copies of documents issued by the State of Jalisco, Mexico
detailing title descriptions and operation expansion applications for the Cinquo (sic) Minas Mining
Company between 1901 and 1911.
David Duncan, P. Geo., of GoGold and Jose Carlos Flores of SPM visited the Maureen and Mike
Mansfield Library in March 2019 and spent a week reviewing the correspondence, reports and
maps in the James Watson Gerard Series: VIII Papers. The items available are listed in Table 6.3.
High-resolution digital images of the reports and maps were scanned by the staff from the Archives
and Special Collections in the Maureen and Mike Mansfield Library. The scanned images were
subsequently processed by staff of Servicios y Proyectos Mineros de México, S.A. de C.V.
(“SPM”) from April to September to geo-reference and digitize the historical mine workings,
stopes and underground assay data. SPM is a geological consulting/contracting service company
that does all the field work for GoGold.
This information was then entered into SPM’s Minesight 3-D modelling software, along with other
data including surface topography, mapping and sampling, legacy drilling and underground
sampling, and the current exploration mapping, sampling and drilling by GoGold. The channel
sample information found is listed in Table 6.4 and high-grade mined areas in underground stopes
are shown in Figure 6.3.
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TABLE 6.3
REPORTS AND MAPS OF THE CINCO MINAS COMPANY 1908 TO 1930
Information 1908 1909 1910 1911 1912 1913 1914 1915 1916 1917 1918 1919 1920 1921 1922 1923 1924 1925 1926 1927 1928 1929 1930
Monthly
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
Reports
Annual
yes
yes
yes
yes
yes
yes
yes
yes
yes
Report
Yearly
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
Production
Geological
yes
yes
yes
yes
yes
yes
yes
Report
Metallurgical
yes
yes
Reports
Mill
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
Recoveries
Report
on
yes
yes
yes
yes
yes
yes
yes
yes
Reserves
Mine
Plan
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
yes
Maps
Mine
Longitudinal
yes
yes
yes
yes
yes
yes
yes
yes
yes
Map
Source: Gerard (1951)
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TABLE 6.4
CHANNEL SAMPLES WITH SILVER AND GOLD
ASSAYS IN THE CINCO MINAS MINE 1908 TO 1930
CMMC Mine
Level Name
Famosa (aka Lucerito)
San Juan
Destajos
Trinidad
Cinco Minas (aka Tunnel)
100
200
300
400
500
600
750
900
1050
1200
1350
1500
1650
1800
Total
Number of
Samples
121
266
641
139
1,552
703
559
1,744
117
no data
1,100
921
327
no data
19
1
572
no data
no data
8,782
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FIGURE 6.3
LOCATION OF STOPE ASSAY DATA FOUND IN THE GERARD PAPERS AT
THE UNIVERSITY OF MONTANA
Source: GoGold (2019)
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6.5
1931 TO 1980 - SEVERAL COMPANIES
In late 1930, local Mexicans formed a mining co-operative and exploited the Deposit at a reduced
scale until 1953, sending the hand sorted mineralized material to custom process plants (Nebocat,
2002).
During 1952-53, a successor company, Nueva Cinco Minas S.A. and Cia. El Aguila (the adjacent
property on strike to the northwest) continued exploration in the area. They focused on the southern
extension of the Cinco Minas Vein near the Cerro Colorado Prospect.
Munroe (2006) states that P. S. Friesen visited the Property in 1968 and concluded there
“was substantial tonnage for the development on a large scale, low-grade, open pit operation”.
The Property was inactive for several years until Cia. Minera Las Cuevas, S.A. (“MLC”) operated
an exploration program between 1981 and 1982. Nebocat (2002) reports that their work included
surface and underground mapping, sampling and diamond drilling along an approximately 300 m
long segment of the Cinco Minas Vein. Nebocat (2002) reports that J. R. Black visited the
underground workings at El Troce in January 1981 and reported that MLC had been conducting a
sampling program.
Nebocat (2002) reports MLC completed several drill holes on each of the El Aguila and Cinco
Minas properties during 1981 and 1982. The Authors have not seen any of the reports to which
Nebocat refers and they are not considered in this Technical Report.
6.6
1990 TO 2016 - MINERA SAN JORGE, S.A. DE C.V.
Minera San Jorge S.A. de C.V. (“MSJ”) is a private Mexican-owned company that acquired the
Cinco Minas properties during the 1990s. Munroe (2006) reports that MSJ retained an independent
consulting geologist Craig Byington to visit the Property and review the work carried out by MLC.
Nebocat (2002) reports MSJ commissioned Henkle and Associates in April 1999 to compile
statistics for samples collected in the El Abra area workings by the MLC staff in 1982.
MSJ retained a Dr. Cuellar in 1998 to perform a Mineral Resource/Mineral Reserve Estimate of
the El Abra, San Pedro, San Juan and Cerro Colorado Zones. The San Juan Mine was sampled for
MSJ by Rosas Haro in February 1997. Rosas Haro also sampled the crosscuts along the Lucerito
(Famosa) haulage-way. It was subsequently found that he sampled mainly footwall quartz–
rhyolite, which is very low-grade. An examination of the haulage-way showed that the crosscuts
were mostly filled in with debris when the co-operative open pit mined the Cinco Minas Vein.
In 2001, MSJ approached FIFO, a Mexican government agency charged to help develop mining
in the country. FIFO collected and forwarded some samples to the Geological Survey of Mexico
for assays, metallurgical testing, and mineralogical studies (polished sections).
6.7
2002 TO 2005 - TUMI RESOURCES INC.
On October 15, 2002, TUMI Resources Inc. (“TUMI”) signed an option and option agreement
with MSJ to earn up to 60% interest in the Cinco Minas Property. TUMI completed a thorough
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technical review and three separate drilling campaigns, which are summarized in Table 6.5
(Nebocat, 2004). TUMI terminated the Option Agreement with MSJ on May 18, 2005.
TABLE 6.5
SUMMARY STATISTICS OF THE TUMI DRILL PROGRAM
Phase
Phase 1
Phase 1
Phase 2
Phase 3
Total
6.8
Type
RC
DD
RC
RC
Holes
23
7
14
22
65
Metres
1,688
253
1,411
1,605
4,957
Samples
450
CRMs
37
Duplicates
42
216
404
1,070
20
43
100
14
99
155
2005 TO 2007 - BANDERA GOLD LTD.
On December 1, 2005, Bandera Gold Ltd. (“Bandera”) announced it had signed an Option
Agreement with MSJ to acquire 60% interest in the Cinco Minas and Gran Cabrera properties by
making option payments of $300,000, issuing 2,800,000 common shares of Bandera to MSJ, and
providing financing of $7,600,000 to MSJ over a five-year period for the exploration and
development of the properties.
Munroe (2006) provides an excellent historical review of the Cinco Minas Property from reports
and maps provided by MSJ and the work carried out by Bandera on the properties.
On September 15, 2007, Bandera announced considerable advances in the exploration, mining and
processing work on the Cinco Minas Property in the joint venture with MSJ. New roads were built
to historical waste rock dumps at the Las Amarillas and Magdalena historical workings to provide
material to process through the new 60 tpd “pilot” process plant. Exploration work continued
along the Cinco Minas Vein to the north of the CMMC mine and to the south near Cerro Colorado.
Williams (2007) reports MSJ initiated an underground mining program on the Destajos Horizon
of the El Abra workings and intersected the southeastern extremes of the historical workings. Work
involved clearing and supporting the workings where required and enlarging them for new
equipment (Figure 6.4).
Williams (2007) states that the pilot process plant is operational and states that adjusting “grind
parameters to achieve 80% passing -200 mesh will enable us to reduce the retention time in the
cyanide circuit as well as achieve higher recoveries. This could be achieved from a low-cost Cerro
Colorado operation with an added grade sweetener from El Abra.”
Williams (2007) reports the first “metal” was poured on September 7, 2007, and the second pour
was on September 22, 2007.
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FIGURE 6.4
PHOTOGRAPH OF THE EL ABRA WORKINGS
A.) LOOKING NORTHWEST ALONG
B.) LOOKING SOUTHWEST ALONG
EL ABRA WORKINGS
EL ABRA WORKINGS
Source: GoGold (2019)
6.9
2008 TO 2016 - LITIGATION BETWEEN BANDERA AND MSJ
Williams (2008) reported on January 4, 2008 that Bandera received notification from MSJ that the
Option Agreement had been terminated. Williams (2008) reported on January 7, 2008 that Bandera
“has contributed approximately $7M in project funding and issued share capital to secure title in
the Properties and has fulfilled its obligation to earn a 60% interest in the Properties”.
Bandera’s Form 51-102F1 MD&A for the three months ended February 28, 2015, reported:
“On February 26, 2008, the Company commenced legal action in Mexico with
respect to its interests in Cinco Minas and Gran Cabrera (the “Assets”).
Compensation being claimed by the Company includes enforcement of the Option
Agreement and damages arising from non-compliance by MSJ. A court of law in
Guadalajara, Mexico has awarded, as a preventative measure in favor of the
Company, encumbrances which have been filed against the applicable assets and
mining concessions with the Mexico Mining Public Registry (“MPR”). The
outcome of the claims for remedies and damages is not determinable; therefore, no
amounts have been recorded in the consolidated financial statements.”
On March 7, 2008, the State Court, Commercial Division of Guadalajara, Mexico, issued
preventive measures in favor of Bandera consisting of: (i) the encumbrance of assets of the
defendants for an amount of US$6M; (ii) the registration of the lawsuit on the files of each of the
mining concessions subject to the Option Agreement before the MPR in Mexico City; and (iii) a
prohibition for defendants, the legal representative of MSJ, to leave the Court’s jurisdiction (the
Mexican State of Jalisco) until this case is settled, unless having appointed an attorney to act on
his behalf while he is away. In order for these preventive measures to be put in place and remain
applicable, Bandera was required to deposit a refundable warranty bond of $502,319 (6M Mexican
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Pesos), in order to respond to any damages and injury that the defendants may suffer as a result of
the said preventive measures being emplaced.
On April 18, 2016, Bandera announced it terminated the litigation on its Cinco Minas and
Gran Cabrera mining properties and stated, “Bandera has diligently defended its rights to enforce
the original option agreement signed by the parties, only to be obstructed at every turn by
procedural maneuvers common in the justice system in Mexico”. Bandera recorded an Impairment
of the Cinco Minas and Gran Cabrera properties of $4,815,055 on April 15, 2018. During the
period from 2008 from 2016, MSJ was involved in additional litigation with employees.
The Los Ricos South Property has been well known to have strong geological potential. However,
the development of the asset was delayed by litigation. In 2016, the Mexican courts awarded title
of the Property to a private Mexican owner from which GoGold purchased the Property in 2019.
Title to the Property was subsequently transferred to GoGold.
6.10
PREVIOUS MINERAL RESOURCE ESTIMATE
The previous Mineral Resource Estimate by P&E (2023) for the Los Ricos South Property
included Measured and Indicated Mineral Resources of 98.6 Moz AgEq grading 276 g/t AgEq
contained in 11.1 Mt, and Inferred Mineral Resources of 13.6 Moz AgEq grading 185 g/t AgEq
contained in 2.3 Mt (Table 6.6). This Mineral Resource was considered to be amenable to
conventional open pit and underground longhole mining methods. The previous Mineral Resource
Estimate supported the 2023 Preliminary Economic Assessment of Los Ricos South Project (P&E,
2023). The effective date of the previous Mineral Resource Estimate was September 12, 2023.
The previous Mineral Resource Estimate is superseded by the current Mineral Resource Estimate
reported in Section 14 of this Technical Report.
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TABLE 6.6
2023 MINERAL RESOURCE ESTIMATE – PIT-CONSTRAINED AND OUT-OF-PIT (1-9)
Mining
Method
PitConstrained5
Pit - Cerro C.6
Out-of-Pit7,8
Eagle
Out-of-Pit7,8
Abra Main
Total
Classification
Tonnes
(M)
Measured
Indicated
Meas & Ind
Inferred
Inferred
Measured
Indicated
Meas & Ind
Inferred
Measured
Indicated
Meas & Ind
Inferred
Measured
Indicated
Meas & Ind
Inferred
3.9
2.8
6.7
0.5
0.9
0.7
1.2
1.9
0.1
1.1
1.4
2.5
0.8
5.7
5.4
11.1
2.3
Au
(g/t)
1.08
0.68
0.91
0.58
0.72
3.60
3.13
3.30
3.63
1.22
1.58
1.42
1.42
1.42
1.45
1.43
1.02
Average Grade
Ag
Cu
AuEq AgEq
(g/t)
(%)
(g/t)
(g/t)
142
0.03
2.94
231
89
0.03
1.87
146
120
0.03
2.49
195
99
0.04
1.91
150
31
0.01
1.12
88
298
0.35
7.94
621
164
0.37
5.79
453
214
0.36
6.59
516
122
0.54
6.00
470
194
0.06
3.79
297
178
0.21
4.18
327
185
0.15
4.00
313
133
0.41
3.73
292
172
0.07
3.72
291
129
0.15
3.33
260
151
0.11
3.53
276
85
0.17
2.36
185
Au
(koz)
135.9
60.7
196.6
9.6
20.9
80.7
117.5
198.2
7.8
44.7
71.5
116.2
36.8
261.4
249.7
511.0
75.0
Contained Metal
Ag
Cu
AuEq
(koz)
(Mlb)
(koz)
17,858
2.3
369.1
8,022
1.9
167.3
25,880
4.2
536.4
1,632
0.4
31.4
905
0.2
32.8
6,679
5.4
178.1
6,176
9.5
217.5
12,855
15.0
395.6
261
0.8
12.9
7,093
1.6
138.8
8,013
6.6
188.4
15,106
8.1
327.2
3,431
7.2
96.6
31,631
9.3
686.0
22,210
18.0
573.2
53,841
27.3
1,259.2
6,230
8.6
173.7
AgEq
(koz)
28,898
13,097
41,995
2,460
2,568
13,940
17,028
30,969
1,006
10,865
14,753
25,618
7,566
53,703
44,878
98,582
13,601
Notes: Meas = Measured, Ind = Indicated.
1.
Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources may
be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.
2.
The Inferred Mineral Resource in this estimate has a lower level of confidence than that applied to an Indicated Mineral Resource and must
not be converted to a Mineral Reserve. It is reasonably expected that the majority of the Inferred Mineral Resource could be upgraded to an
Indicated Mineral Resource with continued exploration.
3.
The Mineral Resources were estimated in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM
Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions
and adopted by the CIM Council.
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4.
5.
6.
7.
8.
9.
Historically mined areas were depleted from the Mineral Resource model.
The pit-constrained AgEq cut-off grade of 38 g/t AgEq was derived from US$1,800/oz Au price, US$23.00/oz Ag price, 85% Ag and 95% Au
process recovery, US$25/tonne process and G&A cost. The constraining pit optimization parameters were $2.10/t mineralized material and
waste mining cost, and 45-degree pit slopes.
Cerro Colorado Resource constrained to open pit mining methods only; out-of-pit Mineral Resources are restricted to the Eagle and Abra
mineralized veins, which exhibit historical continuity and reasonable potential for extraction by cut and fill and longhole mining methods.
The out-of-pit AgEq cut-off grade of 130 g/t AgEq was derived from US$1,800/oz Au price, US$23.00/oz Ag price, 85% Ag and 95% Au
process recovery, US$33/tonne process and G&A cost, and a $50/tonne mining cost. The out-of-pit Mineral Resource grade blocks were
quantified above the 130 g/t AgEq cut-off, below the constraining pit shell and within the constraining mineralized wireframes. Out–of-Pit
Mineral Resources are restricted to the Eagle and Abra Veins, which exhibit historical continuity and reasonable potential for extraction by
cut-and-fill and longhole mining.
AgEq and AuEq were calculated at an Ag/Au ratio of 78.2:1 for pit constrained and out-of-pit Resources.
Totals may not sum due to rounding.
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7.0
GEOLOGICAL SETTING AND MINERALIZATION
7.1
REGIONAL GEOLOGY
The Hostotipaquillo – Cinco Minas mining districts (Figure 7.1) occur at the intersection of two
extensive calc-alkaline magmatic arcs: 1) the older Sierra Madre Occidental (“SMO”) volcanic
province; and 2) the younger Mexican Volcanic Belt (“MVB”), also known as the Trans-Mexico
Volcanic Arc (“TMVA”) (Figure 7.2). The SMO volcanic province trends northwest along the
Pacific margin of Mexico and parallels the western coastline. It extends for approximately
1,700 km from the USA border to the Mexican state of Guerrero. The MVB covers the boundary
between the SMO and Cretaceous to Paleocene batholith and volcano-sedimentary sequences of
the Jalisco Block (“JB”) (Ferrari et al., 1999).
Two major volcanic sequences occur within the SMO volcanic province (Ferrari et al., 1999;
Garcia, 2019). The older volcanic sequence ranges in age from 100 Ma to 42 Ma (late Cretaceous
to Eocene), is 1.0 km to 1.5 km in thickness, and consists primarily of andesites and minor
rhyolites. The younger sequence, referred to as the upper volcanic series, overlies the older
andesite series. The age of the younger sequence is predominantly 37 Ma to 32 Ma, with the latest
volcanism occurring at approximately 18 Ma. The younger sequence is dominated by rhyodacite
to rhyolite ignimbrites with intercalated mafic lavas, suggesting bimodal volcanism. The
volcanism in the western SMO represents the largest known concentration of pyroclastic flows
and ignimbrites in the world. The SMO is related to the subduction of the Farallon plate. The
TMVA is reportedly attributed to the subduction of the Rivera and Cocos Plates, which includes
the Jalisco Block.
A volcanic plateau deformed by a series of horsts and grabens, forming prominent mesas and
canyons, occurs at the area of intersection of the south end of the SMO (Western Sierra Madre
physiographic province) and the MVB (Figure 7.2; Ferrari et al., 1999; Garcia, 2019). The
dominant major structural features in this area are the north-south oriented Bolaños and Colima
Grabens, which are separated by the west-northwest trending, apparently left-laterally displaced
Zacoalco Graben. The Hostotipaquillo-Cinco Minas District is located approximately at the
intersection of the Bolaños and Zacoalco Grabens and is located along, and bisected by, the
boundary of the SMO Block to the north and the Jalisco Block to the south.
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FIGURE 7.1
GEOLOGICAL MAP OF MEXICO*
* The Cinco Minas is the historical mine on the Los Ricos South Property.
Source: Modified by P&E (2024) from GoGold (2019)
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FIGURE 7.2
THE MAIN CENOZOIC VOLCANIC PROVINCES OF MEXICO
Note: MVB = Mexican Volcanic Belt
Source: Modified by P&E (2024) from Ferrari et al (1999)
7.2
LOCAL GEOLOGY
The geology of the Hostotipaquillo District is characterized by late Oligocene to Pliocene volcanic
and sub-volcanic intrusive rocks deformed by a set of northwest and east-west trending,
graben-forming normal faults (Figure 7.3; Garcia, 2019). Oligocene and Miocene volcanics are
primarily andesite flows, rhyolite ash flow and air fall tuffs, and rhyolite and dacite flow-domes
that are partially covered by Pliocene to Recent basalt flows. The northwest-trending graben that
extends across most of the district is one of several late Miocene to Quaternary tectonic depressions
formed in the area of the intersection of the south SMO and the TMVA, and is part of the larger
regional west-northwest trending Zacoalco Graben system. The Property geology is shown in
Figure 7.4.
The Rio Santiago River flows northwest through the district along the northeast margin of the
Hostotipaquillo district graben structure, including, from northwest to southeast, the
La Trini-Mololoa-Monte del Favor group of mines, Gran Cabrera group of mines,
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Santo Domingo-La Española Mine group, and Cinco Minas-El Aguila Mines vein systems.
These faults form prominent scarps that are the canyon walls on the southwest and south side of
Rio Santiago River. The mineralized vein systems in these faults form dip slopes in the river
canyon walls at several locations, such as Cabrera and Santo Domingo-La Española.
Andesite occurs in various colors and textures. Northwest of the El Aguila Mine, near the bottom
of the vein and in San Miguel Creek and the mouth of La Calera Creek, the andesite is greenish
grey in color and very fine-grained. It contains abundant quartz phenocrysts and previous
investigators classified the rock as quartz andesite.
At the Village of Cinco Minas, outcrops of the andesite form the hanging wall to the Cinco Minas
Vein. They are of reddish-purple, porphyritic andesite and overlie andesite tuffs
(Rivera and Vazquez). These andesites have been observed in an open pit exposure at the
El Abra workings (P&E, 2019). These volcanics have been down-dropped along the dip-slope of
the Cinco Minas Vein by a large, post-mineralization fault, such that they appear to conformably
overlie the vein/fault surface.
In the El Abra Mine, rhyolites occur overlying the andesite in the lower parts of the vein.
On surface, the rhyolites are found principally in outcrops above the vein and underlie most of the
higher hills found to the northeast of it. The rhyolites are pink and light green in colour and contain
quartz phenocrysts. Quartz phenocrystic rhyolite was observed along crosscuts connected to the
La Famosa level haulage, located in the footwall of Cinco Minas Vein (P&E, 2019).
Andesite and rhyolite tuffs are present. The andesite tuffs outcrop in Cinco Minas Creek
are light green in colour, fine-grained, and locally have purplish inter-bands and show indications
of internal folding. The latter type outcrops in the higher parts of the hillside in the extreme
northwest part of Cinco Minas Vein near the San Juan historical workings. Here they are pale pink
in colour and contain abundant quartz and biotite phenocrysts and feldspars phenocrysts that are
kaolinized.
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FIGURE 7.3
MAP SHOWING THE STRUCTURES AND VOLCANIC CENTRES IN THE
HOSTOTIPIQUILLO DISTRICT, MEXICO
Note: The dashed line marks the boundary between the SMO and the Jalisco Block.
Los Ricos is located approximately 10 km east of the Town of Hostotipaquillo.
Source: Modified by P&E (2024) from Garcia (2019)
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FIGURE 7.4
PROPERTY GEOLOGY
Source: GoGold (2020)
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Breccias occur above the rhyolite northeast of El Pitayo and form stratiform layers that strike
northwesterly and dip 32° to the northeast. They consist of angular fragments of red and green
volcanics 1 mm to 5 mm across. The matrix is rhyolite in composition. The orientation and
distribution of fragments suggests a vent source to the west.
Younger basalt overlay all the units noted above. Two groups of basalt are distinguished, one of
which occurs below the Cinco Minas Vein. They overlay the rhyolite northeast of El Capizayo and
are fine-grained. Petrographic analysis indicates the basalt to be hornblende and enstatite
porphyritic. Their stratigraphic position suggests that deposition early in the volcanic succession,
possibly part of a bimodal suite.
7.3
STRUCTURAL GEOLOGY
The structural geology of Los Ricos and the Monte del Favor Prospects, located 30 km northwest
of the Cinco Minas area (Figure 7.5), have been studied by Garcia (2019). Accordingly, the La
Trini – Monte Del Favor Prospects are genetically associated with the Monte del Favor Fault,
which trends N 135°. The structural relationships associated with this fault zone comply with the
geometric relationships of the system proposed by Riedel (1929), and indicate left-lateral sense of
movement.
On the other hand, the Los Ricos South Project is spatially associated with the Cinco Minas Fault
system, located approximately 11 km northeast from the Monte del Favor Fault. Kinematic
indicators in the Cinco Minas area indicate right-lateral movement, which implies that the block
contained between the two faults had a translational movement towards the northwest.
In the Los Ricos South area, a detailed study collected 863 measurements of fractures at
85 stations along the mineralized structure, which is approximately 3 km long with a general
northwest-southeast trend. Analysis of all the fracture measurements indicated three phases of
superimposed deformation:
1) Phase 1. Pre-mineralization deformation, characterized by right-lateral
movement oriented northwest-southeast, in which local areas underwent brittle
deformation to create the permeability that channelized flow of the mineralizing
solutions;
2) Phase 2. Deformation characterized by normal failure and extension subparallel to
Phase 1. Quartz textures and cross-cutting structural relationships indicate that
Phase 2 deformation began after the start of mineralization, reactivated the
right-lateral main fault and associated R, T, R', X and P Faults, and terminated prior to
the end of mineralization; and
3) Phase 3. Weak post-mineralization deformation with an oblique or orthogonal
orientation to Phases 1 and 2, characterized by right-lateral and left-lateral senses of
movement. This phase of deformation dislocated mineralized structure.
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FIGURE 7.5
STRUCTURAL GEOLOGICAL SETTING OF THE LOS RICOS SOUTH PROJECT AREA
Los Ricos South
Project
Source: GoGold (2019) and P&E (2020)
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7.4
MINERALIZATION
The mineralized vein at Los Ricos South, which is as much as approximately 30 m wide, has had
at least three quartz vein formation and metal precipitation events (Nebocat, 2004). The vein strikes
across the Property for 3.5 km, following a northwest-southeast trending fault, and dips
approximately 65° to 70° to the west.
In 1923, the Cinco Minas Company retained an independent consultant named George Garrey to
carry out and deliver a report on his geological examinations of the underground workings (Garrey,
1923). Garrey (1923) examined the underground workings and reports the chief gangue mineral is
quartz, whereas crystalline quartz is plentiful and colourless, cream colored, milky or locally even
amethyst, the bulk of the quartz is a dense fine-grained grey, yellowish of greenish porcelain quartz
intermixed with chalcedonic quartz of a slightly later origin. This quartz is associated with minor
grey, brown and white calcite. Some of the calcite was deposited contemporaneously with the
quartz, however, most of it post-dates the latter. The contemporaneous calcite is more abundant in
the vein near surface than at depth. Comb quartz and open vugs in the quartz vein were more
abundant in the upper workings near surface. Siderite, rhodochrosite and adularia were also
observed as rare occurrences, whereas kaolinite was also present near secondary enriched oxidized
mineralized materials. Chlorite and possibly epidote were also in evident in association with the
quartz, or in partially replaced rock fragments.
The chief sulphide minerals are galena, chalcopyrite, sphalerite, pyrite, covellite, bornite and silver
sulphides (Figure 7.6). The higher-grade sulphide zones are generally associated with fine-grained
galena and minute specks of chalcopyrite. Coarse-grained sulphides generally carry low metal
values. Recognizable silver sulphides are rare. However, specks of pyrargyrite were observed in
the high-grade sulphides that might have been tetrahedrite, argentite or other silver sulphide
minerals. Native silver flakes and specks were observed coating fractures in high-grade sulphide
mineralized zones.
There appear to have been either several periods of vein deposition or a long period of
mineralization interrupted periodically by faulting and brecciation of the quartz. The
mineralization appears to have formed relatively early. The location of the mineralized shoots,
which appear to rake at 65° to 70° northwestwards, appears to have been controlled by:
•
segregation of the minerals during deposition from the original primary solutions;
•
periods of movement and brecciation that followed the first period of mineralization;
and
•
formation of the diagonal slips or cross-faults that confined or diverted the later primary
mineralizing solutions into these breccia-filled channels.
The portions of the quartz vein between the various mineralized shoots show little evidence of
brecciation of the original quartz. The richest mineralized shoots appear to have been associated
with the greatest width of quartz and the portions of the veins showing the greatest number of
periods of brecciation.
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FIGURE 7.6
PHOTOGRAPH OF DRILL CORE SHOWING SULPHIDE MINERALS
Source: GoGold (2019)
The Magdalena and the San Pedro-San Juan mineralized shoots, although of minor importance,
appear to have a true shoot-like character and plunge approximately 65° to the northwest. On the
Destajos Level and above, there was originally almost continuous mineralization from the
southeast edge of the San Nicolas Stope to the northwest edge of the Main Destajos Stopes at
approximately 320 m northwest of the main shaft.
Garrey (1923) reports that the San Nicolas, San Diego, Destajos, Cinco Minas and North Cinco
Minas Deposits were probably part of what appears to have been one large, mineralized shoot,
subsequently broken-up along faults into smaller shoots. Drill hole LRGG-19-079 intersected the
Los Ricos Vein from 128.2 m to 144.7 m (Figure 7.7), which displays typical gold and silverbearing quartz veins in the area of the historical Cinco Minas underground mine.
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FIGURE 7.7
PHOTOGRAPHS OF THE
HOLE LRGG-19-079
LOS
RICOS
VEIN
IN
DRILL
CORE,
Source: GoGold (2019)
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8.0
DEPOSIT TYPES
Los Ricos (Cinco Minas) is a classic Neogene (previously Tertiary) age, volcanic-hosted and
low-sulphidation epithermal precious metal deposit.
Epithermal deposits of Au (±Ag) are a type of lode gold deposit that consists of mineralized veins
and disseminations formed at or near (≤1.5 km) the Earth’s surface (Hedenquist, 2000; Taylor,
2007; Figure 8.1). The deposits occur in association with hot springs and young volcanic centres.
The host rocks are volcanic and volcaniclastic sedimentary, sedimentary and metamorphic rocks.
The mineralization is dominated by Au and Ag; however, Cu, Pb and Zn may also be present in
variable amounts.
Epithermal Au deposits are classified on the basis of the sulphidation state of the sulphide
mineralogy as belonging to one of three sub-types (Hedenquist, 2000; Taylor, 2007):
•
Low-Sulphidation: previously called adularia-sericite, these low-sulphidation subtype
deposits are considered to be formed by near-neutral pH fluids, as a result of being
dominated by meteoric waters. However, some deposits contain magma-derived C and
S.
•
High-Sulphidation: previously called quartz-(kaolinite)-alunite, alunite-kaolinite,
enargite-Au, or high-sulphur deposits, these highly acidic deposits generally occur
proximal to magmatic sources of heat and volatiles and form from acidic hydrothermal
fluids containing magmatic S, C and Cl.
•
Intermediate-Sulphidation: some deposits with mostly low-sulphidation
characteristics have sulphide mineralized assemblages that represent a sulphidation
state between that of high-sulphidation and low-sulphidation deposits.
In high-sulphidation systems, native Au and electrum are typically associated with pyrite, enargite,
covellite, bornite and chalcocite. In addition to sulphosalts and base metal sulphides, tellurides and
bismuthinite are present in some deposits. Total sulphide contents are generally higher in highsulphidation than low-sulphidation subtype deposits. However, high sulphide contents may also
characterize transitional polymetallic low-sulphidation deposits. Where base metals are present in
high-sulphidation deposits, the Cu abundance can vary significantly and typically dominate that
of Zn. Principal gangue minerals are quartz (“vuggy silica”), alunite and barite (particularly
associated with Au). High sulphidation deposits form proximal to shallowly emplaced magmatic
intrusions.
In low-sulphidation systems, native Au and electrum occur in vein deposits that contain only up
to a few percent sulphides. The principal gangue minerals are calcite, chlorite, adularia, barite,
rhodochrosite, fluorite and sericite. Calcite is characteristic of low-sulphidation deposits, because
of the low acidity of the hydrothermal mineralizing fluids. Low-sulphidation deposits form in fault
zones, sedimentary rocks and vein complexes relatively more distal from magmatic intrusions
(Taylor, 2007).
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FIGURE 8.1
HYDROTHERMAL FLUID SYSTEM MODEL SHOWING EPITHERMAL AND
PORPHYRY RELATED MINERAL DEPOSIT TYPES
low
Source: Sillitoe (2010)
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9.0
EXPLORATION
GoGold’s 2019 exploration program consisted of mapping, sampling and diamond drilling along
the Cinco Minas Vein system from the San Juan prospect located north of the historical
Cinco Minas workings to south of the Cerro Colorado Deposit, an overall distance of 3.5 km.
The objective of the program was to identify and test targets with potential to host near-surface
precious metal mineralization. The program consisted of:
9.1
•
Digital compilation studies of historical records and maps obtained from the Gerard
Archives;
•
Acquisition of new 1 m resolution topographic information across the Property;
•
Geological mapping and prospecting;
•
Trenching and sampling; and
•
Diamond drilling at the Cinco Minas and Cerro Colorado Deposits.
DIGITAL TERRAIN MODEL
GoGold commissioned PhotoSat of Vancouver, BC to produce a DTM for the Los Ricos South
Property with 1 m contours (Figure 9.1). The coordinate system is WGS84 UTM Zone 13 and
elevations are heights above the EGM2008 geoid. All historical records and current exploration
and drilling work were recorded using this DTM, resulting in consistent, highly accurate base maps
and datasets for the Property.
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FIGURE 9.1
AREA COVERED WITH THE 2019 1 M DIGITAL TERRAIN MODEL BY PHOTOSAT
Source: GoGold (2020)
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9.2
GEOLOGICAL MAPPING AND PROSPECTING
Geological mapping surveys were initiated in the spring of 2019 with the goal of locating and
following the Cinco Minas Vein and alteration zone along strike to the north of the historical mine
workings and to the south end of the Property. The work was completed at 1:1,000 scale.
Outcrop exposures are abundant on the Property and many historical pits, shafts, adits and waste
dumps were located. All the information was transferred to an ArcGIS database.
Figure 9.2 shows the local area geology, four zones of the Los Ricos South Vein systems, and the
gold equivalent (“AuEq”) assay sample results from the sampling program and geological
mapping survey. Figures 9.3 to 9.5 show the zones in greater detail.
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FIGURE 9.2
LOCAL GEOLOGY MAP
Source: GoGold (2020)
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FIGURE 9.3
NORTHERN GEOLOGY MAP EL AGULILA AREA
Source: GoGold (2020)
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FIGURE 9.4
CENTRAL GEOLOGY MAP EL ABRA GEOLOGY
Source: GoGold (2020)
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FIGURE 9.5
SOUTHERN GEOLOGY MAP CERRO COLORADO AREA
Source: GoGold (2020)
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9.3
TRENCHING AND CHANNEL SAMPLING
In order to evaluate the Los Ricos South Vein in the areas between the historical workings,
a trenching program was carried out. Trenches were excavated across the vein and alteration zone
exposures at intervals of 20 m along the strike of the vein. A tractor mounted backhoe was used
where access allowed; otherwise, the trenches were dug by hand. All sample location information
was recorded using the GPS and all the sample information was transferred to the ArcGIS database.
See Figure 9.6 for a general trench location map in the Los Lamas area and Figures 9.7 and 9.8 for
Trenches 10 and 11 at Los Lamas. Figures 9.9 to 9.12 are maps showing detailed geology with
trench and channel sample locations and AuEq assay values for the trench channel samples.
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FIGURE 9.6
TRENCH LOCATION MAP OF THE LOS LAMAS AREA
Source: GoGold (2020)
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FIGURE 9.7
TRENCH NUMBER 10,
LOS LAMAS
FIGURE 9.8
TRENCH NUMBER 11,
LOS LAMAS
Los Lamas Trench number 10 was advanced
by pick and shovel.
Los Lamas Trench number 11 was advanced
by pick and shovel.
Source: GoGold (2019)
Source: GoGold (2019)
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FIGURE 9.9
NORTHERN TRENCH MAP EL AGUILA AREA
Source: GoGold (2020)
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FIGURE 9.10
CENTRAL TRENCH MAP EL ABRA
Source: GoGold (2020)
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FIGURE 9.11
SOUTHERN TRENCH MAP LAS LAMAS
Source: GoGold (2020)
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FIGURE 9.12
SOUTHERN TRENCH MAP CERRO COLORADO
Source: GoGold (2020)
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9.4
UNDERGROUND MAPPING AND SAMPLING
The northern extension of the San Juan Vein is known as the Rascadero Vein and is exposed in
the El Troce workings. The El Troce workings were surveyed, mapped and sampled as shown in
Figures 9.13 and 9.14 and the results summarized in Table 9.1.
FIGURE 9.13
LOCATION OF EL TROCE WORKINGS
Source: GoGold (2020)
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FIGURE 9.14
SAMPLING EL TROCE WORKINGS
Source: GoGold (2019)
TABLE 9.1
ASSAY RESULTS EL TROCE SAMPLES
Location
Channel ID
El Troce UG
El Troce UG
El Troce UG
El Troce UG
El Troce UG
El Troce UG
El Troce UG
El Troce UG
El Troce UG
El Troce UG
LRTRH-0351
LRTRH-0352
LRTRH-0353
LRTRH-0354
LRTRH-0355
LRTRH-0356
LRTRH-0357
LRTRH-0358
LRTRH-0359
LRTRH-0360
From
(m)
0.0
0.0
0.0
0.0
0.0
0.0
0.0
0.0
0.0
0.0
To
(m)
2.0
3.0
0.6
2.2
2.1
2.5
2.1
2.4
2.3
2.1
Length
(m)
2.0
3.0
0.6
2.2
2.1
2.5
2.1
2.4
2.3
2.1
Gold
(g/t)
0.98
2.34
2.43
0.58
0.57
2.12
2.18
1.00
6.80
1.55
Silver
(g/t)
206.7
430.9
492.2
154.6
85.8
128.4
108.5
44.3
149.9
179.9
AuEq75
(g/t)
3.73
8.09
8.99
2.64
1.71
3.83
3.62
1.59
8.80
3.94
Note: AuEq75 = gold equivalent at 75/1 Ag/Au ratio @ 100% processing recovery.
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9.5
EAGLE CONCESSION EXPLORATION
Following acquisition of the Eagle Concession in October, 2022, GoGold completed preliminary
geological mapping, sampling, geophysical IP surveying, and commenced a drill program.
Mapping and geophysics indicate that the mineralized Eagle structure extends >2.5 km to the
northwest. Significant geophysical anomalies were identified that formed targets for subsequent
drill testing (Figure 9.15). The 2022 drill program results are described in Section 10 of this
Technical Report.
FIGURE 9.15
EAGLE AND MAIN AREA IP GEOPHYSICAL RESPONSE
Source: GoGold (Press Release dated October 18, 2022)
9.6
EXPLORATION POTENTIAL
In addition to the exploration work completed, the Authors established that the Los Ricos South
Abra and Eagle mineral deposits contain an additional Exploration Target of 1.8 to 2.2 Mt at 375
to 425 g/t AgEq for 22 to 30 Moz AgEq. The Exploration Target is based on the estimated strike
length, depth and thickness of the known mineralization. This Exploration Target is shown in
Figure 9.16.
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FIGURE 9.16
LONGITUDINAL PROJECTION OF EXPLORATION TARGET
Source: P&E (2023)
The potential quantities and grades of the Exploration Target are conceptual in nature. There
has been insufficient work done by a Qualified Person to define this estimate as a Mineral
Resource. The Company is not treating this estimate as Mineral Resources, and readers should
not place undue reliance on this estimate. Even with additional work, there is no certainty that
this estimate will be classified as Mineral Resources or that it will ever prove to be economically
recoverable.
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10.0
DRILLING
As of the effective date of this Technical Report, GoGold has completed 524 drill holes totalling
89,380 m at Los Ricos South. Due to the extensive amount of drilling, the collar locations have
been listed in Appendix F and significant intersection tables have been listed in Appendix G.
Highlights of the drill hole program at Los Ricos South are summarized below, based in part on
information available on GoGold’s website and press releases. A summary of the drill programs
at the Los Ricos South Property is presented in Table 10.1.
TABLE 10.1
SUMMARY OF DRILL PROGRAMS ON THE LOS RICOS SOUTH PROPERTY
Year
Company
2003
Tumi Resources
2004
Tumi Resources
2019
GoGold
2020
GoGold
2021
GoGold
2022
GoGold
2023
GoGold
Subtotal Tumi
Subtotal GoGold
Total
10.1
Number of Holes
Drilled
42
22
112
123
12
231
46
64
524
588
Metres Drilled
(m)
3,243
1,571
16,918
19,654
2,473
42,776
7,559
4,814
89,380
94,194
EAGLE CONCESSION DRILLING
GoGold announced the acquisition of the Eagle Concession in October 2022, which covers the
northern strike extension of the Main (Abra) Deposit on the Los Ricos South Property and
represents an extension to the 2021 Mineral Resource Estimate at Los Ricos South.
During negotiations for the concession acquisition and securing legal ownership, GoGold
completed due diligence drilling, including drill hole LRGAG-21-014, which intercepted 20,269
g/t silver equivalent (“AgEq”) over 0.7 m, contained within 11.9 m of 2,260 g/t AgEq which was
contained within a wider intersection of 72.6 m of 461 g/t AgEq. This represents the highest values
drilled in the district, with drill hole LRGAG-21-031 having comparable results of 428 g/t AgEq
over 68 m. The mineralized intercepts were consistent with geophysical targets on the new
concession.
Figure 10.1 shows a longitudinal projection of the Eagle Concession and Main Area. Cross-section
projections are presented in Figure 10.2 through Figure 10.5. Select assay intersections in the
drilling are listed below:
•
•
LRGAG-21-014: 2,260 g/t AgEq over 11.9 m within 72.6 m of 461 g/t AgEq.
LRGAG-21-032: 4,895 g/t AgEq over 0.6 m within 65.8 m of 210 g/t AgEq.
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•
•
•
•
•
•
•
•
LRGAG-21-035: 21,580 g/t AgEq over 0.9 m within 110.6 m of 388 g/t AgEq.
LRGAG-22-043: 3,420 g/t AgEq over 0.7 m within 20.7 m of 315 g/t AgEq.
LRGAG-22-063: 1,565 g/t AgEq over 0.8 m within 15.7 m of 137 g/t AgEq.
LRGAG-22-118: 55.0 m of 2,738 g/t AgEq, including 7.0 m of 20,715 g/t AgEq. Also
includes 63,658 g/t AgEq over 2.0 m.
LRGAG-22-126: 5,818 g/t AgEq over 0.8 m within 23.3 m of 617 g/t AgEq.
LRGAG-22-162: 1,519 g/t AgEq over 31.0 m within 8.1 m of 5,375 g/t AgEq
including 0.7 m of 47,119 g/t AgEq.
LRGAG-22-164: 1,126 g/t AgEq over 50.0 m including 7.8 m of 6,334 g/t AgEq
including 0.8 m of 46,822 g/t AgEq.
LRGAG-22-165: 1,603 g/t AgEq over 3.2 m including 22.1 m of 273 g/t AgEq.
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FIGURE 10.1
EAGLE LONGITUDINAL PROJECTION
Source: GoGold (Press Release dated March 29, 2023)
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FIGURE 10.2
EAGLE CROSS-SECTION PROJECTION DRILL HOLES 11, 12, 13, 14
Source: GoGold (Press Release dated October 18, 2022)
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FIGURE 10.3
EAGLE CROSS-SECTION PROJECTION DRILL HOLES 30, 31
Source: GoGold (Press Release dated October 18, 2022)
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FIGURE 10.4
EAGLE CROSS-SECTION PROJECTION DRILL HOLES 32, 35
Source: GoGold (Press Release dated November 2, 2022)
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FIGURE 10.5
EAGLE CROSS-SECTION PROJECTION DRILL HOLE LRGAG-22-118
Source: GoGold (Press Release dated January 23, 2023)
10.2
MAIN (ABRA) DEPOSIT DRILLING
Drilling on the Main (Abra) Deposit was designed to better define the high-grade portions that
may be amenable to bulk underground mining. These drill holes are in addition to those completed
in 2019 and 2020 which were the basis for the 2021 Los Ricos South PEA.
105 holes totalling 17,321.9 m have been drilled in the Main Deposit area by GoGold since the
2021 PEA. Figure 10.6 shows a longitudinal projection of Main area and Figure 10.7 shows the
pierce points on the 2021 Mineral Resource Model. Cross-section projections are presented in
Figure 10.8 and Figure 10.9. Selected intersections of silver and gold mineralization in the drilling
are as follows:
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•
•
•
•
•
•
•
•
•
•
•
•
•
•
•
•
•
•
LRGG-22-209: 11,103 g/t AgEq over 1.0 m within 33.2 m of 513 g/t AgEq.
LRGG-22-218: 2,558 g/t AgEq over 4.8 m within 15.0 m of 828 g/t AgEq.
LRGG-22-235: 1,276 g/t AgEq over 3.5 m within 13.9 m of 417 g/t AgEq.
LRGAG-22-144: 2,015 g/t AgEq over 0.5 m within 14.2 m of 977 g/t AgEq.
LRGAG-22-145: 6,951 g/t AgEq over 2.4 m within 5.4 m of 1,056 g/t AgEq.
LRGG-22-246: 2,646 g/t AgEq over 1.0 m within 14 m of 302 g/t.
LRGG-22-280: 706 g/t AgEq over 14.0 m including 2.3 m of 4,081 g/t AgEq and 0.6
m of 9,283 g/t AgEq.
LRGG-22-254: 2,360 g/t AgEq over 0.6 m within 11.9 m of 279 g/t AgEq.
LRGG-22-269: 3,218 g/t AgEq over 0.9 m within 17.3 m of 350 g/t AgEq.
LRGF-23-281: 3,183 g/t AgEq over 1.3 m within 11.7 m of 457 g/t AgEq.
LRGF-23-286: 2,030 g/t AgEq over 1.7 m within 23.1 m of 203 g/t AgEq.
LRGF-23-291: 3,382 g/t AgEq over 1.0 m within 34.8 m of 253 g/t AgEq.
LRGF-23-302: 404 g/t AgEq over 30.2 m including 9.7 m of 1,110 g/t AgEq and 0.8
m of 5,468 g/t AgEq.
LRGG-23-306: 5,838 g/t AgEq over 1.5 m within 11.7 m of 979 g/t AgEq.
LRGG-23-309: 1,048 g/t AgEq over 0.6 m within 18.6 m of 202 g/t AgEq.
LRGG-23-316: 491 g/t AgEq over 27.3 m including 7.6 m of 1,670 g/t AgEq, and 0.9
m of 8,115 g/t AgEq.
LRGTO-22-002: 3.56 g/t AgEq over 5.2 m including 0.9 m of 484 g/t AgEq and 0.95
m of 203.2 g/t AgEq.
LRGTO-23-005: 4.23 g/t AgEq over 4.9 m including 0.9 m of 846.5 g/t AgEq and
0.65 m of 221 g/t AgEq.
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FIGURE 10.6
LONGITUDINAL PROJECTION – MAIN AREA
Source: GoGold (Press Release dated November 30, 2022)
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FIGURE 10.7
LONGITUDINAL PROJECTION – LOS RICOS SOUTH RESOURCE AREA WITH PIERCE POINTS
Source: GoGold (Press Release dated August 23, 2023)
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FIGURE 10.8
MAIN AREA CROSS-SECTION PROJECTION DRILL HOLE LRGG-22-209
Source: GoGold (Press Release dated November 16, 2022)
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FIGURE 10.9
MAIN AREA CROSS-SECTION PROJECTION DRILL HOLE LRGG-22-218
Source: GoGold (Press Release dated November 30, 2022)
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11.0
SAMPLE PREPARATION, ANALYSES AND SECURITY
The following section discusses sampling carried out by GoGold at the Los Ricos South Property
between 2019 and 2023.
11.1
SAMPLE PREPARATION
11.1.1
Channel Sampling
GoGold has carried out a general surface and underground sampling program on the Los Ricos
South Property. Sampling included chip channel samples and grab samples following a protocol
of sampling procedures including:
•
Channel sampling controls including keeping records of the sample type, size, number
and location using GPS;
•
The sample locations were photographed;
•
One in every 40 samples were duplicated and sent for analysis;
•
Every 40 samples one blank sample was inserted; and
•
Every 40 samples one control sample of commercial certified reference material
(“CRM”) was inserted.
Identical procedures were used for sampling in the historical mine workings. Samples were taken
by local crews under the supervision of a geologist from SPM. Chip samples were cut with a
hammer and chisel, collected on a tarp and placed in a plastic bag to be labelled and sent to the
laboratory for precious metal assay and ICP multi-element analysis.
11.1.2
Drill Core Sampling
The protocol for handling, sampling and assaying diamond drill core samples was developed in
2011 by David Duncan, P.Geo., for GoGold’s San Diego project in Durango State during 2012 –
2014 and GoGold’s Santa Gertrudis program in 2015 to 2017. These same protocols are used for
the Los Ricos South drilling program and are described as follows:
•
The drill core was placed in labelled drill core boxes by the drilling contractor with
footage blocks inserted in the trays at the end of each run. The lids were placed on and
subsequently fastened to the drill core boxes.
•
GoGold’s geologists and geo-technicians were present at the drill rig to ensure that drill
core handling, drill core accommodation, box number and depth recording was
properly done by the drilling contractor.
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•
The drill core was transferred from the drill rig to GoGold’s drill core logging, sampling
and storage facilities at Cinco Minas, where the trays were placed in order on the
logging tables and the first inspection is made prior to cleaning and washing the drill
core of any drilling muds.
•
All depth marker tags were checked for completeness and accuracy with special
attention paid to possible mining voids.
•
The SPM geo-technicians aligned the drill core pieces, assess and measure drill core
recoveries and RQD and photograph the drill core.
•
Bulk density measurements were reported for all drill holes by GoGold geotechnicians. The geo-technicians selected an intact cylinder of drill core 10 cm to 20
cm in length, record the weight, coat the sample in paraffin wax, and dry and record
the weight with wax, submerge the sample in a graduated cylinder filled with water,
record the change in volume in water, divide the weight with paraffin by the volume
displaced to determine the bulk density. Measurements for each box of drill core were
made from the top to the bottom of the drill hole, thus providing excellent
representative coverage through the hanging wall units, the Los Ricos quartz vein, and
into the footwall units.
•
The SPM geologists logged the drill core and laid out the areas to be sampled by the
geo- technicians.
•
Boxes of drill core were transferred to the sampling room, where the drill core was
sawn in half by a diamond saw.
•
The half drill core samples were placed in plastic bags along with a sample tag ID and
tied closed with zip locks under the supervision of the SPM geologists. Sample tags
had three portions; one for the drill core tray, the sample bag and one left in the sample
book.
•
Up to 10 sample bags were placed in larger rice bags, which were tied closed with zip
locks and labelled.
•
The remainder of the sample was returned to the drill core box, the lids replaced, and
the boxes were transferred to drill core racks at GoGold’s secure drill core storage
facility at Cinco Minas.
•
All samples were collected by SPM personnel and delivered to the Actlabs laboratory
in Zacatecas or ALS Chemex (“ALS”) in Guadalajara. The drill core and samples are
under GoGold’s or SPM’s supervision, from the time of pick-up of the drill core at the
drill site until they are delivered to laboratory staff. All drill core and sample splits are
kept in a secure storage facility at Cinco Minas. SPM use their own vehicles to transport
the samples to the Actlabs sample preparation facility in Zacatecas, or the ALS facility
in Guadalajara, for both sample preparation and analyses. The samples are generally
received by the lab within two days.
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•
Assay data was reported electronically from Actlabs and ALS to GoGold and SPM.
ActLabs protocol crushes samples to a nominal minus 10 mesh (1.7 mm), mechanically split
(riffle) to obtain a representative sample and then pulverized to at least 95% minus 150 mesh (106
microns). Samples were analyzed for gold and silver, as well as an array of other elements. Gold
analysis was carried out by fire assay with atomic absorption spectroscopy (“AAS”) finish.
Reporting limits for this test method were 0.005–10 ppm. Results exceeding 10 ppm Au were
reanalyzed using fire assay with a gravimetric finish and reported in g/t. Silver analysis was carried
out by total digestion with ICP finish. Reporting limits for this test method were 0.3–100 ppm.
Results exceeding 100 ppb Ag were reanalyzed using fire assay with a gravimetric finish, and
reported in g/t.
ALS crushed the samples and prepared 200 to 300 gram, pulp samples with 90% passing Tyler
150 mesh (106 μm). The pulps were assayed for gold using a 30-gram charge by fire assay (Code
AA23) and over limits greater than 10 g/t are re-assayed using a gravimetric finish (Code MEGRAV21). Silver and multi-element analysis was completed using total digestion (Code MEICP61 Total Digestion ICP). Over limits greater than 100 g/t silver were re-assayed using a
gravimetric finish (ME-GRA21).
The Actlabs’ Quality System is accredited to international quality standards through the ISO/IEC
17025:2017 and ISO 9001:2015. The accreditation program includes ongoing audits, which verify
the QA system and all applicable registered test methods. Actlabs is also accredited by Health
Canada.
ALS has developed and implemented strategically designed processes and a global quality
management system at each of its locations. The global quality program includes internal and
external inter-laboratory test programs and regularly scheduled internal audits that meet all
requirements of ISO/IEC 17025:2017 and ISO 9001:2015. All ALS geochemical hub laboratories
are accredited to ISO/IEC 17025:2017 for specific analytical procedures. Both ALS and Actlabs
are independent of GoGold and SPM.
It is the Author’s opinion that sample preparation, security and analytical procedures for the Los
Ricos South Project 2019 to 2023 drilling were adequate for the purposes of this Mineral Resource
Estimate.
11.2
BULK DENSITY
Bulk density measurements were determined by GoGold using the water immersion method on
diamond drill core samples. Samples were collected at 20 m intervals downhole and data from all
the major units in the hanging wall, mineralized zones and footwall were logged. As described in
section 14 of this Technical Report, the average bulk density was 2.53 tonnes per cubic metre
(t/m3), with a slightly higher bulk density observed for the higher-grade domains (Table 11.1). A
total of seven measurements were also taken for the Cerro Colorado Deposits, with the average
bulk density calculated at 2.49 t/m3, and the median bulk density at 2.50 t/m3.
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TABLE 11.1
SUMMARY OF BULK DENSITY STATISTICS
Item
Waste
Abra Main
Abra Plum North
Abra Plum South
Eagle Main
Eagle Plum FW
Eagle Plum HW
Total
Count
2,822
527
169
10
384
199
404
4,515
Avg Bulk
Density
(t/m3)
2.46
2.55
2.70
2.54
2.57
2.71
2.79
2.53
Median Bulk
Density
(t/m3)
2.50
2.54
2.67
2.55
2.57
2.65
2.69
2.54
Independent verification sampling carried out in August 2019 and May 2023 has confirmed
GoGold’s onsite measurements. A total of 22 due diligence samples were measured independently
at Actlabs (10 determinations by pycnometer on pulps in 2019) and ALS (12 determinations by
water displacement on drill core in 2023), returning a mean value of 2.64 t/m3, median value of
2.64 t/m3, minimum value of 2.49 t/m3 and a maximum value of 2.77 t/m3.
11.3
2019 QUALITY ASSURANCE/QUALITY CONTROL REVIEW
GoGold implemented and monitored a thorough quality assurance/quality control (“QA/QC” or
“QC”) program for the diamond drilling undertaken at the Los Ricos South Project over the 2019
period. QC protocol included the insertion of QC samples into every batch sent off for analysis,
including certified reference material (“CRMs”), blanks and field duplicates.
CRMs and blanks were inserted approximately every 1 in 60 samples. In addition, field duplicates
consisting of ¼ drill core were collected approximately every 70 samples.
11.3.1
Performance of Certified Reference Materials
CRMs were inserted into the sample stream approximately every 60 samples. Two CRMs were
used during the 2019 drill program to monitor for gold performance only; the AR09002X and
AR09003X CRMs. The former a low-grade gold CRM and the latter of higher grade.
Criteria for assessing CRM performance are based as follows. Data falling within ± 2 standard
deviations from the accepted mean value pass. Data falling outside ± 3 standard deviations from
the accepted mean value, or two consecutive data points falling between ± 2 and ± 3 standard
deviations on the same side of the mean, fail.
Results are presented in Table 11.2.
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TABLE 11.2
SUMMARY OF 2019 AU CERTIFIED REFERENCE MATERIALS USED AT LOS RICOS SOUTH
Certified
Reference
Material
AR09002X
AR09003X
Certified
Mean
Value
(ppm)
0.151
1.231
ActLabs Results
+/- 1SD
(ppm)
+/- 2SD
(ppm)
0.02
0.08
0.04
0.16
No.
Results
No. (-)
Failures
No. (+)
Failures
54
35
0
0
0
0
Average
Result
(ppb)
0.15
1.23
Source: P&E (2021)
The AR09002X gold CRM was manufactured by Shea Clark Smith / Minerals Exploration &
Environmental Geochemistry of Reno, Nevada for Animus Resources in 2009. The CRM was
prepared from mineralized rock derived from the Santa Gertrudis Mine. It is certified for Au. There
were 54 data points for this reference material and there were no recorded failures for this CRM.
The AR09003X gold CRM was also manufactured by Shea Clark Smith / Minerals Exploration &
Environmental Geochemistry of Reno, Nevada for Animus Resources in 2009. The CRM was
prepared from mineralized rock derived from the Santa Gertrudis Mine. It is certified for Au. There
were 35 data points for this CRM and no failures were returned for the AR09003X CRM.
The Author considers that the CRM data demonstrate acceptable accuracy in the 2019 data.
11.3.2
Performance of Blanks
All blank data for Au and Ag were graphed. If the assayed value in the certificate was indicated as
being less than detection limit the value was assigned the value of half the detection limit for data
treatment purposes. An upper tolerance limit of three times the detection limit was set. There was
a total of 89 data points to examine. All data plotted at or below the set tolerance limits, except for
a single Ag sample (LRC-002260 which returned a result of 0.9 ppm) that plots just above the set
tolerance limit, and the Author does not consider contamination to be an issue for the 2019 drill
data.
11.3.3
Performance of Field Duplicates
Field duplicate data were examined for the 2019 drill program for both gold and silver. There was
a total of 77 duplicate pairs in the data set. Data were scatter graphed (Figures 11.1 and 11.2) and
found to have acceptable precision for the field level for both gold and silver, with respective Rsquared values of 0.999 and 0.997.
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FIGURE 11.1
PERFORMANCE OF AU FIELD DUPLICATES FOR 2019 DRILLING AT LOS
RICOS SOUTH
Source: P&E (2021)
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FIGURE 11.2
PERFORMANCE OF AG FIELD DUPLICATES FOR 2019 DRILLING AT LOS
RICOS SOUTH
Source: P&E (2021)
11.4
2020 QUALITY ASSURANCE/QUALITY CONTROL REVIEW
GoGold continued to implement and monitor a thorough QC program for the diamond drilling
undertaken at the Los Ricos South Project throughout 2020. The rate of QC sample insertion was
increased from the 2019 program, however, again included the routine insertion of CRMs, blanks
and field duplicates into the sample stream.
CRMs were inserted approximately every 1 in 10 samples and blanks at a rate of 1 in 25. In
addition, field duplicates consisting of ¼ drill core were collected approximately every 25 samples.
11.4.1
Performance of Certified Reference Materials
CRMs were inserted into the sample stream approximately every 10 samples. Five CRMs were
used during the 2020 drill program; the AR09002X, AR09003X and OxB139 gold CRMs, as well
as the SN104 and SQ88 CRMs, which are certified for both gold and silver. Criteria for assessing
CRM performance remained as described in section 11.3.1 of this Technical Report.
A summary of the CRM performance results for the 2020 program is presented in Table 11.3.
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TABLE 11.3
SUMMARY OF 2020 CERTIFIED REFERENCE MATERIALS USED AT LOS RICOS SOUTH
Certified
Mean
Value
(ppm)
Monitoring Gold
AR09002X
0.151
AR09003X
1.231
OxB130
0.125
SN104
9.182
SQ88
39.723
Certified
Reference
Material
Monitoring Silver
SN104
46.7
SQ88
160.8
ActLabs Results
+/- 1SD
(ppm)
+/- 2SD
(ppm)
0.02
0.08
0.006
0.184
0.947
1.4
5.1
No.
Results
No. (-)
Failures
No. (+)
Failures
Average
Result
(ppb)
0.04
0.16
0.012
0.368
1.894
15
96
35
142
73
0
0
0
0
0
0
0
0
0
0
0.15
1.23
0.13
9.20
39.31
2.8
10.2
142
73
0
0
0
0
47.26
158.35
Source: P&E (2021)
There were 15 data points for Au for the AR09002X CRM and 96 data points for Au for the
AR09003X CRM. There were no recorded failures for either CRM.
The OxB130 gold CRM was supplied by Rocklabs Reference Materials of Aukland, New Zealand.
The CRM was prepared from basalt and feldspar minerals with minor quantities of finely divided
gold-containing minerals that have been screened to ensure there is no gold nugget effect. There
were 35 data points for this CRM. No failures were returned for the OxB130 CRM. There was a
slight high bias noted.
The SN104 and SQ88 CRMs were supplied by Rocklabs Reference Materials of Aukland, New
Zealand. Both CRMs were prepared from feldspar minerals, basalt and iron pyrites with minor
quantities of finely divided gold and silver-containing minerals that have been screened to ensure
there is no gold nugget effect. The SN104 and SQ88 CRMs are certified for both gold and silver.
There were 142 data points for the SN104 CRM and 73 data points for the SQ88 CRM. No failures
were returned for either the SN104 or SQ88 CRM for either gold or silver. There was a slight high
bias noted for silver in the SN104 data, earlier on in the program, and slight low biases noted for
both gold and silver in the SQ88 data.
The Author considers that the CRMs demonstrate reasonable accuracy in the 2020 data.
11.4.2
Performance of Blanks
All blank data for Au and Ag were graphed. If the assayed value in the certificate was indicated as
being less than detection limit the value was assigned the value of half the detection limit for data
treatment purposes. An upper tolerance limit of three times the detection limit was set. There were
a total of 146 data points to examine.
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All data plotted at or below the set tolerance limits, except for three Ag samples (LRC-007757,
LRC-011739 and LRC-012336, which returned results of 1.6 ppm, 6.1 ppm and 4.1 ppm
respectively). The Author does not consider contamination to be an issue for the 2020 drill data.
11.4.3
Performance of Field Duplicates
Field duplicate data were examined for the 2020 drill program for both gold and silver. There was
a total of 146 duplicate pairs in the data set. Data were scatter graphed and found to have acceptable
precision for the field level for both gold and silver, both with R-squared values of 0.973 (Figures
11.3 and 11.4).
FIGURE 11.3
PERFORMANCE OF AU FIELD DUPLICATES FOR 2020 DRILLING AT LOS
RICOS SOUTH
Source: P&E (2021)
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FIGURE 11.4
PERFORMANCE OF AG FIELD DUPLICATES FOR 2020 DRILLING AT LOS
RICOS SOUTH
Source: P&E (2021)
11.5
2020 UMPIRE SAMPLING PROGRAM
GoGold carried out a comprehensive umpire sampling program of a selection of the 2019 and 2020
drill core samples to verify the primary lab (Actlabs) results. A total of ten drill holes from the
2019 program and eight from the 2020 program were chosen, ensuring that the selected drill holes
were spread out along the length of the deposit, extended to depth and also temporally represented
the 2019 and 2020 drill programs.
The entire sampled length of all 18 selected drill holes were re-assayed at an umpire laboratory,
with samples starting in barren hanging wall material, transitioning to the mineralized zone and
ending up in the low grade to barren footwall. Re-assaying entire drill hole lengths in this manner
also gives a good representation of all ranges of grades within the Deposit.
A total of 559 reject samples from the 2019 drill core were assayed at ALS and 317 reject samples
from the 2020 drill core were assayed at Bureau Veritas, representing approximately 10% of the
primary samples. Each batch assayed contained a range of QC samples, including three CRMs and
one or two blanks and samples were assayed by the same method as the original primary laboratory
analysis.
Table 11.4 outlines a summary of drill holes selected for umpire sampling and reject samples sent
for assaying.
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TABLE 11.4
UMPIRE SAMPLING PROGRAM DRILL HOLES
Sample No.
From
Sample No.
To
Reject
Samples
QC
Samples
2019 Drill Holes
LRGG-19-011 ALS Chemex
LRGG-19-017 ALS Chemex
LRGG-19-022 ALS Chemex
LRGG-19-039 ALS Chemex
LRGG-19-050 ALS Chemex
LRGG-19-054 ALS Chemex
LRGG-19-076 ALS Chemex
LRGG-19-085 ALS Chemex
LRGG-19-086 ALS Chemex
LRGG-19-092 ALS Chemex
Total
LRC-002053
LRC-002389
LRC-003268
LRC-003730
LRC-004291
LRC-004619
LRC-005458
LRC-006563
LRC-006042
LRC-006750
LRC-002101
LRC-002455
LRC-003306
LRC-003794
LRC-004355
LRC-004656
LRC-005527
LRC-006632
LRC-006084
LRC-006821
46
64
37
62
72
36
66
67
41
68
559
5
5
5
5
5
5
5
5
5
5
50
2020 Drill Holes
LRGG-20-095 Bureau Veritas
LRGG-20-106 Bureau Veritas
LRGG-20-119 Bureau Veritas
LRGG-20-129 Bureau Veritas
LRGG-20-137 Bureau Veritas
LRGG-20-143 Bureau Veritas
LRGG-20-155 Bureau Veritas
LRGG-20-161 Bureau Veritas
Total
LRC-006970
LRC-007653
LRC-007989
LRC-010641
LRC-010765
LRC-011113
LRC-011760
LRC-011869
LRC-007062
LRC-007700
LRC-010019
LRC-010667
LRC-010810
LRC-011162
LRC-011819
LRC-011899
76
39
26
22
37
41
50
26
317
5
5
5
5
5
5
5
5
40
Drill Hole ID
Umpire Lab
Source: P&E (2021)
The Author has reviewed the umpire assay results for both the 2019 and 2020 programs and
comparison was made between the primary lab results and the umpire lab results by way of line
graph and scatter plots (Figures 11.5 to 11.8). The data for both ALS and Bureau Veritas indicate
no material biases between the umpire laboratories and gold and silver assays from Actlabs, the
primary laboratory.
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FIGURE 11.5
2019 UMPIRE SAMPLING ASSAYS FOR AU
Source: P&E (2021)
FIGURE 11.6
2019 UMPIRE SAMPLING ASSAYS FOR AG
Source: P&E (2021)
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FIGURE 11.7
2020 UMPIRE SAMPLING ASSAYS FOR AU
Source: P&E (2021)
FIGURE 11.8
2020 UMPIRE SAMPLING ASSAYS FOR AG
Source: P&E (2021)
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11.6
JUL 2020 - MAY 2023 QUALITY ASSURANCE/QUALITY CONTROL REVIEW
GoGold continued implementing industry-standard QC protocol for the next phase of drilling at
Los Ricos South from July 2020 to May 2023. CRMs were inserted approximately every 1 in 9
samples, blanks approximately every 1 in 21 samples, and field duplicates consisting of ¼ drill
core were collected approximately every 21 samples.
11.6.1
Performance of Certified Reference Materials
CRMs were inserted into the analysis stream approximately every 9 samples. GoGold continued
to utilize the AR09003X, OxB130, SN104, and SQ88 CRMs throughout July 2020 to May 2023
and also introduced two new CRMs, from Rocklabs: the Oxi164 (gold-only) and SN118 (silver
and gold) CRMs. Criteria for assessing CRM performance remained as described in section 11.3.1
of this Technical Report. The Oxi164 gold CRM was supplied by Rocklabs and was prepared from
basalt and feldspar minerals with minor quantities of finely divided gold-containing minerals,
which were screened to ensure that there is no gold nugget effect. The SN118 CRM was also
supplied by Rocklabs and was prepared from feldspar minerals, basalt and iron pyrites with minor
quantities of finely divided gold and silver-containing minerals, which were screened to ensure
that there is no gold nugget effect.
A summary of the CRM performance results for the July 2021 to May 2023 drilling at the Project
is presented in Table 11.5 and performance charts for each CRM are displayed in Figures 11.9 to
11.17.
TABLE 11.5
SUMMARY OF CERTIFIED REFERENCE MATERIALS USED AT LOS RICOS SOUTH
Certified
Reference
Material
Certified
Mean
Value
(ppm)
Monitoring Gold
AR09003X
1.231
OxB130
0.125
Oxi164
1.790
SN104
9.182
SN118
8.917
SQ88
39.723
Monitoring Silver
SN104
46.7
SN118
49.9
SQ88
160.8
Lab Results
+/- 1SD
(ppm)
+/- 2SD
(ppm)
0.08
0.006
0.036
0.184
0.168
0.947
1.4
2.1
5.1
No.
Results
No. (-)
Failures
No. (+)
Failures
Average
Result
(ppb)
0.16
0.012
0.072
0.368
0.336
1.894
14
36
902
199
741
313
0
0
7
1
8
1
0
0
8
0
12
0
1.230
0.131
1.805
9.136
8.967
39.54
2.8
4.2
10.2
199
741
313
0
0
8
6
0
4
47.62
50.78
156.12
Source: P&E (2024)
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The Author considers that CRM performance was acceptable, with an overall failure rate of around
1.5% for both the gold and silver data. As indicated in Table 11.5, the majority of the CRMs
performed well, with very few failures recorded. The SQ88 CRM showed the highest failure rate
for silver, at 3.8%, followed by the SN104 CRM, with a failure rate of 3.0%. The SN118 CRM
revealed the highest failure rate for gold, at 2.7%, followed by the Oxi164 CRM, with a failure
rate of 1.7%. All other CRMs recorded failure rates below 0.5%. The Author reviewed all CRM
failures and found all failures to be minority failures within a particular batch, with several other
CRMs passing within each respective batch. The Author considers that the CRM data demonstrate
acceptable accuracy in the July 2020 to May 2023 Los Ricos South data.
FIGURE 11.9
PERFORMANCE OF AU AR09003X CRM FOR JUL 2020 – MAY 2023
DRILLING AT PROJECT
Source: P&E (2024)
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FIGURE 11.10
PERFORMANCE OF AU OXB130 CRM FOR JUL 2020 – MAY 2023
DRILLING AT PROJECT
Source: P&E (2024)
FIGURE 11.11
PERFORMANCE OF AU OXI164 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
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FIGURE 11.12
PERFORMANCE OF AU SN104 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
FIGURE 11.13
PERFORMANCE OF AG SN104 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
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FIGURE 11.14
PERFORMANCE OF AU SN118 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
FIGURE 11.15
PERFORMANCE OF AG SN118 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
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FIGURE 11.16
PERFORMANCE OF AU SQ88 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
FIGURE 11.17
PERFORMANCE OF AG SQ88 CRM FOR JUL 2020 – MAY 2023 DRILLING
AT PROJECT
Source: P&E (2024)
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11.6.2
Performance of Blanks
All blank data for gold and silver were reviewed by the Author. If the assayed value in the
certificate was indicated as being less than detection limit, the value was assigned the value of onehalf the detection limit for data treatment purposes. An upper tolerance limit of five times the
detection limit was set. There were 945 data points to examine.
All silver data and the vast majority of gold data plots at or below the set tolerance limits (Figures
11.18 and 11.19) and the Author does not consider the very few gold outliers to be significant to
the integrity of the data.
FIGURE 11.18
PERFORMANCE OF AU BLANK FOR JUL 2020 – MAY 2023 DRILLING AT
PROJECT
Source: P&E (2024)
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FIGURE 11.19
PERFORMANCE OF AG BLANK FOR JUL 2020 – MAY 2023 DRILLING AT
PROJECT
Source: P&E (2024)
11.6.3
Performance of Field Duplicates
Field duplicate data for gold and silver were examined for the July 2020 to May 2023 drilling at
Los Ricos South. There were 945 duplicate pairs in the dataset. Data were scatter graphed (Figures
11.20 and 11.21) and found to have acceptable precision at the field level for gold and silver, with
R-squared values of 0.966 and 0.998, respectively, and the majority of the data plotting close to
the 1:1 line.
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FIGURE 11.20
PERFORMANCE OF AU FIELD DUPLICATES FOR JUL 2020 – MAY 2023
DRILLING AT PROJECT
Source: P&E (2023)
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FIGURE 11.21
PERFORMANCE OF AG FIELD DUPLICATES FOR JUL 2020 – MAY 2023
DRILLING AT PROJECT
Source: P&E (2023)
11.7
2023 UMPIRE SAMPLING PROGRAM
GoGold carried out a comprehensive umpire sampling program of a selection of the 2020 to 2022
Los Ricos South drill core samples, to verify the primary labs’ (Actlabs and ALS) results. A total
of four drill holes from the 2020 program, two from the 2021 program and 23 from the 2022
program were chosen to verify, ensuring that the selected drill holes were spread out along the
length of the deposit, extended to depth and also temporally represented the 2020 to 2022 drilling.
The entire sampled length of all 29 selected drill holes were re-assayed at an umpire laboratory,
with samples starting in barren hanging wall material, transitioning to the mineralized zone and
ending up in the low grade to barren footwall. Re-assaying entire drill hole lengths in this manner
also gives a good representation of all ranges of grades within the deposit.
A total of 985 pulp samples from the original drill core samples were assayed at Bureau Veritas,
representing approximately 5% of the primary samples. Each batch assayed contained a range of
QC samples, including CRMs and blanks and samples were assayed by the same method as the
original primary laboratory analysis.
Table 11.6 outlines a summary of drill holes selected for umpire sampling and pulp samples sent
for assaying.
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The Author has reviewed the umpire assay results for the 2020 to 2022 drilling at Los Ricos South,
and comparison was made between the primary lab results and the umpire lab results, with the aid
of line graph and scatter plots (Figures 11.22 and 11.23). The data indicate no material biases in
the gold and silver assays between the primary and umpire laboratories.
TABLE 11.6
UMPIRE SAMPLING PROGRAM DRILL HOLES
Drill Hole ID
Area
Umpire Lab
LRGG-20-177
LRGG-20-185
LRGG-20-193
LRGG-20-203
LRGG-22-216
LRGG-22-228
LRGG-22-232
LRGG-22-241
LRGG-22-256
LRGG-22-262
LRGG-22-271
LRGAG-21-008
LRGAG-21-016
LRGAG-22-027
LRGAG-22-030
LRGAG-22-044
LRGAG-22-054
LRGAG-22-063
LRGAG-22-072
LRGAG-22-085
LRGAG-22-095
LRGAG-22-106
LRGAG-22-112
LRGAG-22-120
LRGAG-22-134
LRGAG-22-144
LRGAG-22-156
LRGAG-22-161
LRGAG-22-173
Total
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Abra
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
El_Aguila
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Bureau Veritas
Sample No.
From
LRC-012559
LRC-013010
LRC-013752
LRC-014089
LRC-106130
LRC-108117
LRC-107446
LRC-106601
LRC-108412
LRC-109041
LRC-142126
LRC-095435
LRC-096355
LRC-097139
LRC-099209
LRC-097455
LRC-097693
LRC-097901
LRC-102168
LRC-102299
LRC-103232
LRC-103653
LRC-103844
LRC-106052
LRC-140424
LRC-142344
LRC-142614
LRC-143402
LRC-147188
Sample No.
To
LRC-012614
LRC-013035
LRC-013785
LRC-014107
LRC-106150
LRC-108137
LRC-107467
LRC-106629
LRC-108427
LRC-109081
LRC-142148
LRC-095466
LRC-096382
LRC-097170
LRC-099248
LRC-097510
LRC-097718
LRC-097929
LRC-102188
LRC-102317
LRC-103263
LRC-103688
LRC-103868
LRC-106086
LRC-140476
LRC-142366
LRC-142664
LRC-143464
LRC-147263
Pulp
Samples
56
26
34
19
21
21
22
29
16
41
23
32
28
32
40
56
26
29
21
19
32
36
25
35
53
23
51
63
76
985
Source: P&E (2023)
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FIGURE 11.22
2023 UMPIRE SAMPLING ASSAYS FOR AU
Source: P&E (2023)
FIGURE 11.23
2023 UMPIRE SAMPLING ASSAYS FOR AG
Source: P&E (2023)
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11.8
CONCLUSIONS
It is the Author’s opinion that sample preparation, security and analytical procedures for the Los
Ricos South Project drill programs were adequate, and that the data is of good quality and
satisfactory for use in the current Mineral Resource Estimate. It is recommended that GoGold
continue with the current sampling and data collection procedures as further drilling at the Project
is undertaken to further define, increase confidence and expand on the current Mineral Resource.
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12.0
DATA VERIFICATION
12.1
DATABASE VERIFICATION
The Project data is stored in the GVMapper database. This database is secure, operated by a single
database administrator in the SPM office located in Hermosillo, Mexico and contains data
checking routines designed to prevent common data entry errors.
Assay data for the legacy RC holes completed by TUMI Resources in 2003-2005 were checked
by the Authors. The review consisted of checking the digital data against source documents to
ensure proper data entry, as well as data integrity checks for overlapping intervals, data beyond
total depth of hole. No errors were identified during the review. To date, none of the original assay
certificates from the TUMI work have been viewed by the Authors.
Industry standard validation checks were carried out on the supplied databases, and minor
corrections made where necessary. The Authors typically validate a Mineral Resource database by
checking for inconsistencies in naming conventions or analytical units, duplicate entries, interval,
length or distance values less than or equal to zero, blank or zero-value assay results,
out-of-sequence intervals, intervals or distances greater than the reported drill hole length,
inappropriate collar locations, and missing interval and coordinate fields.
The Authors also conducted verification of the Los Ricos South Project drill hole assay database
for gold and silver in 2021, by comparison of the database entries with assay certificates, supplied
to the Authors by Actlabs Guadalupe, Zacatecas, México, in comma-separated values (csv) format
and Portable Document Format (pdf) format. Assay data ranging from 2019 through 2020 were
verified for the Project. Approximately 87% (9,053 out of 10,435 samples) of the database was
checked for gold and silver, and approximately 81% (3,019 out of 3,750 samples) of the
constrained data were verified. Very few minor discrepancies were noted in the data, which were
not material to the Mineral Resource Estimate.
The Authors again conducted verification of the Los Ricos South drill hole assay database for gold
and silver in 2023, by comparison of the database entries with assay certificates, downloaded
directly by the Authors from the ALS Webtrieve™ online portal in comma-separated values (csv)
format and Portable Document Format (pdf) format. Assay data ranging from 2020 through 2023
were verified for the Project. Approximately 87% (17,219 out of 19,725 samples) of the updated
data was verified for gold and silver. Very few minor discrepancies were noted in the data, which
were not material to the Mineral Resource Estimate.
The Authors also conducted verification of the updated Los Ricos South drill hole assay data for
gold, silver, copper, lead and zinc in 2024, by comparison of the database entries with assay
certificates, downloaded directly by the Authors from the ALS Webtrieve™ online portal in
comma-separated values (csv) format and Portable Document Format (pdf) format. Assay data
ranging from April 2023 to May 2023 were verified for the Project. Approximately 76% (276 out
of 364 samples) of the updated data was verified for gold and silver. No errors were encountered
in the data.
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12.2
P&E SITE VISIT AND INDEPENDENT SAMPLING
The Los Ricos South Project was visited by independent P&E Qualified Persons Mr. Fred Brown,
P.Geo., August 15 and 16, 2019, and by Mr. David Burga, P.Geo., May 16 and 17, 2023, for the
purpose of completing site visits that included review of the logging and drill core storage facilities,
drilling sites, outcrops, GPS location verifications, discussions and due diligence sampling.
Mr. Brown collected ten samples from nine diamond drill holes during the August 15 and 16, 2019
site visit. All samples were selected from drill holes completed in 2019.
A selection of broad mineralized intervals was sent in advance by Mr. Burga to GoGold and these
intervals were pulled from storage and brought to the drill core facility in advance of Mr. Burga’s
arrival. GoGold employees were not aware of specific samples to be taken ahead of the site visit.
Mr. Burga collected 12 samples from 12 diamond drill holes completed in 2020, 2022 and 2023.
A range of high, medium, and low-grade samples were selected from the stored drill core. Samples
were collected by taking a quarter of the remaining half drill core, with the other quarter drill core
returned to the drill core box. Individual samples were placed in plastic bags with a uniquely
numbered tag, after which all samples were collectively placed in a larger bag and delivered by
Mr. Brown to the ALS Global laboratory in Guadalajara, Mexico (2019) or by Mr. Burga to the
Activation Laboratories Ltd. facility in Ancaster, Ontario (2023) for analysis.
Samples at ALS and Actlabs were analyzed for gold and silver by fire assay with a gravimetric
finish. Bulk density determinations were determined by pycnometer on all 2019 samples and by
water displacement on all 2023 samples.
ALS has developed and implemented strategically designed processes and a global quality
management system at each of its locations. The global quality program includes internal and
external inter-laboratory test programs and regularly scheduled internal audits that meet all
requirements of ISO/IEC 17025:2017 and ISO 9001:2015. All ALS geochemical hub laboratories
are accredited to ISO/IEC 17025:2017 for specific analytical procedures.
The Actlabs’ Quality System is accredited to international quality standards through ISO/IEC
17025:2017 and ISO 9001:2015. The accreditation program includes ongoing audits, which verify
the QA system and all applicable registered test methods. Actlabs is also accredited by Health
Canada. Both ALS and Actlabs are independent of GoGold and SPM.
Results of the Los Ricos South site visit verification samples for gold and silver are presented in
Figures 12.1 through 12.4.
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Results of the site visit due diligence samples are presented in Figures 12.6 and 12.7.
FIGURE 12.6
FIGURE 12.1
LOS RICOS SITE VISIT DUE DILIGENCE SAMPLE RESULTS FOR GOLD,
RA
ESULTS
THE AUGUST 2019 AU VERIFICATION SAMPLES (ALS)
UGUSTOF
2019
Los Ricos Project
P&E Site Visit Due Diligence Samples for Au: Aug 2019
3.0
GoGold
P&E
2.5
Au g/t
2.0
1.5
1.0
0.5
0.0
Drill Hole Number
Source: P&E (2021)
FIGURE 12.7
FIGURE 12.2
LOS RICOS SITE VISIT DUE DILIGENCE SAMPLE RESULTS FOR SILVER,
RA
ESULTS
THE AUGUST 2019 AG VERIFICATION SAMPLES (ALS)
UGUSTOF
2019
Los Ricos Project
P&E Site Visit Due Diligence Samples for Ag: Aug 2019
350
300
GoGold
P&E
Ag g/t
250
200
150
100
50
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Drill Hole Number
Source: P&E (2021)
Based upon the evaluation of the QA/QC program and P&E’s due diligence sampling, it is the
author’s opinion that the results are suitable for use in the current Technical Report.
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FIGURE 12.3
RESULTS OF THE MAY 2023 AU VERIFICATION SAMPLES (ACTLABS)
Source: P&E (2023)
FIGURE 12.4
RESULTS OF THE MAY 2023 AG VERIFICATION SAMPLES (ACTLABS)
Source: P&E (2023)
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12.3
CONCLUSION
Verification of the Los Ricos South Project data, used for the current Mineral Resource Estimate,
has been undertaken by the Authors, including site visits, due diligence sampling, verification of
drill hole assay data from electronic assay files, and assessment of the available QA/QC data. The
Authors consider that there is good correlation between the gold and silver assay values in
GoGold’s database and the independent verification samples collected by the Authors and
analyzed at ALS and Actlabs. The Authors are satisfied that sufficient verification of the drill hole
data has been undertaken and that the supplied data are of good quality and appropriate for use in
the current Mineral Resource Estimate.
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13.0
MINERAL PROCESSING AND METALLURGICAL TESTING
13.1
INTRODUCTION
The earliest metallurgical data on the Los Ricos South (“LRS”) Property comes from the Cinco
Minas Mining Company (“CMMC”) which indicates the 500 tons per day flotation and
cyanidation processing operation had recovery rates for both silver and gold greater than 90%
between 1918 and 1930.
Metallurgical testwork programs were undertaken in 2021, 2022 and 2023 in support of previous
Project evaluations before this Feasibility Study (“FS”) started. The testwork programs were
completed at SGS and Metso facilities.
A series of laboratory testing (FS testwork program), commencing in February 2024, of the LRS
mineralization was initiated at SGS Lakefield (“SGS”) located in Ontario, Canada. Samples were
comprised of composites exclusively from the Eagle, Abra and Cerro Colorado (“CC”) Zones.
As part of the FS testwork program, Metso Outotec Canada (“Metso Outotec”) conducted
advanced solid-liquid testwork, with the objective to investigate gold, copper and silver recovery
potential using vacuum filtration washing, pre-leach and post leach thickening and tailings
dewatering. The samples were exclusively from the Eagle and Abra Zones.
The testwork program undertaken to support the FS included:
•
•
•
•
•
•
•
•
13.2
Head grade elemental, gold and silver deportment and mineralogical analysis of
Deposit samples;
Comminution testwork;
Investigation of pre-concentration options, including gravity separation;
Cyanide leaching;
Precious metal recovery testing (Merrill Crowe process);
Solid-liquid separation testing;
Sulphidization, acidification, recycling, and thickening (“SART”) testing; and
Cyanide detoxification.
METALLURGICAL TESTWORK
Several metallurgical test programs have been completed on the LRS Project, under various
owners, since 1910. These programs were designed to quantify metallurgical performance and
response to different processing options over a range of samples. A summary of the historical test
programs is included in Table 13.1.
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TABLE 13.1
LRS METALLURGICAL TESTWORK PROGRAM SUMMARY
Year
Laboratory/
Location
19101911
Unknown
20202021
SGS Lakefield
20212022
SGS Lakefield
Metso Outotec
2023
SGS Lakefield
Testwork Performed
Investigation into the recovery of gold
and silver, by pre-concentration
(gravity) and cyanidation of tailings and
whole ore.
Preliminary comminution testing, whole
ore cyanide leaching and acid-based
accounting (“ABA”)
The
SGS
program
included
comminution
testing,
feed
characterization, whole ore leaching,
flotation, Merrill Crowe, SART and
ABA testwork.
Metso Outotec carried out solid/liquid
separation testwork.
The
SGS
program
included
comminution
testing,
feed
characterization, mineralogy, whole ore
leaching, flotation, Merrill Crowe,
SART,
gravity
pre-concentration
testwork,
solid/liquid
separation
testwork and ABA testwork.
The
SGS
program
included
comminution, sample characterization,
gravity pre-concentration, whole ore
cyanidation leaching, Merrill Crowe,
SART testwork, mineralogy, gold and
silver
deportment
and
cyanide
detoxification testwork.
Metso Outotec carried out solid/liquid
separation testwork.
2024
SGS Lakefield
Metso Outotec
13.2.1
GoGold Resources Testwork
Samples Tested
Unknown
Two composite samples
from the Abra Zone.
One master composite
and
six
variability
samples from the Abra
Zone.
Three Eagle Zone hole
composite samples
Three
composite
samples each from the
Eagle and Abra Zones.
One composite sample
from the CC Zone.
Variability
samples
including seven from the
Eagle Zone, five from
the Abra Zone and two
from the CC Zone
Metallurgical testing carried out by GoGold from 2020 to 2024 is summarized in this section.
Results from past programs (2020-2021, 2021-2022 and 2023) that are relied upon to support the
process plant design for the FS are included.
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13.2.1.1
Head Assays
SGS-Lakefield Program (2020-2021)
The two Abra Zone composites used in the SGS leaching testwork program were analyzed for
head assay, with the results summarized in Table 13.2.
TABLE 13.2
2020-2021 ABRA ZONE HEAD ASSAYS
Sample ID
Unit
Comp 1
Comp 2
Au
g/t
0.92
3.25
Ag
g/t
195
314
S
%
0.41
0.56
S=
%
0.38
0.51
Cu
%
0.066
0.113
Fe
%
1.5
1.22
Pb
%
0.56
1.84
Sb
g/t
<20
<20
Zn
g/t
8,060
7,720
The Comp 1 sample graded 0.92 g/t Au and 195 g/t Ag, while Comp 2 had higher grades of 3.25
g/t Au and 314 g/t Ag. The ICP results indicated the presence of Cu in both samples, with measured
levels of 653 g/t and 1,130 g/t in Comp 1 and Comp 2, respectively.
SGS-Lakefield Program (2021-2022)
Head assay results of the one composite and six variability samples from the Abra Zone in the
LRS Deposit tested in the 2021-2022 program are shown in Table 13.3. The gold and silver assay
values in the table were determined by fire assay and are the average of duplicate tests.
TABLE 13.3
2021-2022 ABRA ZONE COMPOSITE AND VARIABILITY HEAD ASSAYS
Element
Unit
Au
Ag
S
S=
Cu
Pb
g/t
g/t
%
%
%
%
Master
Composite
1.32
125
0.14
0.13
0.220
0.53
1
0.75
48.8
0.14
0.15
0.034
0.07
Variability Sample ID
2
3
4
5
2.65
1.22
1.36
2.01
170
224
250
315
0.07
0.08
0.11
0.24
0.10
0.09
0.12
0.23
0.027
0.042
0.044 0.059
0.05
0.16
0.11
0.39
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6
1.82
436
0.31
0.30
0.068
1.14
Page 146 of 577
TABLE 13.3
2021-2022 ABRA ZONE COMPOSITE AND VARIABILITY HEAD ASSAYS
Element
Unit
As
Sb
Zn
g/t
g/t
%
Master
Composite
<50
<20
0.13
1
<50
<20
0.14
Variability Sample ID
2
3
4
5
<50
<50
<50
<50
<20
<20
<20
<20
0.90
0.25
0.16
0.44
6
<50
<20
0.56
The gold head grade of the master composite was 1.32 g/t while the variability samples ranged
from 0.75 g/t to 2.65 g/t. The silver head grade of the master composite was 125 g/t and the
variability samples ranged from 48.8 g/t to 436 g/t. The samples generally had low total sulphur
and sulphide sulphur values, ranging from 0.07% to 0.31% ST and 0.09% to 0.30% S=,
respectively. The master composite copper head grade was 0.22%, while the six variability sample
copper levels were lower and ranged from 0.027% to 0.068%.
SGS-Lakefield Program (2023)
Head sample analyses were completed on the Eagle Zone 2023 metallurgical composites to
determine the precious metal concentrations as well as the content of other elements. Assays
included Au and Ag in duplicate by fire assay, sulphur speciation for total sulphur (ST) and
sulphide sulphur (S=) by Leco analyzer, Cu by atomic absorption, and a semi-quantitative
Inductively Coupled Plasma (ICP) scan analysis. Results are included in Table 13.4.
TABLE 13.4
2023 EAGLE ZONE HOLE COMPOSITE HEAD ASSAYS
Hole Comp Hole Comp Hole Comp
012
031
014/035
Element
Unit
Au (Avg)
g/t
2.87
5.02
4.65
Ag (Avg)
g/t
110
132
188
ST
%
0.89
0.69
0.54
S=
%
0.84
0.67
0.49
Cu
%
0.22
0.47
0.28
Fe
%
3.12
4.23
2.13
Pb
%
0.93
0.52
0.53
Zn
%
1.24
0.89
0.56
As
g/t
<30
34
52
Sb
g/t
<10
<10
<10
The gold and silver head grades of the three Eagle hole composite samples ranged between 2.87
g/t to 5.02 g/t and 110 g/t and 188 g/t, respectively. The samples generally had low total sulphur
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and sulphide sulphur values, ranging from 0.49% to 0.84% ST and 0.54% to 0.49% S=,
respectively. The samples contain appreciable base metals concentrations with copper head grades
between 0.22% and 0.47%, lead between 0.52% and 0.93% and zinc levels between 0.56% and
1.24%.
SGS-Lakefield Program (2024)
In the 2024 SGS Program, three master composites of the Abra and Eagle Zones and one master
composite of the CC Zone was generated for metallurgical testing. Seven Eagle Zone, five Abra
Zone and two CC Zone variability samples were also generated. Multi-element and ICP
spectroscopy head assay analysis was completed on all samples. A summary of the head assays
and select ICP results for the key Abra, Eagle and CC master and variability composites are
presented in Table 13.5, Table 13.6, and Table 13.7.
Screened Metallics (“SM”) assays are generated by screening a 0.5 kg aliquot of the sample at 150
microns. The oversize is fire assayed to extinction and the undersize is sub-sampled and fire
assayed. The two are combined to generate an overall assay for the sample. This assay method is
useful for samples that contain high levels of coarse, free gold. The SM assays do not show
significant coarse free gold content in any of the samples.
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TABLE 13.5
2024 ABRA ZONE SAMPLES HEAD ASSAYS
Element
Au by Fire Assay (FA)
Au by Screened Metallics
method (SM)
Ag by FA
Ag by SM
ST
S=
CT
Cu
Fe
Pb
Zn
As
Hg
Sb
Low
Unit Grade
Comp
g/t
0.59
Med
Grade
Comp
1.31
High
Var.
Var.
Var.
Var.
Var.
Grade Sample Sample Sample Sample Sample
Comp
#1
#2
#3
#4
#5
2.59
0.32
0.48
1.02
1.37
1.77
g/t
0.50
1.35
2.73
0.28
0.48
0.91
1.40
1.32
g/t
g/t
%
%
%
%
%
%
%
g/t
g/t
g/t
76
85
0.55
0.46
0.13
0.22
2.22
0.19
0.38
<30
0.7
<20
158
184
0.16
0.12
0.11
0.09
1.51
0.15
0.16
<30
0.7
<20
213
243
0.95
0.84
0.06
0.30
1.52
0.58
1.24
<30
1.1
<20
49
43
0.17
0.14
0.01
0.10
1.63
0.12
0.31
44
1.0
<20
89
82
0.15
0.12
0.03
0.17
2.97
0.08
0.29
40
1.2
<20
122
121
0.45
0.40
0.02
0.21
2.12
0.16
0.49
<30
0.9
<20
218
229
0.40
0.33
0.01
0.11
2.20
0.31
0.93
43
1.3
<20
367
345
0.94
0.83
0.03
0.24
2.32
0.51
0.96
37
2.4
<20
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TABLE 13.6
2024 EAGLE ZONE SAMPLES HEAD ASSAYS
Element
Unit
Low
Grade
Comp
Medium
Grade
Comp
High
Grade
Comp
Var.
Sample
#1
Var.
Sample
#2
Var.
Sample
#3
Var.
Sample
#4
Var.
Sample
#5
Var.
Sample
#6
Var.
Sample
#7
Au by FA
g/t
0.91
2.36
12.0
0.39
0.48
0.90
1.36
1.72
3.17
5.37
Au by SM
g/t
0.82
2.04
9.24
0.37
0.45
0.88
1.39
1.67
2.77
5.58
Ag by FA
g/t
93
219
532
22
41
82
98
173
196
144
Ag by SM
g/t
104
212
520
27
41
87
112
173
213
156
ST
%
0.65
0.80
1.61
0.83
1.02
1.10
1.12
1.01
3.19
1.65
S=
%
0.56
0.68
1.38
0.75
0.95
0.95
1.02
0.89
2.89
1.49
CT
%
0.04
0.05
0.05
0.03
0.05
0.04
0.06
0.11
0.06
0.06
Cu
%
0.38
0.31
0.47
0.22
0.21
0.19
0.69
1.37
0.48
0.85
Fe
%
3.23
3.10
3.40
4.75
2.86
3.03
3.68
4.51
4.08
3.62
Pb
%
0.29
0.48
1.19
0.12
0.13
0.22
1.11
0.48
1.46
1.04
Zn
%
0.38
0.49
1.11
0.24
0.35
0.63
1.22
0.67
2.20
1.88
As
g/t
64
40
<30
<30
<30
31
112
<30
45
182
Hg
g/t
2.6
1.7
2.3
0.7
0.8
2.5
2.9
1.7
2.7
1.5
Sb
g/t
<20
<20
<20
<20
<20
<20
<20
<20
<20
<20
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TABLE 13.7
2024 CC ZONE SAMPLES HEAD ASSAYS
Element
Au by FA
Au by SM
Ag by FA
Ag by SM
ST
S=
CT
Cu
Fe
Pb
Zn
As
Hg
Sb
As
Unit
Comp
g/t
g/t
g/t
g/t
%
%
%
%
%
%
%
g/t
g/t
g/t
g/t
1.00
0.89
46
47
0.51
0.45
0.57
0.02
1.67
0.03
0.05
<30
<0.3
<10
<30
Var.
Sample
#1
1.08
0.91
67
73
0.82
0.75
0.08
0.02
1.35
0.02
0.02
<30
0.5
17
<30
Var.
Sample
#2
1.62
1.82
59
60
1.72
1.65
0.45
0.01
2.18
<0.01
<0.01
<30
<0.3
<10
<30
Gold and silver sample grades for each zone ranged between:
•
•
•
Abra; 0.28 g/t to 2.73 g/t gold and 43 g/t to 367 g/t silver;
Eagle; 0.37 g/t to 12.0 g/t gold and 22 g/t to 532 g/t silver; and
CC; 0.89 g/t to 1.82 g/t gold and 46 g/t to 73 g/t silver.
Total carbon and sulphur contents in the Eagle and Abra composites were low, with carbon ranging
from 0.03% to 0.25% and sulphur from 0.12% to 1.65%. Arsenic levels were generally below
detection but reached up to 182 g/t in the Eagle sample and 44 g/t in Abra. Copper content varied
from 0.21% to 1.37% in Eagle and 0.09% to 0.30% in Abra. Lead and zinc ranged from trace to
1.19% for lead in and from trace to 2.20% for zinc. Eagle samples had the highest lead and zinc
concentrations. All samples contained <20 g/t antimony, which did not affect leaching or recovery.
Mercury levels ranged from <0.3 g/t to 2.9 g/t, with no observed adverse recovery effects but
requiring capture in downstream processes.
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13.2.1.2
Mineralogy
SGS-Lakefield Program (2021-2022)
A subsample of the Abra master composite underwent semi-quantitative XRD analysis using the
Rietveld method. Full mineral abundance data is reported in Table 13.8.
TABLE 13.8
2021-2022 ABRA ZONE XRD
ANALYSIS RESULTS
Minerals
Quartz
Orthoclase
Calcite
Epidote
Chlorite
Stilpnomelane
Muscovite
Biotite
Chalcopyrite
Cinoclase
Albite
Total
% w/w
73.0
7.0
3.3
0.7
5.6
0.7
3.2
1.6
0.2
1.7
3.0
100.0
Major minerals (>30%) included quartz, while detected minor minerals (2% to 10%) were
orthoclase, chlorite, calcite, muscovite and albite. Trace minerals (<2%) included cinoclase,
biotite, epidote, stilpnomelane and chalcopyrite.
SGS-Lakefield Program (2023)
A detailed mineralogical analysis of the 2023 Eagle hole composite samples was performed using
scanning electron microscopy (“QEMSCAN”) and XRD. Mineral composition results, in weight
percentages, are presented in Table 13.9.
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TABLE 13.9
2023 EAGLE ZONE HOLE QEMSCAN AND XRD RESULTS
Minerals
Quartz
Chlorite
Calcite
Hematite
Orthoclase
Albite
Galena
Sphalerite
Pyrite
Goethite
Ilmenite
Kaolinite
Total
Hole Comp
012
52.8
5.9
18.8
2.9
17.6
0.4
0.6
1.0
100.0
Unit
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
Hole Com
031
76.2
5.1
0.6
3.7
10.1
0.7
0.2
0.3
0.5
0.2
0.1
2.3
100.0
Hole Comp
014/035
65.5
4.5
16.9
1.8
10.0
0.7
0.3
0.2
0.3
100.0
The major mineral (>30%) was quartz, while moderate (10% to 30%) minerals include calcite and
orthoclase. Minor minerals (2% to 10%) were chlorite and hematite. Trace minerals (<2%)
included albite, galena, sphalerite, pyrite, goethite, ilmenite and kaolinite.
SGS-Lakefield Program (2024)
Quantitative mineralogy evaluation of the Eagle, Abra and CC master composite samples by
scanning electron microscopy (QEMSCAN) and XRD was completed. The mineral compositions
of the prepared composites are shown in Table 13.10. Values displayed are in weight percentages.
TABLE 13.10
2024 EAGLE, ABRA, AND CC ZONE QEMSCAN AND XRD RESULTS
Minerals
Unit
Eagle
Eagle
Eagle Abra
Abra
Abra
Cerro
High Medium Low
High Medium Low
Colorado
Grade Grade Grade Grade Grade Grade
Comp
Comp Comp Comp Comp Comp Comp
Quartz
Chlorite
Albite
Kaolinite
Sphalerite
Galena
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
83.2
4.1
0.9
0.7
0.4
0.5
86.0
2.8
0.7
0.5
0.1
0.2
84.3
3.2
0.8
0.4
0.2
0.1
85.9
2.6
0.8
0.4
1.0
0.3
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87.4
3.4
0.8
0.2
0.1
0.1
77.9
4.9
2.8
0.5
0.2
0.1
76.3
6.1
0.8
1.9
0.1
0.1
Page 153 of 577
TABLE 13.10
2024 EAGLE, ABRA, AND CC ZONE QEMSCAN AND XRD RESULTS
Minerals
Unit
Eagle
Eagle
Eagle Abra
Abra
Abra
Cerro
High Medium Low
High Medium Low
Colorado
Grade Grade Grade Grade Grade Grade
Comp
Comp Comp Comp Comp Comp Comp
Chalcopyrite
Rutile
Hematite
Calcite
Pyrite
Orthoclase
Flourite
Epidote
Magnetite
Total
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
% w/w
0.3
1.1
2.1
0.1
0.7
5.1
0.9
100.0
0.2
0.7
2.3
0.6
4.8
0.7
0.2
0.2
100.0
0.2
0.8
2.2
0.4
7.4
0.1
100.0
0.3
0.8
0.7
0.2
0.2
5.2
0.6
0.9
0.0
100.0
0.0
0.7
0.8
0.5
0.1
4.9
0.1
0.8
0.1
100.0
0.1
0.8
1.3
0.7
0.1
10.2
0.1
0.1
100.0
0.2
0.8
3.4
0.2
9.0
0.2
0.3
100.0
The major mineral in the Eagle, Abra, and CC Zone samples was quartz (77.9% to 87.4%), while
minor minerals (2% to 10%) included chlorite (2.8% to 4.9%), and orthoclase (4.8% to 10.2%).
Eagle contains higher minor mineral levels of hematite (2.1% to 2.3%) and CC contained a higher
amount of calcite (3.4%). Sulphide minerals include sphalerite, galena, chalcopyrite and pyrite,
totalling up to 2% in each sample. Overall, there is no indication of smectites, which are poor
filtering clays.
13.2.1.3
Gold and Silver Deportment
SGS-Lakefield Program (2024)
Gold and silver deportment studies were completed on the Eagle, Abra and CC master composite
samples using a TESCAN Integrated Mineral Analyzer (“TIMA”). The silver deportment results
have been summarized below as silver occurs in greater abundance than gold in the Deposit. Figure
13.1 shows the silver deportment results by mineral type.
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FIGURE 13.1
ELEMENTAL SILVER DEPORTMENT OF THE EAGLE, ABRA AND CC ZONE
SAMPLES
Source: Ag TIMA SGS Report (2024)
The major silver bearing minerals for the seven samples include:
•
Acanthite, which ranged between 33.9% to 89.0% for all samples;
•
Stromeyerite, which ranged between 6.2% to 39.5% for all samples;
•
Aguilarite, which ranged between 33.7% and 22.2% for the Eagle medium and CC
samples, respectively;
•
Silver, which ranged between 16.5% and 24.3% for the Eagle low and Abra low
samples respectively, and from 2.8% to 6.3% over the remaining samples; and
•
Pyrite, which measured 32.7% for the CC sample.
Figure 13.2 shows the silver mineral association results. The major silver associated minerals
include:
•
Free silver, which ranges between 6.6% to 56.2% for all samples apart from the Eagle
low and CC samples which measured 0%;
•
Liberated silver minerals, which range between 12.5% to 55.1%;
•
Quartz feldspars, which ranges between 10.8% to 30.3%;
•
Complex minerals, which range between 6.35% to 21.5%; and
•
Galena, which measured 28.2% for the Abra low-grade sample.
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FIGURE 13.2
SILVER MINERAL DEPORTATION OF THE EAGLE, ABRA, AND CC ZONE
SAMPLES
Source: Ag TIMA SGS Report (2024)
A silver grain size assessment indicated the following:
•
Eagle high, where 47.8% of the silver is coarse grained (>75 µm) and the remainder is
below 55 µm;
•
Eagle med, where 30.2% of the silver is coarse grained (70-75 µm) and the remainder
is below 40 µm;
•
Eagle low, where 25.9% of the silver is in the 45 to 50 µm range and the remainder is
below 25 µm;
•
Abra high/med, where 40.3% to 41.2% of the silver is coarse grained (>75 µm) and the
remainder is below 55 µm; and
•
Abra low/CC, where all silver is under 30 µm.
13.2.1.4
Comminution Testwork
SGS-Lakefield Program (2020-2021)
A Los Ricos Vein and Los Ricos Host composite sample were tested for bond ball mill index
(“BWi”), while the Los Ricos Vein sample was solely tested for bond low-energy impact (“CWi”),
bond rod mill index (“RWi”) and bond abrasion index (“Ai”). The RWi and BWi grindability tests
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were performed at respective closing screen sizes of 1,180 µm and 106 µm. Results are
summarized in Table 13.11.
TABLE 13.11
2020-2021 LOS RICOS VEIN AND LOS RICOS HOST
COMMINUTION TEST RESULTS
Sample Name
Los Ricos Vein
Los Ricos Host
CWi
(kWh/t)
7.4
-
RWi
(kWh/t)
15.8
-
BWi
(kWh/t)
19.7
15.3
Ai
(g)
0.72
-
The Los Ricos Vein sample CWi result of 7.4 kWh/t placed it within the 29th hardness percentile
and the RWi of 15.8 kWh/t classified the sample as moderately hard with respect to the SGS
database. The Ai result of 0.74 g placed the sample at the 92nd abrasiveness percentile with respect
to the SGS database.
The Los Ricos Vein and Los Ricos Host samples reported a corresponding BWi of 19.7 kWh/t and
15.3 kWh/t, which respectively classified each sample as very hard and moderately hard with
respect to the SGS database.
SGS-Lakefield Program (2021-2022)
An Abra master composite sample (LRGG-21-Comp) underwent a full JK Drop-Weight Test, a
CWi test, RWi and BWi grindability tests and an Ai test. Four variability samples were also
subjected to SAG Mill Comminution (“SMC”), RWI, BWI and Ai testing. The RWi and BWi
grindability tests were performed at respective closing screen sizes of 1,180 µm and 106 µm. The
results are detailed in Table 13.12.
TABLE 13.12
2021-2022 ABRA ZONE COMMINUTION TEST RESULTS
Sample
Name
LRGG-19-017
LRGG-19-079
LRGG-19-005
LRGG-20-121
LRGG-21-Comp
Axb
(via
JKDWT)
46.9
Axb
(via SMC)
44.4
43.6
37.1
41.0
-
CWi
RWi
BWi
(kWh/t) (kWh/t) (kWh/t)
11.1
16.6
16.6
16.4
17.3
16.3
19.5
19.5
18.5
19.2
20.6
Ai
(g)
0.651
0.665
0.595
0.667
0.743
SMC testing resulted in Axb parameter values of 41.0 to 44.4 for samples LRGG-19-017, LRGG19-079 and LRGG-20-121, indicating medium competence (51.9th to 59.2nd percentile), while
sample LRGG-19-005 had an Axb value of 37.1, indicating moderately competent ore (67.7th
percentile). The JK drop weight testing on the LRGG-21 composite sample returned an Axb value
of 46.9, indicating medium competence (47th percentile).
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The LRGG-21 composite sample CWi result of 11.1 kWh/t indicates medium competency (58th
percentile).
The RWi index results for the five samples ranged from 16.3 kWh/t to 17.3 kWh/t, giving a
hardness range of 70th to 79th percentile, putting all samples in the moderate category when
compared to the SGS database.
The BWi index results for the samples ranged from 18.5 kWh/t to 20.6 kWh/t, giving a hardness
range of 87th to 95th percentile, putting all five samples in the hard category when compared to the
SGS database.
The Ai index results for the samples ranged from 0.595 g to 0.743 g, giving an abrasive range of
84th to 92nd percentile, putting all samples in the moderately abrasive to abrasive category when
compared to the SGS database.
SGS-Lakefield Program (2023)
In 2023, four Eagle Zone hole composite samples underwent SMC, BWI and Ai testing. The BWi
grindability tests were performed at a closing screen size of 106 µm. The test results are detailed
in Table 13.13.
TABLE 13.13
2023 EAGLE ZONE COMMINUTION TEST RESULTS
Axb
(via SMC)
BWi
(kWh/t)
Ai
(g)
Hole 23 Comp
37.8
16.2
0.719
Hole 27 Comp
35.2
18.9
1.037
Hole 37 Comp
41.4
19.0
0.671
Hole 76 Comp
43.0
20.6
0.468
Sample Name
SMC testing resulted in Axb parameter values ranging from 35.2 to 43.0 for all samples, giving a
hardness range of 55th to 73rd percentile, indicating medium to moderate competency. BWi index
results ranged from 16.2 kWh/t to 20.6 kWh/t, giving a hardness range of 72nd to 95th percentile,
indicating hard to very hard material when compared to the SGS database. Ai index results ranged
from 0.468 g to 1.037 g, giving an abrasive range of 73rd to 98th percentile, putting these samples
in the moderately abrasive to very abrasive range when compared to the SGS database.
SGS-Lakefield Program (2024)
In the FS, a comminution testwork program was completed on five Eagle, four Abra and two CC
Zone variability samples. Each variability sample was subjected to SMC, RWI, BWI and Ai
testing. The RWi and BWi grindability tests were performed at respective closing screen sizes of
1,180 µm and 106 µm. The results are presented in Table 13.14.
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TABLE 13.14
2024 EAGLE, ABRA AND CC ZONE
COMMINUTION TEST RESULTS
Sample ID
Eagle Comp #1
Eagle Comp #2
Eagle Comp #3
Eagle Comp #4
Eagle Comp #5
El Abra Comp #1
El Abra Comp #2
El Abra Comp #3
El Abra Comp #4
CC Comp #1
CC Comp #2
Axb
RWi
BWi
(via SMC) (kWh/t) (kWh/t)
40.1
16.8
17.9
43.9
16.2
17.5
42.7
15.0
18.0
36.3
17.2
19.3
42.6
16.3
19.0
55.9
15.7
16.9
49.6
17.3
18.9
41.2
16.5
17.9
52.9
15.3
18.0
70.8
14.4
15.2
39.3
17.4
19.0
Ai
(g)
0.736
0.728
0.780
0.641
0.461
0.493
0.429
0.489
0.386
0.164
0.624
The Eagle Zone SMC Axb parameter values ranged from 36.3 to 43.9, resulting in a hardness
percentile of 53 to 70 in the SGS database. The Abra Zone SMC Axb parameter values ranged
from 41.2 to 55.9, resulting in a hardness range of 34th to 59th percentile in the SGS database. The
CC Zone SMC Axb parameter values were 39.3 and 70.8, resulting in a hardness range of 22nd to
64th percentile in the SGS database. The results indicate Eagle is the most competent ore zone,
followed by Abra and then CC.
The Eagle Zone RWi index results ranged from 15.0 kWh/t to 17.2 kWh/t, giving a hardness range
of 56th to 78th percentile in the SGS database. The Abra Zone RWi index results ranged from 15.3
kWh/t to 17.3 kWh/t, giving a hardness range of 60th to 79th percentile in the SGS database. The
CC Zone RWi index results were 14.4 kWh/t and 17.4 kWh/t, resulting in a hardness range of 50th
to 80th percentile in the SGS database.
The Eagle Zone BWi index results ranged from 17.5 kWh/t to 19.3 kWh/t, giving a hardness range
of 82nd to 91st percentile in the SGS database. The Abra Zone BWi index results ranged from 16.9
kWh/t to 18.9 kWh/t, giving a hardness range of 78th to 89th percentile in the SGS database. The
CC Zone BWi index results were 15.2 kWh/t and 19.0 kWh/t, resulting in a hardness range of 61st
to 90th percentile in the SGS database. The BWi results suggest that Eagle and Abra have similar
hardness and are the hardest ore zones.
The Eagle Zone Ai index results ranged from 0.461 g to 0.780 g, giving an abrasiveness range of
72nd to 93rd percentile in the SGS database. The Abra Zone sample Ai index results ranged from
0.386 g to 0.493 g, giving an abrasiveness range of 63rd to 75th in the SGS database. The CC sample
Ai index results were 0.164 g and 0.624 g, resulting in an abrasiveness range of 29th to 86th
percentile in the SGS database. The Ai results suggest that Eagle is the most abrasive ore zone,
followed by Abra and then CC.
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13.2.1.5
Gravity Concentration Testwork
SGS-Lakefield Program (2024)
The FS metallurgical test program evaluated the response of the Eagle Zone high-grade composite
sample to a gravity pre-concentration step ahead of cyanide leaching. Testing from earlier
programs had demonstrated that a pre-concentration step either via gravity separation or flotation
offered an opportunity to provide improved recoveries.
A 10 kg of the Eagle Zone high-grade composite sample was ground to a k80 of 75 µm and passed
through a Knelson MD-3 gravity concentrator. The concentrate was further upgraded on a Mozley
table, to produce a Mozley concentrate at a target mass recovery of 0.01% of the original feed
sample. The gravity tailings were pre-aerated for four hours and then leached for 96 hours, with
an initial NaCN concentration of 3.0 g/l. Cyanide and lime consumed during the leaching was 14.4
kg/t and 0.34 kg/t, respectively. The gravity recovery and gravity tailings leaching results are
shown in Table 13.15.
TABLE 13.15
EAGLE ZONE HIGH-GRADE SAMPLE GRAVITY PRE-CONCENTRATION AND
GRAVITY TAILINGS LEACH TEST RESULTS
Gravity Distribution
Sample
ID
Eagle
High
Grade
Cyanidation
Extraction
Overall
Recovery
Conc
Mass
(%)
Conc
Au
(%)
Tail
Au
(%)
Conc
Ag
(%)
Tail
Ag
(%)
Au
(%)
Ag
(%)
Au
(%)
Ag
(%)
0.011
5.1
94.9
2.1
97.9
92
89
93
89
Approximately 5.1% of the gold and 2.1% of the silver in the sample reported to the gravity
concentrate through the pre-concentration step. 92% and 89% of the remaining gold and silver,
respectively, in the gravity tailings respectively leached into solution after 96 hours of cyanide
leaching. The overall recovery to gravity concentration and gravity tailings leach was 93% for gold
and 89% for silver, which matched the recoveries of the corresponding Eagle high-grade whole
ore leach (“WOL”) cyanidation test (CN-39).
Gravity gold recovery was low at 5.1% and was not carried into further testing. Gravity
concentration was not included in subsequent testing and not recommended in the plant flowsheet
for similar reasons.
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13.2.1.6
Leach Testwork
SGS-Lakefield Program (2020-2021)
Two composite samples from the Abra Zone were tested to evaluate the response of cyanide
leaching using standard conditions that focussed on the extraction of gold, silver and copper. The
range of leach conditions included:
•
•
•
•
•
•
•
Primary grind k80 size varied between 73 µm to 93 µm;
No pre-aeration;
40% w/w solids;
Lime added to maintain pH at 10.5 to 11;
1 g/L to 2 g/L of NaCN maintained;
Air sparging during leaching; and
Leach retention time of 72 hours to 96 hours.
The test results are presented in Table 13.16.
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TABLE 13.16
2020-2021 ABRA ZONE BOTTLE ROLL LEACH TEST RESULTS
Comp
1
1
1
2
2
2
Test
No.
CN1
CN3
CN4
CN2
CN5
CN6
Feed
Size
k80
(µm)
93
89
88
78
73
73
NaCN
added
(g/L)
2
1
2
2
1
2
Reagent Cons.
(kg/t)
NaCN
3.01
3.45
3.84
4.43
4.23
5.75
CaO
0.48
1.21
1.31
0.43
1.69
0.95
Au Extraction
(%)
8h
~88
89.1
90.5
~92
89.4
92.5
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24 h
~88
91.1
96.9
~92
81.5
96.4
48 h
~88
96.1
91.5
~92
79.8
96.8
72 h
88.3
83.6
84.6
92
91.3
97.0
96 h
90.8
91.5
94.0
94.0
Au Head
Grade
(g/t)
Calculated
0.73
1.03
1.00
2.75
3.32
3.19
Page 162 of 577
Ag Extraction
(%)
8h
34.2
20.8
37.4
28.7
17.8
33.1
24 h
58.2
31.8
58.0
55.0
42.8
54.2
48 h
80.9
53.6
70.2
69.7
61.5
73.1
72 h
78.4
70.0
84.8
80.0
77.2
90.8
96 h
81.4
85.3
90.8
90.8
Ag Head
Grade
(g/t)
Calculated
209
217
213
325
306
605
Cu Extraction
(%)
8h
57.3
58.2
62.3
60.8
59.0
67.1
24 h
61.8
61.9
63.2
66.1
66.5
67.0
48 h
63.5
62.3
66.8
67.7
70.2
67.4
72 h
63.9
66.6
67.3
66.8
70.6
67.4
Cu Head
Grade (%)
96 h
62.1
66.0
71.2
66.4
Calculated
0.08
0.07
0.07
0.13
0.13
0.13
Initial tests (CN-1 and CN-2) showed gold extractions of 88% and 92% after 72 hours, with
leaching nearly complete in 8 hours. Silver extractions were approximately 78% and 80%,
indicating longer retention benefits silver recovery. Tests CN-3 to CN-6 assessed grind size,
cyanide concentration, and increased leach time to 96 hours.
Gold residue grades for Comp 1 were 0.09 g/t to 0.10 g/t Au, and for Comp 2, 0.19 g/t to 0.21 g/t
Au. Gold extractions were similar at 1 g/L and 2 g/L cyanide, but silver extractions improved with
finer grind size, longer retention, and higher cyanide. The highest gold and silver extractions were
approximately 91.5% and 85% for Comp 1, and 94% and 91% for Comp 2.
Copper residue grades for Comp 1 decreased from 0.28% to 0.017% with higher retention time,
finer grind, and higher cyanide, while Comp 2 ranged from 0.036% to 0.44%. Copper extractions
were similar at 2 g/L cyanide, but the highest extraction (66% for Comp 1, 71.2% for Comp 2)
was at 1 g/L cyanide.
SGS-Lakefield Program (2021-2022)
Thirteen bottle roll tests were completed on a single Abra Zone master composite sample to
evaluate the response of cyanide leaching using standard conditions that focussed on the extraction
of gold, silver and copper. The range of leach conditions included:
•
•
•
•
•
•
•
•
Primary grind k80 size varied between 42 µm and 97 µm;
Pre-aeration time of up to 4 hours (when included);
40% w/w solids;
Lime added to maintain pH at 10.5 to 11;
1 g/L to 3 g/L of NaCN maintained;
Lead nitrate (included for some tests);
Air sparging during leaching; and
Leach retention time of 48 hours to 96 hours.
The test results are presented in Table 13.17.
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TABLE 13.17
2021-2022 ABRA ZONE MASTER COMPOSITE LEACH BOTTLE ROLL TEST RESULTS
Test
No.
CN-1
Cn-2
CN-3
CN-4
CN-5
CN-6
CN-7
CN-8
CN-9
CN-10
CN-11
CN-12
CN-13
Feed
Size
k80
(µm)
NaCN
Added
(g/L)
72
72
73
52
97
74
74
75
68
42
69
68
71
1
2
3
2
2
2
2
2
2
2
2
2
2
Reagent
Cons.
(kg/t)
Au Extraction
(%)
NaCN CaO 8h 24h 48h 72h
4.54
0.62 82 83
86
90
5.03
0.59 89 92
94
94
6.34
0.46 90 93
94
94
5.56
0.59 93 95
96
96
5.01
0.50 90 94
94
96
5.26
0.53 90 93
94
93
5.00
0.97 85 88
90
91
5.15
0.52 92 92
94
93
4.63
0.82 4.47
0.89 4.63
0.87 93
4.72
0.80 92
4.47
0.06 -
96h
91
94
95
95
94
94
92
95
93
93
92
Au
Head
Grade
(g/t)
Calc
1.52
1.55
1.55
1.55
1.54
1.64
1.65
1.58
1.46
1.29
1.54
1.49
1.48
Ag Extraction
(%)
8h
16
38
45
37
38
40
47
52
-
24h
40
51
59
53
51
53
65
65
-
Note: Calc = calculated, h = hours.
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48h
70
65
70
71
65
67
65
78
73
-
72h
79
72
76
80
71
73
81
84
77
-
96h
81
76
79
81
75
77
82
85
81
83
82
Ag
Head
Grade
(g/t)
Calc
138
142
140
140
138
140
141
140
126
131
132
132
129
Cu Extraction
(%)
8h
-
24h
-
48h
79
-
72h
79
-
96h
75
81
81
76
77
78
74
81
80
82
80
Cu
Head
Grade
(%)
Calc
0.19
0.18
0.18
0.19
0.18
0.18
0.18
0.18
0.21
0.21
0.21
0.21
0.21
Eight initial tests (CN-1 to CN-8) were conducted to assess factors such as NaCN addition, grind
size, lead nitrate addition, and aeration effects. CN-8 showed the best leach response with gold
and silver extractions of 95% and 85%, respectively, at a 75 µm grind size, 2.0 g/L NaCN, oxygen
sparging, and a 96-hour retention time. Copper extraction was 81%. Increasing NaCN or finer
grinding showed no significant improvements in gold and silver extraction.
Five additional optimization tests were done, including one to further examine grind size (CN-10),
two for leach retention time (CN-11 to CN-12), and one to test natural NaCN decay after 24 hours
(CN-13). CN-13 yielded gold and silver extractions of 92% and 82%, respectively, with a 71 µm
grind size, 4-hour pre-aeration, and a 96-hour total leach time. Gold and silver residue grades were
higher, and extractions lower compared to CN-8.
Tests CN-9, CN-11, and CN-12 showed no significant difference in gold extraction after 48 hours
versus 96 hours, but silver extraction increased from 73% to 83% with the extra 48 hours.
A bottle roll leach test was conducted on six Abra Zone samples using CN-13 conditions: 75 µm
grind size, 40% solids, pH 10.5 to 11.0, 4 hours pre-aeration, and NaCN decay after 24 hours for
a total of 96 hours. The variability test results are presented in Table 13.18.
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TABLE 13.18
2021-2022 ABRA ZONE VARIABILITY SAMPLE BOTTLE ROLL LEACH TEST RESULTS
Test
No.
CN-15
CN-16
CN-17
CN-18
CN-19
CN-20
Variability
Sample No.
1
2
3
4
5
6
Reagent Cons.
(kg/t)
NaCN
0.87
0.77
1.07
1.26
1.52
2.07
CaO
0.39
0.23
0.26
0.15
0.21
0.24
Au Extraction
(%)
12h
96
91
88
93
90
89
24h
93
89
87
91
88
77
48h
96
90
90
91
92
91
72h
94
92
90
92
90
92
96h
97
96
95
95
94
94
Au
Head
Grade
(g/t)
Calc
0.95
3.01
1.31
1.92
2.38
2.21
Note: Calc = calculated, h = hours.
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Ag Extraction
(%)
12h
62
57
50
47
38
36
24h
67
66
58
58
54
53
48h
72
76
69
71
75
72
72h
74
82
76
77
83
80
96h
83
83
83
82
88
84
Ag
Head
Grade
(g/t)
Calc
59.9
184
231
261
322
448
Cu
Extraction
(%)
Cu Head
Grade
(%)
96h
50
30
45
50
49
58
Calc
0.03
0.03
0.04
0.04
0.06
0.07
The six variability tests achieved gold extractions of 94% to 97% and silver extractions of 82% to
88% after 96 hours. Tailing grades ranged from 0.03 g/t to 0.15 g/t for gold and 10.0 g/t to 72.7
g/t for silver. NaCN consumption was 0.77 kg/t to 2.07 kg/t, lime consumption was 0.15 kg/t to
0.39 kg/t, and copper extraction ranged from 30% to 58%.
A 20 kg bulk leach (BL-1) was completed on the master composite sample for subsequent Merrill
Crowe and SART testing. The bulk leach resulted in gold, silver, and copper extractions of 94%,
80%, and 84%, respectively, with tailing grades of 0.10 g/t for gold, 29.6 g/t for silver, and 0.037%
for copper.
SGS-Lakefield Program (2023)
A leach bottle roll test was completed on each of the three Eagle Zone hole composite samples to
evaluate the response of cyanide leaching using standard conditions that focussed on the extraction
of gold, silver and copper. The bottle roll testing was performed at a target grind size k80 of 75 µm,
40% solids, a pH of 10.5 to 11.0 maintained with lime, 4 hours of pre-aeration with air sparging,
an initial NaCN concentration of 3.0 g/L for the first 24 hours and then allowed to decay naturally
for the remaining 72 hours of the leach and a total leach retention time of 96 hours.
A 30 kg bulk leach was also completed on each hole composite sample (BL-1, BL-2 and BL-3) to
provide sufficient leach solution for the programs subsequent Merrill Crowe and SART testing.
The bulk leach was performed at the same conditions as the bottle roll tests.
The bottle roll and bulk leach test results are presented in Table 13.19.
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Page 167 of 577
TABLE 13.19
2023 EAGLE ZONE HOLE COMPOSITE SAMPLE LEACH TEST RESULTS
Reagent
Cons.
(kg/t)
NaCN
Reagent
Cons.
(kg/t)
CaO
Au
Extraction
(%)
96 h
CN Au
Residue
(g/t)
Au Head
Grade
(g/t)
Calc
Ag
Extraction
(%)
96 h
CN Ag
Residue
(g/t)
Ag Head
Grade
(g/t)
Calc
Cu
Extraction
(%)
96 h
CN Cu
Residue
(g/t)
96 h
Cu Head
Grade
(g/t)
Calc
Sample
Hole
Test
No.
CN
Residue
k80
(µm)
012
CN-13
69
6.12
0.03
94
0.19
3.24
82
20.5
111
88
0.046
0.37
031
CN-14
66
11.5
0.00
95
0.33
6.35
76
33.8
140
80
0.094
0.46
014/035
CN-15
64
7.67
0.02
96
0.16
4.23
92
17.3
208
80
0.057
0.28
012
BL-1
73
5.87
1.03
92
0.21
2.64
86
16.2
118
77
0.052
0.23
031
BL-2
85
12.1
0.00
87
0.80
6.02
77
34.8
150
78
0.11
0.48
014/035
BL-3
87
8.98
0.00
83
0.62
3.53
78
46.8
210
79
0.062
0.29
Note: Calc = calculated.
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Page 168 of 577
The Eagle hole sample bottle roll tests achieved extractions ranging from 94% to 96% for gold
and from 76% to 92% for silver after 96 hours. The residue grades ranged from 0.16 g/t to 0.33 g/t
for gold and from 17.3 g/t to 33.8 g/t for silver. NaCN consumption ranged from 6.12 kg/t to 11.5
kg/t and lime consumption ranged from 0.00 kg/t to 0.03 kg/t. Copper tailings residue and
extractions ranged from 0.046% to 0.057% and 80% to 88%, respectively.
The Eagle bulk leach tests achieved extractions ranging from 83% to 92% for gold and from 77%
to 86% for silver after 96 hours. The residue grades ranged from 0.21 g/t to 0.8 g/t for gold and
from 16.2 g/t to 46.8 g/t for silver. NaCN consumption ranged from 5.87 kg/t to 12.1 kg/t and lime
consumption ranged from 0.00 kg/t to 1.03 kg/t. Copper tailings residue and extractions ranged
from 0.052% to 0.11% and 77% to 79%, respectively. The bulk leach extractions were lower than
those of the corresponding bottle roll tests, however, the bulk leach tests were solely performed to
generate enough sample for downstream flowsheet testwork.
SGS-Lakefield Program (2024)
The FS testwork program consisted of bottle roll leach tests on master and variability composites
from the Eagle, Abra and CC Zones to evaluate their response to WOL cyanide leaching. These
tests assessed gold, silver and copper extractions under various conditions, including particle size,
cyanide concentration, effect of pre-aeration and the use of oxygen and air for leach aeration.
Bottle roll tests were conducted at 40% w/w solids, with target grind sizes (k80) ranging from 50
to 100 µm, 4 hours of pre-aeration using air sparging, pH of 10.5 to 11.0 maintained with lime,
an initial NaCN concentration of 2.0 g/L to 3.0 g/L for the first 24 hours and then allowed to decay
naturally for the remaining 72 hours of the leach and a leach retention time of 96 hours. Eagle
Zone master composite and variability sample test results are presented in Table 13.20 and Table
13.21.
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Page 169 of 577
TABLE 13.20
2024 EAGLE ZONE MASTER COMPOSITE BOTTLE ROLL LEACH TEST RESULTS
Composite
Name
Test ID
Grind
k80
(µm)
Low
Low
Low
Low
Low
Low
Medium
Medium
Medium
Medium
Medium
Medium
High
High
High
High
High
High
CN-1
CN-2
CN-3
CN-4
CN-5
CN-6
CN-7
CN-8
CN-9
CN-10
CN-11
CN-12
CN-39
CN-41
CN-40
CN-42
CN-43
CN-44
72
93
55
72
72
72
69
99
48
66
68
72
71
102
54
72
71
72
NaCN
Consumption
(kg/t)
CaO
Consumption
(kg/t)
Au
Extraction
(%)
8.62
8.32
8.97
8.17
8.65
7.94
7.23
7.05
8.10
6.88
7.37
7.37
10.8
12.6
13.6
8.9
11.2
10.3
0.99
0.91
0.98
0.98
0.95
0.92
0.93
0.84
0.96
0.92
1.38
0.39
0.48
0.42
0.48
0.44
0.46
0.39
90
89
92
90
91
91
93
92
95
93
91
93
93
90
95
92
94
94
Leach
Residue
Au
(g/t)
0.09
0.10
0.08
0.10
0.08
0.08
0.17
0.21
0.12
0.16
0.14
0.16
0.78
1.03
0.61
0.86
0.69
0.71
Calculated
Head
Grade Au
(g/t)
0.91
0.92
0.95
0.99
0.88
0.89
2.39
2.61
2.34
2.44
1.50
2.26
10.4
10.6
11.0
11.0
10.6
10.9
Ag
Extraction
(%)
85
82
85
74
83
85
88
85
91
85
83
89
89
87
91
80
86
90
Leach
Residue
Ag
(g/t)
16
19
15
24
16
14
24
29
18
28
34
22
58
72
50
108
74
50
Calculated
Cu
Head
Extraction
Grade Ag
(%)
(g/t)
104
67
105
69
103
71
94
68
94
71
92
70
193
77
187
80
192
80
193
80
199
81
195
78
536
68
543
68
531
66
534
65
546
69
517
69
Leach
Residue
Cu
(%)
0.12
0.12
0.12
0.13
0.11
0.12
0.07
0.06
0.06
0.06
0.06
0.07
0.14
0.14
0.15
0.15
0.14
0.14
Calculated
Head
Grade Cu
(%)
0.41
0.39
0.41
0.40
0.38
0.40
0.31
0.30
0.30
0.30
0.31
0.31
0.44
0.44
0.45
0.43
0.45
0.46
TABLE 13.21
2024 EAGLE ZONE VARIABILITY SAMPLE BOTTLE ROLL LEACH TEST RESULTS
Variability
Sample ID
Test ID
#1
#1
#2
#2
#3
#3
#4
#4
#5
CN-46
CN-46P
CN-47
CN-47P
CN-48
CN-48P
CN-49
CN-49P
CN-50
CN
NaCN
Residue
Consumption
k80
(kg/t)
(µm)
74
5.34
77
4.44
75
9.12
75
6.74
77
8.46
76
6.30
69
12.9
69
13.1
69
23.5
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CaO
Consumption
(kg/t)
Au
Extraction
(%)
0.12
0.46
0.22
0.39
0.15
0.32
0.30
0.38
0.66
85
88
83
94
89
96
92
93
79
Page 170 of 577
Leach
Residue
Au
(g/t)
0.06
0.05
0.08
0.03
0.12
0.03
0.14
0.10
0.35
Calculated
Head
Grade Au
(g/t)
0.41
0.40
0.45
0.40
1.01
0.84
1.75
1.40
1.65
Ag
Extraction
(%)
82
79
90
86
87
88
81
82
2
Leach
Residue
Ag
(g/t)
4
5
4
6
11
10
19
19
164
Calc.
Head
Grade Ag
(g/t)
22
24
40
44
82
85
100
105
167
Cu
Extraction
(%)
57
57
81
81
66
71
51
50
73
Leach
Residue
Cu
(%)
0.10
0.10
0.04
0.04
0.07
0.06
0.36
0.38
0.37
Calculated
Head
Grade Cu
(%)
0.23
0.23
0.21
0.21
0.21
0.21
0.74
0.76
1.38
TABLE 13.21
2024 EAGLE ZONE VARIABILITY SAMPLE BOTTLE ROLL LEACH TEST RESULTS
Variability
Sample ID
Test ID
#5
#6
#6
#7
#7
CN-50P
CN-51
CN-51P
CN-52
CN-52P
CN
NaCN
Residue
Consumption
k80
(kg/t)
(µm)
69
22.8
67
10.6
69
11.8
71
13.8
70
12.3
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CaO
Consumption
(kg/t)
Au
Extraction
(%)
0.66
0.51
0.51
0.53
0.67
76
92
97
87
99
Page 171 of 577
Leach
Residue
Au
(g/t)
0.38
0.29
0.09
0.77
0.05
Calculated
Head
Grade Au
(g/t)
1.59
3.35
2.69
5.78
5.10
Ag
Extraction
(%)
2
88
87
80
79
Leach
Residue
Ag
(g/t)
167
25
25
29
30
Calc.
Head
Grade Ag
(g/t)
171
202
191
141
146
Cu
Extraction
(%)
57
64
62
53
52
Leach
Residue
Cu
(%)
0.10
0.18
0.20
0.42
0.44
Calculated
Head
Grade Cu
(%)
0.23
0.50
0.53
0.89
0.92
Six bottle roll tests were conducted on each of the Eagle low, medium and high-grade master
composite samples to examine the effect of grind size addition (CN-1 to CN-3, CN-7 to CN-9 and
CN-39 to CN-41), evaluate initial NaCN concentrations (CN-4, CN-10 and CN-42), examine the
effect of air during leaching (CN-5, CN-11 and CN-43), examine the effect of pre-aeration and
leaching utilizing oxygen for sparging (CN-6, CN-12 and CN-44).
The grind size sensitivity tests demonstrated that gold and silver had the highest recoveries, and
the residue tailings grade was lowest, at the finest k80 grind size target of 50 µm. The effect of a
lower initial cyanide concentration of 2.0 mg/L generally measured higher gold and silver leach
residue tailings grades and lower recoveries when compared to the tests at 3.0 mg/L. The tests that
used oxygen for pre-aeration and sparging had slightly higher gold (+1%) and silver (+1%)
recoveries and lower gold (0.01 g/t to 0.07 g/t) and silver (2 g/t to 8 g/t) tailings residue grades
than the tests which solely had pre-aeration.
The copper leach extractions ranged between 66% to 81%.
The leach extractions at a primary grind k80 size of 75 µm were investigated for the seven Eagle
variability composite samples with and without pre-aeration (4 hours) using air sparging and an
initial NaCN concentration of 3.0 g/L for the first 24 hours and then allowed to decay naturally for
the remaining 72 hours of the leach. The tests without pre-aeration all exhibited higher residual
tailings grades for gold (by +0.01 g/t to +0.72 g/t) and negligible silver difference (by -3 g/t to +1
g/t). The difference in cyanide consumption over the tests without the pre-aeration step ranged
between -1.2 kg/t to 2.4 kg/t when compared with the tests with pre-aeration. There was no clear
trend on the benefits of pre-aeration on leach extractions or reagent consumptions.
Abra zone master composite and variability sample leach test results are presented in Table 13.22
and Table 13.23.
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Page 172 of 577
TABLE 13.22
2024 ABRA ZONE MASTER COMPOSITE BOTTLE ROLL LEACH TEST RESULTS
Composite
Name
Test
ID
Low
Low
Low
Low
Low
Low
Medium
Medium
Medium
Medium
Medium
Medium
High
High
High
High
High
High
CN-13-R2
CN-14
CN-15
CN-16
CN-17
CN-18-R1
CN-19
CN-20
CN-21
CN-22
CN-23
CN-24
CN-25
CN-26
CN-27
CN-28
CN-29
CN-30
Grind
Target
k80
(µm)
67
99
53
66
68
68
71
98
58
71
71
70
68
89
54
69
69
69
Leach Calculated
NaCN
CaO
Au
Ag
Residue
Head
Consumption Consumption Extraction
Extraction
Au
Grade Au
(kg/t)
(kg/t)
(%)
(%)
(g/t)
(g/t)
6.1
0.2
94
0.03
0.52
85
5.4
0.3
93
0.04
0.57
84
6.5
0.3
95
0.03
0.55
88
6.09
0.22
94
0.03
0.52
85
5.36
0.28
93
0.04
0.57
85
6.54
0.28
95
0.03
0.55
86
2.48
0.33
95
0.08
1.46
87
2.33
0.32
93
0.10
1.47
87
2.75
0.28
95
0.08
1.42
88
2.08
0.35
94
0.09
1.43
85
3.43
0.49
94
0.08
1.43
88
2.47
0.58
94
0.08
1.38
89
8.43
0.29
94
0.15
2.66
86
8.10
0.22
93
0.19
2.7
85
11.8
0.24
95
0.14
2.66
87
8.23
0.27
94
0.15
2.66
86
8.21
0.50
93
0.19
2.70
85
7.41
0.26
95
0.14
2.66
87
Leach
Residue
Ag
(g/t)
12
13
10
12
12
11
20
21
19
24
19
18
31
35
29
31
35
29
Calculated
Head
Grade Ag
(g/t)
78
83
85
78
78
78
159
158
160
160
163
157
226
230
224
226
230
224
Cu
Extraction
(%)
67
66
67
64
64
70
64
64
57
57
65
57
69.8
70.7
67.4
70
71
67
Leach
Residue
Cu
(%)
0.07
0.07
0.07
0.08
0.08
0.06
0.03
0.03
0.04
0.04
0.03
0.04
0.08
0.08
0.09
0.08
0.08
0.09
Calculated
Head
Grade Cu
(%)
0.21
0.21
0.21
0.22
0.22
0.20
0.08
0.08
0.09
0.09
0.09
0.09
0.26
0.27
0.28
0.26
0.27
0.28
Leach
Residue
Cu
(%)
0.04
0.04
0.03
0.16
0.07
0.19
0.05
0.05
0.04
0.04
Calculated
Head
Grade Cu
(%)
0.10
0.10
0.19
0.31
0.21
0.34
0.11
0.10
0.22
0.22
TABLE 13.23
2024 ABRA ZONE VARIABILITY SAMPLE BOTTLE ROLL LEACH TESTS
Variability
Sample ID
#1
#1
#2
#2
#3
#3
#4
#4
#5
#5
Test
ID
CN-53
CN-53P
CN-54
CN-54P
CN-55
CN-55P
CN-56
CN-56P
CN-57
CN-57P
CN
Residue
P80
(µm)
71
69
77
72
77
73
72
70
71
70
NaCN
CaO
Consumption Consumption
(kg/t)
(kg/t)
7.26
3.44
7.46
5.74
11.7
6.9
4.85
3.61
13.8
9.4
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0.11
0.10
0.13
0.27
0.10
0.23
0.14
0.24
0.12
0.27
Au
Extraction
(%)
91
77
93
80
93
88
95
77
95
87
Page 173 of 577
Leach
Residue
Au
(g/t)
0.03
0.08
0.04
0.11
0.08
0.13
0.08
0.40
0.11
0.27
Calculated
Head
Grade Au
(g/t)
0.29
0.35
0.50
0.52
1.11
1.11
1.70
1.74
1.94
2.02
Ag
Extraction
(%)
86
82
88
87
93
92
95
95
95
94
Leach
Residue
Ag
(g/t)
6
8
11
12
9
10
12
13
20
22
Calculated
Head
Grade Ag
(g/t)
44
44
89
93
126
127
234
238
377
388
Cu
Extraction
(%)
61
61
84
49
68
44
53
52
82
82
Six bottle roll tests were conducted on each Abra low, medium and high-grade master composite
sample to examine the effect of grind size addition (CN-13 to CN-15, CN-19 to CN-21 and CN25 to CN27), evaluate initial NaCN concentrations (CN-16, CN-22 and CN-28), examine the effect
of air during leaching (CN-17, CN-23 and CN-29), examine the effect of pre-aeration and leaching
utilizing oxygen for sparging (CN-18, CN-24 and CN-30).
The grind size sensitivity tests demonstrated that gold and silver recovery typically increased, and
residue tailings grades decreased at finer grind sizes. The lower 2.0 g/L leach cyanide
concentration leach results were comparable to tests with an initial concentration of 3.0 g/L,
however residual gold and silver tailings grades measured higher with reduced gold and silver
recoveries. The tests that used oxygen for pre-aeration and sparging had similar recoveries and
tailings residue grades to the tests which solely had pre-aeration.
The copper leach extractions ranged between 57.5% to 70.7%.
The leach extractions at a primary grind k80 size of 75 µm were investigated for the five Abra
variability samples with and without pre-aeration (4 hours) using air sparging and an initial NaCN
concentration of 3.0 g/L for the first 24 hours and then allowed to decay naturally for the remaining
72 hours of the leach. All tests with pre-aeration exhibited higher residual tailings grades for gold
(by +0.05 g/t to +0.32 g/t) and negligible silver difference (by +1 to +2 g/t). The cyanide
consumption over the tests without pre-aeration increased by 1.2 kg/t to 4.8 kg/t when compared
with the pre-aeration tests.
CC Zone master composite and variability sample leach test results are presented in Table 13.24
and Table 13.25.
P&E Mining Consultants Inc.
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Page 174 of 577
TABLE 13.24
2024 CC MASTER COMPOSITE LEACH BOTTLE ROLL TEST RESULTS
Composite
Name
Test ID
Composite
Composite
Composite
Composite
Composite
Composite
CN-31
CN-32
CN-33
CN-34-R1
CN-35
CN-36
CN
Residue
k80
(µm)
69
100
54
68
70
70
NaCN
Consumption
(kg/t)
CaO
Consumption
(kg/t)
Au
Extraction
(%)
2.00
1.19
2.21
1.02
2.76
1.49
0.75
0.99
0.71
0.94
1.20
1.03
93
91
92
93
93
94
Leach
Residue
Au
(g/t)
0.08
0.10
0.09
0.07
0.08
0.07
Calculated
Head
Grade Au
(g/t)
1.08
1.06
1.17
1.02
1.06
1.20
Ag
Extraction
(%)
90
85
89
88
89
89
Leach
Residue
Ag
(g/t)
5
7
6
6
5
5
Calculated
Head
Grade Ag
(g/t)
49
48
53
52
47
47
Cu
Extraction
(%)
47
29
52
30
48
44
Leach
Residue
Cu
(%)
<0.01
0.02
0.01
0.02
0.01
0.01
Calculated
Head
Grade Cu
(%)
0.02
0.03
0.02
0.03
0.02
0.02
Leach
Residue
Cu
(%)
<0.01
0.01
<0.01
<0.01
Calculated
Head
Grade Cu
(%)
0.02
0.02
0.02
0.02
TABLE 13.25
2024 CC VARIABILITY SAMPLE BOTTLE ROLL LEACH TEST RESULTS
Variability
Sample
Test ID
#1
#1
#2
#2
CN-58
CN-58P
CN-59
CN-59P
CN
Residue
k80
(µm)
70
69
73
73
NaCN
Consumption
(kg/t)
CaO
Consumption
(kg/t)
Au
Extraction
(%)
2.01
2.00
2.80
2.27
0.88
0.89
0.73
0.81
95
94
94
90
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Leach
Residue
Au
(g/t)
0.06
0.06
0.13
0.17
Calculated
Head
Grade Au
(g/t)
1.09
0.99
1.97
1.76
Ag
Extraction
(%)
87
85
87
86
Leach
Residue
Ag
(g/t)
10
11
6
8
Calculated
Head
Grade Ag
(g/t)
74
74
46
57
Cu
Extraction
(%)
34
37
48
49
Six bottle roll tests were conducted (CN-31 to CN-36) on the CC master composite sample to
examine the effect of grind size addition (CN-31 to CN-33), to evaluate initial NaCN
concentrations (CN-34-R1), to examine the effect of air during leaching (CN-35), and to examine
the effect of pre-aeration and leaching utilizing oxygen for sparging (CN-36).
Test CN-31 and CN-33 demonstrated that there was negligible impact of grind size between 69 µm
and 54 µm on gold and silver leach tailings residue grades and recovery. For test CN-34-R1 to
CN-36, test CN-36 showed the best leach response, achieving respective gold and silver leach
tailings residues and leach extractions of 0.07 g/t and 5 g/t and 94% and 89%, respectively, at a
k80 grind size of 70 µm, NaCN addition of 3.0 g/L, oxygen pre-aeration and sparging during the
test and a total leach retention time of 96 hours. However, there appeared to be negligible impact
of varying leach conditions on gold and silver recovery. The copper residue tailings grades and
leach extractions ranged between 0.01% to 0.02% and from 29% to 52%, respectively.
Two CC variability samples were subjected to leach tests without pre-aeration and with 4 hours of
pre-aeration using air. The pre-aeration and no pre-aeration tests were run at a target grind k80 size
of 75 µm, an initial NaCN concentration of 3.0 g/L for the first 24 hours and then allowed to decay
naturally for the remaining 72 hours of the leach, 40% pulp density and target pH of 10.5 to 11.
One test showed comparable results with and without pre-aeration, however, tests CN-59 and CN59P demonstrated a higher residual tailings grade for gold and silver of 0.04 g/t and 2 g/t,
respectively, with 4 hours of pre-aeration.
Results and Conclusions from the 2024 Leaching Program
The major results and conclusions from the 2024 leaching program on the Eagle, Abra and CC
zone composite and variability samples are:
•
Leach tests at target grind k80 sizes of 50, 75, and 100 µm were completed on Eagle,
Abra and CC master composite samples to evaluate the effect of primary grind size on
gold and silver recovery. Gold and silver recovery generally increased for the Eagle
and Abra Zone samples at finer primary grind size. The primary grind k80 of 75 µm
was selected for flowsheet design;
•
The effect of initial leach cyanide concentration on Eagle, Abra and CC Zone master
composite samples was investigated at the selected grind size using 4 hours of preaeration with air sparging and initial NaCN concentrations of 2.0 g/L and 3.0 g/L for
the first 24 hours and then allowed to decay naturally for the remaining 72 hours of the
leach. Gold leach extractions for all zone samples were similar or increased by 1%,
silver leach extraction increased between 3% to 9%, copper extraction varied from 3% to +17% and cyanide consumption increased between 0.2 and 2.0 kg/t at the higher
cyanide concentration. An initial cyanide concentration of 3 g/L was selected for
flowsheet design;
•
The effect of pre-aeration at the selected grind size was investigated for the Eagle, Abra
and CC master composite samples with and without pre-aeration (4 hours) with air
sparging and an initial NaCN concentration of 3.0 g/L for the first 24 hours and then
allowed to decay naturally for the remaining 72 hours of the leach. Pre-aeration tests
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on Eagle and Abra samples showed the following results: Eagle had lower gold residual
tailings after 4 hours of pre-aeration (by -0.01 g/t to -0.72 g/t), with negligible silver
difference. Abra exhibited higher residual tailings grades with pre-aeration for gold (by
+0.05 to +0.16 g/t) and negligible silver difference. The difference in cyanide
consumption without the pre-aeration step varied between -1.2 kg/t and +2.4 kg/t for
Eagle Deposit and increased by +1.4 kg/t to +4.1 kg/t for the Abra Deposit when
compared with tests using pre-aeration. Ultimately, pre-aeration is not recommended
for inclusion on the flowsheet;
•
The effect of air or oxygen for leach aeration at the selected grind size was investigated
for the Eagle, Abra and CC Zone master composite samples using 4 hours of preaeration with air or oxygen sparging and an initial NaCN concentration of 3.0 g/L for
the first 24 hours and then allowed to decay naturally for the remaining 72 hours of the
leach. The results showed comparable leach results with gold and silver residue tailings
grades and recoveries for the Eagle, Abra and CC Zones when comparing the oxygen
and air sparging results. Use of air sparging is recommended for flowsheet design; and
•
In summary, leach conditions selected for flowsheet design and recovery estimation
were a grind k80 size of 75 µm, a 96-hour leach retention time, pH of 10.5-11 and an
initial NaCN concentration of 3.0 g/L for the first 24 hours and then allowed to decay
naturally for the remaining 72 hours of the leach.
13.2.1.7
Merrill Crowe Testwork
The Merrill Crowe process efficiently recovers gold and silver in leach pregnant solution using
zinc dust precipitation. The test programs were primarily to generate sufficient solution for SART
and cyanide detoxification testing and secondarily to confirm gold and silver recovery in the
presence of copper and free cyanide, establish zinc dust and lead nitrate addition requirements.
SGS-Lakefield Program (2021-2022)
The 2021-2022 testwork program included three small scale tests (1 L) performed at 2.5, 5 and 10
times the stoichiometric amount of zinc dust addition to gold and silver, followed by a bulk test
(10 L) at the optimal conditions. The testwork was performed on the pregnant leach solution
(“PLS”) solution from the bulk leach of the program master composite (BL-1). In the tests, lead
nitrate was added at one quarter of the amount of zinc dust added. The results of the small scale
and bulk tests are shown in Table 13.26.
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TABLE 13.26
2021-2022 ABRA ZONE MERRILL CROWE TEST RESULTS
Product
MC-1
Barren
MC-2
Barren
MC-3
Barren
BMC-1
Barren
BL-1 PLS
(Head)
Zn:Ag+Au
Test Size
Stoichiometric
Addition Ratio
L
Au
Assays
(mg/L)
Ag
Cu
Extraction
(%)
Au
Ag
Cu
2.5:1
1.0
0.02
1.30
1,090
98.0
98.3
12.8
5:1
1.0
<0.01
0.07
988
99.0
99.9
21.0
10:1
1.0
<0.01
<0.03
941
99.0
100.0
24.7
10:1
10.0
0.02
1.14
1,090
98.0
98.5
12.8
-
1.0 &
10.0
1.02
75.3
1,250
-
-
-
The three small scale tests demonstrate that gold, silver and copper precipitation recovery increases
with increasing zinc stoichiometric addition ratio, with results showing the gold and silver were
efficiently extracted/precipitated at a zinc stoichiometric ratio of 2.5 or more. The copper
precipitated varied between 12.8% and 24.7%.
The results from the bulk Merrill Crowe test (BMC-1) showed that gold and silver were efficiently
extracted/precipitated out of the PLS solution, with percent extractions of 98.0% for gold and
98.5% for silver. The copper removal was determined to be 12.8%.
SGS-Lakefield (2023)
The 2023 testwork program on the three Eagle Zone composite hole samples included three small
scale tests (I.5 L) performed at 2.5, 5 and 10 times the stoichiometric amount of zinc dust addition
ratio to gold and silver, followed by one bulk test (6 L) at the optimal conditions for each sample.
The testwork was performed on each of the three PLS solutions from the bulk leach of each hole
composite (BL1, BL2 and BL3). In the tests, lead nitrate was added at one quarter of the amount
of zinc dust added. The free cyanide levels in the 012, 031 and 014/035 Eagle hole sample leach
PLS were measured at 1,504 mg/L, 980 mg/L and 1,486 mg/L, respectively, sufficient for the
precipitation stage. The results of the small scale and bulk tests are shown in Table 13.27 and Table
13.28.
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TABLE 13.27
2023 EAGLE ZONE HOLE COMPOSITE MERRILL CROWE SMALL SCALE TEST RESULTS
Sample
ID
012
031
014/035
Product
MC-1A
Barren
MC-1B
Barren
MC-1C
Barren
Feed PLS
(BL-1)
MC-2A
Barren
MC-2B
Barren
Feed PLS
(BL-2)
MC-3A
Barren
MC-3B
Barren
MC-3C
Barren
Feed PLS
(BL-3)
Zn:Au+Ag
Stoichiometric
Addition
Ratio
Test
Size
Assays
(mg/L)
Extraction
(%)
L
Au
Ag
Cu
Au
Ag
Cu
2.5:1
1.5
0.11
2.17
-
93.3
96.8
-
5:1
1.5
<0.05
0.62
-
100
99.1
-
10:1
1.5
<0.05
0.28
-
100
99.6
-
-
-
1.65
68.8
1,180
-
-
-
2.5:1
1.5
0.19
0.49
-
94.6
99.4
-
5:1
1.5
0.08
0.19
-
97.7
99.8
-
-
-
3.53
77.5
2,520
-
-
-
2.5:1
1.5
0.06
0.57
-
97.0
99.5
-
5:1
1.5
<0.05
0.08
-
100
99.9
-
10:1
1.5
<0.05
0.03
-
100
100
-
-
-
1.98
111
1,570
-
-
-
The results from the small-scale tests demonstrate that gold and silver precipitation recovery
increases with increasing zinc stoichiometric addition ratio, with results showing the gold and
silver were efficiently extracted/precipitated at a zinc stoichiometric ratio equal to or greater than
5. The copper precipitation efficiency was not measured during the tests.
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TABLE 13.28
2023 EAGLE ZONE HOLE COMPOSITE MERRILL CROWE BULK TEST RESULTS
Sample
ID
012
031
014/035
Zn:Au+Ag
Stoichiometric
Addition Ratio
Test
Size
L
Au
Assays
(mg/L)
Ag
Cu
Extraction
(%)
Au
Ag
Cu
MC-BL1
Barren
10:1
6.0
0.11
2.17
-
100
98.8
-
Feed PLS
(BL-1)
-
-
1.65
68.8
1,180
-
-
-
MC-BL2
Barren
5:1
6.0
0.19
0.49
-
97.5
99.3
-
Feed PLS
(BL-2)
-
-
3.53
77.5
2,520
-
-
-
MC-BL3
Barren
10:1
6.0
0.06
0.57
-
100
99.8
-
Feed PLS
(BL-3)
-
-
1.98
111
1,570
-
-
-
Product
The results from the bulk Merrill Crowe tests show that gold and silver were efficiently
extracted/precipitated out of the PLS solution, with percent extractions of 97.5% to 100% for gold
and 98.9% to 99.8% for silver. A Zn:Au+Ag stoichiometric ratio of 10:1 was used for Eagle sample
012 and 014/035 and a stoichiometric ratio of 5:1 was used for hole 031. Copper precipitation
efficiency was not measured during the test, but pregnant solution feed levels measured between
1,180 mg/L and 2,520 mg/L.
SGS-Lakefield (2024)
The 2024 precious metal recovery testwork was completed using the PLS of bulk cyanidation tests
for the Eagle, Abra and CC variability composite samples.
The samples were filtered and deaerated prior to testing. Zinc was added at five times the
stoichiometric requirement for gold and silver precipitation to promote recovery of the gold and
silver metals. Lead nitrate was added at a quarter of the zinc addition. A summary of the Merrill
Crowe solution analyses, including analyses of the feed solution and barren solution is presented
in Table 13.29. All tests were performed at a pH of 11.0-11.3, a 5:1 Zn:Au+Ag stoichiometric ratio
and a test volume of either 1.0 L or 10.0 L.
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TABLE 13.29
2024 EAGLE, ABRA, AND CC ZONE MERRILL CROWE SMALL SCALE
AND BULK TESTWORK RESULTS
Composite
ID
Abra Low
Abra
Medium
Abra High
Eagle Low
Eagle
Medium
Eagle High
Cerro
Colorado
Sample
ID
Feed
PLS
MC1-D
MC-BL1
Barren
Feed
PLS
MC2-D
MC-BL2
Barren
Feed
PLS
MC3-D
MC-BL3
Barren
Feed
PLS
MC4-C
MC-BL4
Barren
Feed
PLS
MC5-C
MC-BL5
Barren
Feed
PLS
MC7-C
MC-BL7
Barren
Feed
PLS
MC6-C
MC-BL6
Barren
Assays
Au
Ag*
Cu
Fe
CN(mg/L) (mg/L) (mg/L) (mg/L) (mg/L)
Au
(%)
Extraction
Ag
Cu
(%) (%)
Fe
(%)
0.14
14.4
501.0
52.7
695
-
-
-
-
<0.05
0.25
-
-
-
100.0
98.3
-
-
0.05
0.62
272.00
64.80
-
64.3
95.7
45.7
-23.0
0.20
23.9
93.6
35.6
633
-
-
-
-
<0.05
0.77
-
-
-
100.0
96.8
0.08
1.64
94.7
36.2
-
60.0
93.1
-1.2
-1.7
0.42
33.1
322
76.7
646
-
-
-
-
0.05
0.86
-
-
-
88.1
97.4
0.05
0.35
334
74.7
-
88.1
98.9
-3.7
2.6
0.16
14.5
498
47.1
880
-
-
-
-
<0.05
0.36
-
-
-
100.0
97.5
-
-
0.05
0.28
521
50.3
-
68.8
98.1
-4.6
-6.8
0.17
14.3
498.0
46.4
800
-
-
-
-
0.06
1.53
-
-
-
65.0
89.0
-
-
0.05
0.7
429
60.8
-
70.6
95.1
13.9
-31.0
1.55
71.2
530.0
91.8
570
-
-
-
-
<0.05
0.08-
-
-
-
100.0
99.9
-
-
0.09
0.12
507
56.9
-
94.2
99.8
4.3
38.0
0.16
7.69
19.8
29.4
571
-
-
-
-
0.07
1.57
-
-
-
56.0
80.0
-
-
0.05
0.57
507
56.9
-
68.8
92.6
-3.1
-0.7
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Results showed the gold and silver values in the barren solution were low and align with the values
in the previous testwork programs, however, the corresponding gold and silver extractions at a 5:1
zinc addition stoichiometric ratio are lower. For the Eagle and Abra samples, the bench and bulk
gold and silver extractions ranged from 65.0% to 100% and 60.0% to 94.2% for gold and 89.0%
to 99.9% and 93.1% to 99.8% for silver. The reported gold and silver PLS solution grades for all
samples, apart from the Abra high-grade and Eagle high sample, were considerably lower than the
expected gold and silver nominal and maximum solution grades in the plant. The reported gold
solution grades were 0.4 mg/L and 0.9 mg/L, while the reported silver solution grades were 39.9
mg/L and 61.2 mg/L, respectively. The bench scale tests typically outperformed the bulk scale
tests and feedback from SGS Lakefield was that the bulk tests were predominantly run to generate
sufficient solution for downstream testwork and the lower gold and silver extractions could be due
to zinc powder settling in the reactor.
Approximately 56.0% of the gold and 80.0% of the silver precipitated in the single CC sample
bench test. This improved to 68.8% and 92.6% in the bulk test.
The free CN solution concentrations ranged from 570 mg/L to 880 mg/L which were lower than
in the previous testwork programs and lower than what would be expected in the plant. A higher
stoichiometric ratio of 10:1, which was tested in previous campaigns and improved precipitation
recoveries, was not attempted in the FS campaign indicating there is room for further optimization
of the results.
The copper precipitation was measured from -4.6% to 45.7% in the bulk tests and the Abra low
composite was the only sample with a measured copper extraction higher than 13.9% Cu. The PLS
solution feed copper concentrations ranged between 19.6 mg/L and 530 mg/L over the tests.
Negative Cu extractions are related to sampling and assaying errors with streams that are similar
in Cu concentrations.
Lab scale Merrill Crowe precipitation tests were performed primarily to generate solution for
downstream (SART) testing. The results are not indicative of plant scale precipitation efficiencies.
Test data is not required for design.
13.2.1.8
SART Testwork: Three Campaigns
In 2021, SGS conducted multiple programs on the LRS Deposit including Merrill Crowe
precipitation and cyanide recovery via Sulphidization, Acidification, Recycling, and Thickening
(“SART”).
From a series of small batch tests, it was determined that the CNfree recovery ranged from 120%
to 130%, performed at a sodium hydrosulphide (“NaHS”) stoichiometric addition ranging from
100% to 125% and pH 4.
The 2021 results have been detailed in the previous PEA reports and referenced accordingly in this
Technical Report.
During 2023, SGS completed testwork to investigate the gold and silver recovery potential of the
Eagle (Underground) Zone of the LRS Property. Merrill Crowe precipitation and cyanide recovery
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via SART testing completed on four-hole composites determined that the CNfree recovery ranged
from 99% to 161%, performed at a NaHS stoichiometric addition ranging from 100% to 125% and
pH 4.
In 2024, SGS conducted SART testwork on seven metallurgical composites. From a series of small
SART batch tests, it was determined that the CNfree recovery ranged from 72% to 183%, performed
at a NaHS stoichiometric addition ranging from 125% to 400% at pH 4 for all the testing
completed.
Metallurgy Testwork, Cinco Minas Mining Company: 1910 – 1911
Detailed reports competed on the Cinco Minas material concluded that the “ore contains certain
copper minerals deleterious to cyanide treatment.” Long retention times of greater than 48 hours
and fine grinding produced the best results, but the cyanide consumptions indicated a required
SART into the Los Ricos process flow design.
Metallurgy Testwork, 2021
Commencing in March 2021, SGS conducted testwork on ore samples from the LRS Property to
evaluate and optimize various processes for the recovery of gold and silver including Merrill
Crowe precipitation and cyanide recovery.
To assist with the plant’s water balance, as well as to lower the cyanide reagents costs, Merrill
Crowe barren solution is intended to go through SART to recover/recycle free cyanide back to the
start of the leach. Based on the results of the small batch Merrill Crowe testing, a larger bulk test
was performed using the optimized conditions to produce ample Merrill Crowe barren solution for
SART testing.
SART testing determined that the Weak Acid Dissociable (“WAD”) cyanide associated with
copper and zinc in the Merrill Crowe barrens is fully converted to free cyanide at pH 4 with a
NaHS stoichiometric addition ranging from 100% to 125%. Copper and zinc precipitation
efficiencies were approximately 100%.
Cyanidation testing using recycled SART barren solution produced gold and silver extractions of
95% and 85%, respectively, which were the same as the recoveries achieved with fresh NaCN.
Metallurgy Testwork, 2023
In 2023, SGS conducted testwork on four separate drill hole composites (Hole 012, Hole 031 and
Hole 014/035) from the Eagle Deposit to evaluate various processes for the recovery of gold and
silver including Merrill Crowe precipitation and cyanide recovery via SART.
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SART Testing
Using the Merrill Crowe barren solutions, the regeneration of free cyanide was undertaken by the
SART process. During the testwork, the chemical reactions for the precipitation of copper, zinc
and residual silver with NaHS as Cu2S, ZnS and Ag2S, are as follows:
•
•
•
2 Na2Cu(CN)3 + NaHS + 2.5 H2SO4 - Cu2S + 6 HCN + 2.5 Na2SO4;
Na2Zn(CN)4 + NaHS + 1.5 H2SO4 - ZnS + 4 HCN + 1.5 Na2SO4; and
2 NaAg(CN)2 + NaHS + 1.5 H2SO4 - Ag2S + 4 HCN + 1.5 Na2SO4.
The cyanide is liberated in the process as HCN gas in solution, which is then neutralized with lime
to form calcium cyanide for recycle to the leach plant:
•
2 HCN + Ca(OH)2 - Ca(CN)2 + 2 H2O.
The optimum pH for the SART process is approximately a pH of 4.
A series of three small batch SART tests (1 L) were conducted on each sample to examine the
effect of pH and NaSH addition. Upon completion, a SART barren solution subsample was
removed and submitted for assay. The remaining pulp was filtered to recover the copper sulphide
(“Cu2S”) precipitate and the remaining SART barren solution was re-neutralized to pH 10 with
lime.
The barren solutions products were analyzed for Au, Ag, Cu, Zn, CNT, CNfree, and CNWAD. Acid
and lime (for re-neutralization) consumptions were also determined.
Results determined that the CNfree recovery ranged from 99% to 161%, performed at a NaHS
stoichiometric addition ranging from 100% to 125% and pH 4. Detailed results from the SART
testwork are presented in Table 13.30 through Table 13.32.
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TABLE 13.30
HOLE 012 COMPOSITE SART TEST SUMMARY
Sample ID
NaHS
Stoichiometric
Addition
H2SO4
Addition
SART-1 Barren
SART-2 Barren
SART-3 Barren
MC-BL1 Barren
%
100
110
125
-
g/L
5.70
3.40
4.67
-
Reneutralize
Lime
Addition
g/L
5.55
4.39
4.35
-
Solution Analysis
(mg/L)
CNT
2,600
3,190
3,190
3,390
CNWAD
2,330
2,800
2,800
2,400
CNfree
2,371
2,530
2,596
1,800
Zn
3.99
11.7
50.0
182
CNfree
Recovery
Cu
9.29
8.99
8.53
1,230
%
99
105
108
-
Extraction
Zn
97.8
93.6
72.5
-
Cu
99.2
99.3
99.3
-
Source: SGS (2023)
TABLE 13.31
HOLE 031 COMPOSITE SART TEST SUMMARY
Sample ID
NaHS
Stoichiometric
Addition
H2SO4
Addition
SART-4 Barren
SART-5 Barren
SART-6 Barren
MC-BL2 Barren
%
100
110
125
-
g/L
8.61
6.94
7.14
-
Reneutralize
Lime
Addition
g/L
7.15
6.04
5.94
-
Solution Analysis
(mg/L)
CNT
5,990
4,190
4,590
4,590
CNWAD
4,990
3,590
3,590
3,990
CNfree
2,985
3,663
3,870
1,500
Zn
0.52
0.89
1.36
110
Source: SGS (2023)
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CNfree
Recovery
Cu
46.0
4.20
2.80
2,500
%
124
153
161
-
Extraction
Zn
99.7
99.5
99.3
-
Cu
96.3
99.7
99.8
-
TABLE 13.32
HOLE 014/035 COMPOSITE SART TEST SUMMARY
Sample ID
NaHS
Stoichiometric
Addition
H2SO4
Addition
SART-7 Barren
SART-8 Barren
SART-9 Barren
MC-BL3 Barren
%
100
110
125
-
g/L
7.13
5.68
5.85
-
Reneutralize
Lime
Addition
g/L
5.18
5.16
6.00
-
Solution Analysis
(mg/L)
CNT
4,190
4,190
3,990
4,590
CNWAD
3,390
3,390
3,190
2,800
CNfree
3,178
2,955
3,048
1,800
Zn
1.29
10.0
4.96
227
Source: SGS (2023)
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CNfree
Recovery
Cu
6.81
22.8
16.2
1,650
%
132
123
127
-
Extraction
Zn
99.3
94.5
97.3
-
Cu
99.4
98.1
98.7
-
SART testing determined that the CNfree recovery ranged from 99% to 161%, performed at a
NaHS stoichiometric addition ranging from 100% to 125% and pH 4.
Metallurgy Testwork, 2024
In 2024, SGS conducted SART testwork on seven metallurgical composites from the LRS Deposit.
From a series of small SART batch tests, it was determined that the CNfree recovery ranged from
72% to 183%, performed at a NaHS stoichiometric addition ranging from 125% to 400% at pH 4
for all the testing completed. Table 13.33 presents the 2024 SART testwork results.
TABLE 13.33
SART TEST SUMMARY
Sample ID
Abra Low Grade
Abra Medium Grade
Abra High Grade
Eagle Low Grade
Eagle Medium
Eagle High Grade
Cerro Colorado
NaHS
Stoichiometric
Addition
(%)
125 - 275
125 - 275
125 - 275
125 - 275
125 - 275
125 - 400
125 - 275
CNfree
Recovery
(%)
78 -88
105 - 107
83 - 108
157 - 183
104 - 131
72 - 95
91 - 102
Source: SGS (2025)
SART Testing
Using the Merrill Crowe barren solutions, the regeneration of free cyanide was undertaken by the
SART process. During the testwork, the chemical reactions for the precipitation of copper, zinc
and residual silver with NaHS as Cu2S, ZnS and Ag2S, are as follows:
•
•
•
2 Na2Cu(CN)3 + NaHS + 2.5 H2SO4 - Cu2S + 6 HCN + 2.5 Na2SO4;
Na2Zn(CN)4 + NaHS + 1.5 H2SO4 - ZnS + 4 HCN + 1.5 Na2SO4; and
2 NaAg(CN)2 + NaHS + 1.5 H2SO4 - Ag2S + 4 HCN + 1.5 Na2SO4.
The cyanide is liberated in the process as HCN gas in solution, which is then neutralized with lime
to form calcium cyanide for recycle to the leach plant:
•
2 HCN + Ca(OH)2 - Ca(CN)2 + 2 H2O.
The optimum pH for the SART process is approximately a pH of 4.
A series of small batch SART tests (1 L) were conducted on each sample to examine the effect of
pH and NaHS addition. Upon completion, a SART barren solution subsample was removed and
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submitted for assay. The remaining pulp was filtered to recover the Cu2S precipitate and the
remaining SART barren solution was re-neutralized to pH 10 with lime.
The barren solutions products were analyzed for Au, Ag, Cu, Zn, CNT, CNfree, and CNWAD. Acid
and lime (for re-neutralization) consumptions were also determined.
Abra Low-Grade Composite test results determined that the CNfree recovery ranged from 78%
(SART-1F) performed at a 275% NaHS stoichiometric addition, to 88% (SART-1E) performed at
a 200% NaHS stoichiometric addition. In addition, the zinc, copper and silver precipitation ranged
from 93% to 99%, 68% to 100% and 71% to 95%, respectively, for all the tests.
The sulphuric acid and lime re-neutralization additions ranged from 1.35 g/L to 1.54 g/L and 0.80
g/L to 1.12 g/L, respectively.
Table 13.34 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Abra Low-Grade Composite sample.
Abra Medium-Grade Composite test results determined that the CNfree recovery ranged from
105% (SART-2B and SART-2C) performed at 200% and 275% NaHS stoichiometric addition,
respectively, to 107% (SART-2A) performed at a 125% NaHS stoichiometric addition. In addition,
the zinc, copper and silver precipitation ranged from 97% to 100%, 97% to 100% and 95% to 96%,
respectively, for all the tests.
The sulphuric acid and lime re-neutralization additions ranged from 1.07 g/L to 1.16 g/L and 0.67
g/L to 0.76 g/L, respectively.
Table 13.35 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Abra Medium-Grade Composite sample.
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TABLE 13.34
ABRA LOW-GRADE COMPOSITE SART TEST SUMMARY
Sample ID
SART-1A
SART-1B
SART-1C
SART-1D
SART-1E
SART-1F
Feed
NaHS
Stoichiometric
Addition
%
100
110
125
125
200
275
-
H2SO4
Addition
g/L
1.36
1.38
1.43
1.35
1.46
1.54
-
ReNeutralize
Lime
Addition
g/L
1.12
0.80
1.01
0.93
0.96
1.03
-
Solution Analysis
(mg/L)
CNT
749
649
749
649
649
749
749
CNWAD
599
649
599
549
649
549
649
CNfree
534
526
567
556
570
507
300
Zn
0.47
2.26
1.26
0.50
0.55
0.38
30.7
Cu
65.7
46.1
20.0
33
1.6
0.3
202
CNfree
Recovery
CNS
210
200
190.0
210
190
200
190
Au
0.05
0.05
0.05
0.05
<0.05
<0.05
0.05
Ag
0.04
0.18
<0.03
0.03
<0.05
<0.05
0.62
%
82
81
87
86
88
78
-
Extraction
Zn
98.5
92.6
95.9
98.4
98.2
98.8
-
Cu
67.5
77.2
90.1
83.7
99.2
99.8
-
Ag
93.5
71.0
95.2
95.2
91.9
91.9
-
Source: SGS (2025)
TABLE 13.35
ABRA MEDIUM-GRADE COMPOSITE SART TEST SUMMARY
Sample ID
SART-2A
SART-2B
SART-2C
Feed
NaHS
Stoichiometric
Addition
%
125
200
275
-
H2SO4
Addition
g/L
1.16
1.07
1.12
-
ReNeutralize
Lime
Addition
g/L
0.67
0.67
0.76
-
Solution Analysis
(mg/L)
CNT
524
474
499
599
CNWAD
399
399
374
399
CNfree
428
420
420
379
Zn
0.04
0.38
0.57
19.3
Source: SGS (2025)
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Cu
0.3
0.9
2.4
90.1
CNfree
Recovery
CNS
52
51
56
45
Au
0.08
0.08
0.07
0.08
Ag
0.08
0.08
0.07
1.64
%
107
105
105
-
Extraction
Zn
99.8
98
97
-
Cu
99.7
99
97.4
-
Ag
95.1
95.1
95.7
-
Abra High-Grade Composite test results determined that the CNfree recovery ranged from 83%
(SART-3B) performed at 275% NaHS stoichiometric addition to 108% (SART-3C) performed at
a 275% NaHS stoichiometric addition. In addition, the zinc, copper and silver precipitation ranged
from 96% to 98%, 90% to 99% and 51% to 91%, respectively for all the tests.
The sulphuric acid and lime re-neutralization additions ranged from 1.61 g/L to 2.46 g/L and 0.79
g/L to 1.05 g/L, respectively.
Table 13.36 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Abra High-Grade Composite sample.
Eagle Low-Grade Composite test results determined that the CNfree recovery ranged from 157%
(SART-4A) performed at 125% NaHS stoichiometric addition to 183% (SART-4C) performed at
a 275% NaHS stoichiometric addition. In addition, the zinc precipitation ranged from 95% to 98%
for all the tests, while the copper and silver precipitations were reported as 100% and 91%,
respectively for all the tests.
The sulphuric acid and lime re-neutralization additions ranged from 1.97 g/L to 2.27 g/L and 1.33
g/L to 1.51 g/L, respectively.
Table 13.37 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Eagle Low-Grade Composite sample.
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TABLE 13.36
ABRA HIGH-GRADE COMPOSITE SART TEST SUMMARY
Sample ID
SART-3A
SART-3B
SART-3C
Feed
NaHS
Stoichiometric
Addition
%
125
275
200
-
H2SO4
Addition
g/L
1.16
2.46
1.71
-
ReNeutralize
Lime
Addition
g/L
.79
1.01
1.05
-
Solution Analysis
(mg/L)
CNT
749
724
1050
948
CNWAD
574
549
899
649
CNfree
587
541
699
399
Zn
1.85
1.06
0.90
47.3
Cu
26.4
6.6
2.6
251
CNfree
Recovery
CNS
300
280
280
280
Au
.15
0.14
0.14
0.16
Ag
0.17
0.11
0.03
0.35
%
90
83
108
-
Extraction
Zn
96.1
97.8
98.1
-
Cu
89.5
97.4
99.0
-
Ag
51.4
68.6
91.4
-
Source: SGS (2025)
TABLE 13.37
EAGLE LOW-GRADE COMPOSITE SART TEST SUMMARY
Sample ID
NaHS
Stoichiometric
Addition
H2SO4
Addition
ReNeutralize
Lime
Addition
%
125
200
275
-
g/L
1.97
2.10
2.27
-
g/L
1.51
1.34
1.33
-
SART-4A
SART-4B
SART-4C
Feed
Solution Analysis
(mg/L)
CNT
1600
1100
948
948
CNWAD
649
899
849
649
CNfree
1020
1048
1189
649
Source: SGS (2025)
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Zn
2.22
0.95
1.56
47.3
Cu
1.3
0.4
0.4
251
CNfree
Recovery
CNS
290
290
280
280
Au
0.07
0.06
0.05
0.16
Ag
<0.03
<0.03
<0.03
0.35
%
157
161
183
-
Extraction
Zn
95.3
98.0
96.7
-
Cu
99.5
99.8
99.8
-
Ag
91.4
91.4
91.4
-
Eagle Medium-Grade Composite test results determined that the CNfree recovery ranged from
104% (SART-5A) performed at 125% NaHS stoichiometric addition to 131% (SART-5C)
performed at a 275% NaHS stoichiometric addition. In addition, the zinc, copper and silver
precipitation ranged from 99% to 100%, 97% to 100% and 89% to 96%, respectively, for all the
tests.
The sulphuric acid and lime re-neutralization additions ranged from 1.86 g/L to 2.24 g/L and 1.01
g/L to 1.29 g/L, respectively.
Table 13.38 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Eagle Medium-Grade Composite sample.
Eagle High-Grade Composite test results determined that the CNfree recovery ranged from 72%
(SART-7B) performed at 275% NaHS stoichiometric addition to 95% (SART-7C) performed at a
200% NaHS stoichiometric addition. In addition, the zinc, copper and silver precipitation ranged
from 93% to 99%, 97% to 100% and 94% to 100%, respectively for all the tests.
The sulphuric acid and lime re-neutralization additions ranged from 2.10 g/L to 2.64 g/L and 1.33
g/L to 1.50 g/L, respectively.
Table 13.39 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Eagle High-Grade Composite sample.
Cerro Colorado Composite test results determined that the CNfree recovery ranged from 91%
(SART-6C) performed at 275% NaHS stoichiometric addition to 102% (SART-6A) performed at
a 125% NaHS stoichiometric addition. In addition, the zinc, copper and silver precipitation ranged
from 84% to 93%, 86% to 97% and 0% to 84%, respectively for all the tests.
The sulphuric acid and lime re-neutralization additions ranged from 0.80 g/L to 0.83 g/L and 0.60
g/L to 0.80 g/L, respectively.
Table 13.40 details the results from the SART testing conducted on the Merrill Crowe barren
solution from the Cerro Colorado Composite sample.
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TABLE 13.38
EAGLE MEDIUM-GRADE COMPOSITE SART TEST SUMMARY
Sample ID
SART-5A
SART-5B
SART-5C
Feed
NaHS
Stoichiometric
Addition
%
125
200
275
-
H2SO4
Addition
g/L
1.86
2.10
2.24
-
ReNeutralize
Lime
Addition
g/L
1.01
1.09
1.29
-
Solution Analysis
(mg/L)
CNT
824
861
765
998
CNWAD
749
732
736
799
CNfree
830
921
1045
399
Zn
0.46
0.52
0.14
44.1
Cu
11.8
0.5
0.3
360
CNfree
Recovery
CNS
230
230
240
230
Au
0.24
0.21
0.19
0.27
Ag
0.08
<0.03
<0.03
0.70
%
104
115
131
-
Extraction
Zn
99.0
98.8
99.7
-
Cu
96.7
99.9
99.9
-
Ag
88.6
95.7
95.7
-
Source: SGS (2025)
TABLE 13.39
EAGLE HIGH-GRADE COMPOSITE SART TEST SUMMARY
Sample ID
SART-7A
SART-7B
SART-7C
SART-7D
Feed
NaHS
Stoichiometric
Addition
H2SO4
Addition
ReNeutralize
Lime
Addition
%
125
275
200
400
-
g/L
2.10
2.40
2.22
2.64
-
g/L
1.34
1.33
1.45
1.50
-
Solution Analysis
(mg/L)
CNT
849
1100
749
1050
1300
CNWAD
849
899
649
799
1100
CNfree
885
790
1050
1017
399
Zn
1.68
0.58
0.71
5.26
70.8
Source: SGS (2025)
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Cu
13.3
1.5
0.5
11.2
507
CNfree
Recovery
CNS
340
400
420
350
330
Au
0.03
0.19
0.87
0.77
1.09
Ag
1.06
0.89
<0.02
0.55
18.6
%
80
72
95
92
-
Extraction
Zn
97.6
99.2
99.0
92.6
-
Cu
97.4
99.7
99.9
97.8
-
Ag
94.3
95.2
99.9
97.0
-
TABLE 13.40
CERRO COLORADO COMPOSITE SART TEST SUMMARY
Sample ID
SART-6A
SART-6B
SART-6C
Feed
NaHS
Stoichiometric
Addition
%
125
200
275
-
H2SO4
Addition
g/L
0.80
0.83
0.83
-
ReNeutralize
Lime
Addition
g/L
0.80
0.68
0.60
-
Solution Analysis
(mg/L)
CNT
349
331
343
499
CNWAD
275
268
268
374
CNfree
382
360
341
319
Zn
1.39
0.60
1.04
8.43
Source: SGS (2025)
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Cu
0.8
0.5
2.7
19.3
CNfree
Recovery
CNS
24
27
26
22
Au
0.11
0.11
0.11
0.11
Ag
0.16
0.09
0.62
0.57
%
102
96
91
-
Extraction
Zn
83.5
92.9
87.7
-
Cu
95.8
97.4
86.1
-
Ag
71.9
84.2
-8.8
-
SART Testing – SGS 2024 Conclusion and Recommendations
•
The copper mineralization in the seven metallurgical composites was very reactive with
cyanide, leading to high cyanide consumptions and high concentrations of copper in
the Merrill Crowe barren solutions;
•
Treatment of the Merrill Crowe barren by the SART process will liberate the cyanide
associated with the copper cyanide complex as free cyanide, for recycling to leach, and
will also produce a high-grade copper sulphide product (Cu2S) as a by-product for sale;
and
•
Testing of the SART process on Merrill Crowe barren solutions was very successful.
Free cyanide recoveries of 80% to 157% were achieved with 125% of the
stoichiometric addition of sodium hydrosulphide at pH 4 and 88% to 161% free cyanide
recoveries were achieved at 200% of the stoichiometric addition of sodium
hydrosulphide at pH 4.
Table 13.41 summarizes the results from the SART testing program.
TABLE 13.41
SUMMARY OF LOS RICOS SOUTH SART TEST RESULTS
Item
Anticipated Overall Copper Removal
Anticipated Overall Zinc Removal
Anticipated Overall Silver Removal
Target pH for Copper Removal
Target pH for Neutralization
Free Cyanide Recovery Range – pH 4*
Free Cyanide Recovery Range – pH 4 **
Unit
%
%
%
%
%
Value
90
90
90
4.5
10.5
80-157
88-161
Sources
SGS-18113-06, BQE
SGS-18113-06, BQE
SGS-18113-06, BQE
SGS-18113-06, BQE
SGS-18113-06, BQE
SGS -18113-06
SGS -18113-06
Sources: D.E.N.M. (2025), BQE Water (2025)
*125% Stiochiometric of NaHS
**200% Stiochiometric of NaHS
13.2.1.9
Liquid-Solid Separation Testwork
SGS-Lakefield Program (2023)
Solid/Liquid Separation
As part of the 2023 program, solid-liquid separation testing was performed on the bulk cyanide
leached tailing samples of Hole 012 (BL-1), Hole 031 (BL-2) and Hole 014/035 (BL-3).
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Sample Preparation and Characterization
The samples were received in the form of representative pulp and process water was prepared and
added for pulp dilution as required. A particle size analysis was conducted on all samples, and
subsamples were collected from each sample for solid-content analysis. The dried samples were
submitted for SG determination. The results are shown in Table 13.42.
TABLE 13.42
2023 EAGLE SOLID/LIQUID FEED SAMPLE CHARACTERIZATION
Particle Sizing,
k80 (μm)
73
85
78
Sample ID
BL-1 Final Pulp
BL-2 Final Pulp
BL-3 Final Pulp
SG of Dried
Solids
2.75
2.73
2.71
Pulp pH as
tested
10.9
11.3
11.1
The k80 particle sizes for all three samples ranged between 73-85 μm, while the specific gravity
ranged from 2.71 to 2.75. The pH of all the pulp samples tested were all approximately 11.
Dynamic Thickening
Dynamic thickening tests were performed by SGS using a customized 100 mm benchtop dynamic
thickener. However, apart from sample BL-2, the tests were not completed on BL-1 and BL-3 due
to constant plugging of the underflow discharge. The BL-2 dynamic thickening results, using 70
g/t of Magnafloc 333, are shown in Table 13.43.
TABLE 13.43
SUMMARY OF THICKENING RESULTS BY THICKENER UNIT AREA – BL-2 PULP
Flocc’t
Dosage
(g/t)
Unit
Area
(m2/tpd)
Solids
Loading
(t/m2/h)
70
70
70
70
70
0.12
0.10
0.06
0.06
0.04
0.35
0.42
0.52
0.69
1.04
Net
Rise
Rate
(m/h)
1.71
2.07
2.59
3.45
5.18
Underflow
(% w/w
solids)
Overflow
TSS
(mg/L)
Residence
Time
(h)
59.0
58.1
55.2
54.5
53.3
56
79
220
82
381
0.39
0.33
0.26
0.20
0.13
U/F
Yield
Stress
(Pa)
14
12
7
5
7
Underflow densities ranged between 53.3% to 59% w/w solids and generally increased as the
solids flux loading decreased. The underflow yield stress was low and ranged between 5 Pa and
14 Pa. The overflow clarities ranged between 56 mg/L to 381 mg/L suspended solids and were
generally lower at the lower solids flux loadings.
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Due to the constant plugging of the thickener underflow during the dynamic BL-1 and BL-3
testing, it was recommended to consult with thickener equipment vendors for future testing and
design requirements.
SGS-Lakefield Program (2024)
As part of the FS program, Metso completed sample characterization, flocculant screening and
dynamic settling testwork on Eagle and Abra pre- and post-leach tailings samples. Vacuum and
pressure filtration testwork was completed on multiple samples of the Eagle and Abra Zones
detoxified tailings to establish filter performance and sizing criteria.
The samples used in the 2024 liquid-solid separation testwork are listed in Table 13.44.
TABLE 13.44
2024 EAGLE AND ABRA ZONES SOLID-LIQUID TESTWORK SAMPLE LIST
Sample ID
Eagle Zone Pre-Leach Tailings
Abra Zone Pre-Leach Tailings
Eagle Zone Post-Leach Tailings
Abra Zone Post-Leach Tailings
Eagle Zone Leach Detox Tailings
New Eagle Zone Leach Detox Tailings
Abra Zone Leach Detox Tailings
New Abra Zone Leach Detox Tailings
Old Abra Zone Leach Detox Tailings
Test Program
Static and dynamic thickening;
underflow rheology; particle size distribution
Static and dynamic thickening;
underflow rheology; particle size distribution
Vacuum and Pressure filtration testing, particle
size distribution
The Eagle composite pre-leach, post-leach and detox samples were a 50:50 blend of the Eagle
medium and Eagle low-grade composites. The Abra composite pre-leach, post-leach and detox
samples are an equal blend of the Abra high, medium and low-grade composite samples. The New
Eagle, New Abra and Old Abra Zone leach detox samples were composite samples from the
respective zones.
All detox tailings samples were first bulk leached and then were cyanide detoxified using the
air/SO2 process to reduce the residual CNWAD to <5 mg/L.
Sample Characterization
All pre-leach, post-leach and detox tailings samples were submitted to SGS Lakefield for particle
size analysis and specific gravity (“SG”). The SG was tested using a gas pycnometer and the
particle size distribution (“PSD”) was measured using a combined sieve screen and laser analysis.
The laser analysis was performed using a Malvern Mastersizer 2000 laser analyzer on the screened
material that was less than 38 µm. The SG and PSD results are summarized in Table 13.45.
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TABLE 13.45
2024 METSO OUTOTEC SOLID-LIQUID SAMPLE SG AND PSD RESULTS
Sample
SG
Eagle Zone Pre-Leach Tailings
Abra Zone Pre-Leach Tailings
Eagle Zone Post-Leach Tailings
Abra Zone Post-Leach Tailings
Eagle Zone Detox Tailings
New Eagle Zone Detox Tailings
Abra Zone Detox Tailings
New Abra Zone Leach Detox Tailings
Old Abra Zone Leach Detox Tailings
2.72
2.71
2.68
2.67
2.70
2.71
2.67
2.71
2.68
k10
(µm)
4
4
4
4
4
3
4
3
3
k30
(µm)
20
22
20
20
21
18
21
19
20
k50
(µm)
38
41
38
39
40
39
40
37
40
k80
(µm)
75
83
75
81
79
81
79
75
83
The SG ranged from 2.67 to 2.72 while the k80 size ranged from 75 µm to 83 µm across the
samples.
Dynamic Settling
Dynamic settling tests were performed on the Eagle and Abra pre-leach and post-leach composite
samples to assess settling characteristics and inform thickener design parameters.
Respective Eagle and Abra pre-leach composite sample testwork results are presented in Table
13.46 and Table 13.47. Clear overflow (<100 mg/L) was produced for the majority of the 15 test
runs using less than 20 g/t of flocculant. SNF 923 VHM flocculant was selected as it showed the
best settling performance during flocculant screening.
Flocculant dosage rates of 5, 10 and 20 g/t were tested on the Eagle Zone pre-leach sample at a
solids flux loading of 0.8 t/m2.h, with 10 g/t achieving the highest measured underflow density of
62.8% w/w solids. The solids flux loading rates were then progressively increased from 1.0 t/m2.h
to 1.6 t/m2.h at a flocculant dosage rate of 10 g/t. The underflow densities reduced from 62.5% to
58.9% w/w solids with the increased solids flux loading. The underflow yield stress ranged
between 29 Pa and 87 Pa and generally increased with increased underflow density and increased
flocculant dosage. Underflow yield stresses between 40 Pa to 50 Pa are in the upper range of
centrifugal pump capability and yield stresses less than 40 Pa were achieved at underflow densities
less than 60% w/w solids.
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TABLE 13.46
2024 EAGLE ZONE PRE-LEACH DYNAMIC THICKENING RESULTS
Feed
Test No.
1
2
3
4
5
6
7
Flux
(t/m2.h)
0.80
0.80
0.80
1.00
1.20
1.40
1.60
Flocculant
Liquor Rise
Rate
(m/h)
3.15
3.15
3.15
3.94
4.73
5.51
6.30
Type
Dose
(g/t)
923 VHM
20
10
5
10
10
10
10
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Underflow
Meas.
Yield
Solids
Stress
(% w/w)
(Pa)
61.9
87
62.8
63
62.0
36
62.5
55
59.5
33
58.9
41
58.7
29
Page 199 of 577
Overflow
Solids
(mg/L)
<100
<100
<100
<100
<100
<100
<100
TABLE 13.47
2024 ABRA ZONE PRE-LEACH TAILINGS DYNAMIC THICKENING RESULTS
Test
No.
1
2
3
4
5
6
7
8
Feed
Liquor
Flux
Rise Rate
(t/m2.h)
(m/h)
0.80
2.95
0.80
2.95
0.80
2.95
1.00
3.69
1.20
4.43
1.40
5.17
1.60
5.90
1.80
6.64
Flocculant
Type
Dose
(g/t)
923 VHM
20
10
5
10
10
10
10
10
Underflow
Meas.
Yield
Solids
Stress
(% w/w)
(Pa)
63.2
73
62.0
63
60.5
39
60.2
64
58.9
48
57.4
35
56.8
33
55.7
22
Overflow
Solids
(mg/L)
<100
<100
<100
121
<100
<100
<100
123
Flocculant dosage rates of 5, 10 and 20 g/t were tested on the Abra Zone pre-leach sample at a
solids flux loading of 0.8 t/m2.h, with 20 g/t achieving the highest measured underflow density of
63.2% w/w solids. The solids flux loading rates were then progressively increased from 1.0 t/m2.h
to 1.8 t/m2.h at a flocculant dosage rate of 10 g/t. The underflow densities reduced from 60.2% to
55.7% w/w solids with the increased solids flux loading. The underflow yield stress ranged
between 22 Pa and 73 Pa and generally increased with increased underflow density and increased
flocculant dosage. Underflow yield stresses in the range of centrifugal pump capability were
achieved at underflow densities less than 60% w/w solids.
The respective Eagle and Abra Zone post-leach composite sample testwork results are presented
in Table 13.48 and Table 13.49. Overflow clarities between 116 mg/L and 209 mg/L were observed
for the Eagle sample. The overflow clarities for the Abra sample were higher and ranged between
294 mg/L and 485 mg/L. Higher flocculant dosage rates were required for these samples to target
and maintain acceptable overflow clarities. SNF 920 SH flocculant was used for the Eagle sample
and SNF 905 SH flocculant was used for the Abra sample. Six tests were completed on the Eagle
sample and only 4 tests were completed on the Abra sample due to limited sample availability.
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Page 200 of 577
TABLE 13.48
2024 EAGLE ZONE LEACH TAILINGS DYNAMIC THICKENING RESULTS
Test
No.
Feed
Liquor
Flux
Rise
Rate
(t/m2.h)
(m/h)
Flocculant
Type
Dose
(g/t)
Underflow
Meas.
Yield
Solids
Stress
(% w/w)
(Pa)
Overflow
Solids
(mg/L)
1
0.80
2.70
60
59.0
20
116
2
0.80
2.70
70
59.3
20
133
3
0.80
2.70
80
59.4
15
123
4
0.90
3.03
70
56.9
12
149
5
1.00
3.37
70
58.5
12
161
6
1.20
4.04
70
54.7
24
209
920 SH
TABLE 13.49
ABRA LEACH TAILINGS DYNAMIC THICKENING RESULTS
Test
No.
Feed
Liquor
Flux
Rise
Rate
(t/m2.h)
(m/h)
1
0.80
2.40
2
0.80
2.40
3
0.90
2.70
4
1.00
3.00
Flocculant
Type
905 SH
Dose
(g/t)
Underflow
Meas.
Yield
Solids
Stress
(% w/w)
(Pa)
Overflow
Solids
(mg/L)
80
41.1
108
294
70
56.4
40
447
70
51.6
78
485
70
51.9
50
415
Flocculant dosage rates of 60, 70 and 80 g/t were tested on the Eagle post-leach sample at a solids
flux loading of 0.8 t/m2.h, with 80 g/t achieving the highest measured underflow density of 59.4%
w/w solids. The solids flux loading rates were then progressively increased from 0.9 t/m2.h to 1.2
t/m2.h at a flocculant dosage rate of 70 g/t. The underflow densities reduced from 56.9% to 54.7%
w/w solids with increasing solids flux loading. The underflow yield stress ranged between 12 Pa
and 24 Pa for all tests and generally increased with increased underflow density. Underflow yield
stresses are within centrifugal pump capability.
Flocculant dosage rates of 70 g/t and 80 g/t were tested on the Abra Zone post-leach sample at a
solids flux loading of 0.8 t/m2.h, with 70 g/t achieving the highest measured underflow density of
56.4% w/w solids. The solids flux loading rates were then progressively increased from 0.9 t/m2.h
to 1.0 t/m2.h at a flocculant dosage rate of 70 g/t. The corresponding underflow densities were
51.6% and 51.9% w/w solids with increasing solids flux loading. The underflow yield stress ranged
between 40 Pa and 108 Pa for all tests. Increasing the flocculant to 80 g/t showed the underflow
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density impacted due to over-flocculation and underflow ratholing. Apart from test No. 1 and 3,
the underflow yield stresses are at the upper yield stress range of centrifugal pump capability.
All samples appeared to achieve underflow densities greater than 55% with good overflow clarity
and with relatively low underflow yield stresses, except for the Abra leach tailings sample. The
Abra leach tailings sample underflow densities, apart from one test, were lower than the other
samples and the overflow clarity measured higher and was noticeably murkier. Underflow yield
stresses were high, observed due to ratholing effects of over-flocculation during the testwork.
Additional testwork of the Eagle and Abra post-leach samples is recommended in detailed design
to optimize the flocculant addition rates, underflow densities and overflow clarities.
Rheology
The underflow in each dynamic settling test was measured for rheology using a Thermo Scientific
Haake Viscotester iQ and an “OK600” 4 blade vane. The method measures shear stress versus
time with the peak of this curve equating to yield stress. A constant shear rate of 0.1 sec-1 was used.
All samples were tested at an ambient temperature of 21°C. The yield stress (Pa) versus solids
underflow density for each of the four Abra and Eagle pre-leach and post-leach samples are shown
in Figure 13.3.
FIGURE 13.3
EAGLE AND ABRA ZONE PRE-LEACH AND POST-LEACH UNDERFLOW
YIELD STRESS VERSUS UNDERFLOW SOLIDS DENSITY
Source: Ausenco (2025)
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The plots above show increased yield stress with increased underflow solids density for the Eagle
and Abra pre-leach samples, which is expected. Flocculant dosages less than 20 g/t were required
to achieve acceptable performance for the pre-leach sample tests. The yield stress measurements
for the Eagle post-leach sample measured between 12 Pa and 24 Pa, with no observed relationship
of yield stress increasing with increased solids loading. The Abra post-leach sample shows the
yield stress decreasing with increased underflow solids density, which is not as expected. The
flocculant dosages for the post-leach samples ranged between 60 g/t to 80 g/t, leading to Abra
underflow solids ratholing in the test unit causing low underflow densities and high yield stress.
A higher yield stress requires more pumping power. Typically, slurries with yield stresses above
40 Pa to 50 Pa become difficult to transport using centrifugal pumping. All samples measured less
than 50 Pa at 55% w/w solids, however the high yield stresses at lower underflow densities for the
Abra samples have the potential to affect pump selection and will require further testing and
evaluation in detailed design. Additional dynamic settling testwork is recommended especially for
the post-leach samples to optimise flocculant addition, thickener sizing and confirm pump
selection.
Further rheology testwork is recommended at varying solids densities for the Eagle and Abra
Deposits prior to detailed design of pumps, leach tanks and agitator sizing.
Pressure Filtration
Pressure filtration testing was conducted on two Eagle and three Abra (Abra, New Abra and Old
Abra) Zone composite detox tailings samples. As part of sample preparation, each sample was
leached and then cyanide detoxified using the SO2/air to simulate process conditions. The testwork
was completed using Metso’s Labox 100 bench scale test unit in chamber filter (“CPF”) mode.
While in CPF mode, the chamber was oriented horizontally, two filter cloths were used, and the
test only included the pumping and air-drying cycles.
For all samples, the following filtration parameters were used:
•
•
•
•
•
•
Feed solids content ranged from 54% to 56% w/w solids;
FPC 310 filter cloth used;
40-, 50- and 60-mm chamber widths tested;
Slurry temperature approximately 22°C;
0.6 MPa pumping pressure; and
0.8 MPa drying pressure.
The pressure filtration test results for the two Eagle Zone detoxed tailings samples are shown in
Table 13.50 and Table 13.51.
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TABLE 13.50
2024 EAGLE ZONE DETOX TAILINGS SAMPLE PRESSURE FILTRATION TEST RESULTS
Parameters
Chamber
Thickness
Cycle Time
Cake Thickness
Cake Moisture
Units
Test #1
Test #2
Test #3
Test #4
Test #5
Test #6
mm
40
40
50
50
60
60
min
mm
% w/w
11
40
11.2
7
40
12.5
10
50
12.1
7
50
13.6
10.5
60
12.4
7.5
60
13.8
TABLE 13.51
2024 EAGLE ZONE (NEW EAGLE) DETOX TAILINGS SAMPLE
PRESSURE FILTRATION TEST RESULTS
Parameters
Chamber
Thickness
Cycle Time
Cake Thickness
Cake Moisture
Units
Test #1
Test #2
Test #3
Test #4
Test #5
Test #6
mm
40
40
50
50
60
60
min
mm
% w/w
11.0
40.3
14.8
10.5
40.1
15.2
12.0
50.5
16.1
11.5
50.9
16.3
13.5
60.8
16.5
12.5
61.0
16.9
The pressure filter test results for the Abra Zone (Abra, New Abra and Old Abra) detox tailings
samples are shown in Table 13.52, Table 13.53 and Table 13.54.
TABLE 13.52
2024 ABRA ZONE DETOX TAILINGS SAMPLE PRESSURE FILTRATION TEST RESULTS
Parameters
Chamber
Thickness
Cycle Time
Cake Thickness
Cake Moisture
Units
Test #1
Test #2
Test #3
Test #4
Test #5
Test #6
mm
40
40
50
50
60
60
min
mm
% w/w
12
40
20.6
9
40
19.4
12
50
14.1
8.25
50
15.2
12
60
15.0
9
60
16.3
TABLE 13.53
2024 ABRA ZONE (NEW ABRA) DETOX TAILINGS SAMPLE
PRESSURE FILTRATION TEST RESULTS
Parameters
Chamber
Thickness
Cycle Time
Cake Thickness
Cake Moisture
Units
Test #1
Test #2
Test #3
Test #4
Test #5
Test #6
mm
40
40
50
50
60
60
min
mm
% w/w
11.0
41.7
15.9
10.0
42.5
15.6
12.0
50.6
15.2
11.0
50.4
15.3
13.0
61.1
16.7
12.0
61.0
17.2
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Page 204 of 577
TABLE 13.54
2024 ABRA ZONE (OLD ABRA) DETOX TAILINGS SAMPLE
PRESSURE FILTRATION TEST RESULTS
Parameters
Chamber
Thickness
Cycle Time
Cake Thickness
Cake Moisture
Units
Test #1
Test #2
Test #3
Test #4
Test #5
Test #6
mm
40
40
50
50
60
60
min
mm
% w/w
11.0
41.1
15.5
10.5
41.5
16.1
12.0
51.3
17.3
11.0
49.4
18.4
14.0
60.9
17.4
13.0
60.8
18.8
Conclusions from the testwork were:
•
The Eagle Zone detox tailings sample had the best dewatering properties with filter
cake moistures ranging between 11.2% and 13.8% moisture with filter cycle times
ranging between 7 minutes and 11 minutes;
•
The New Eagle Zone detox tailings sample filter cake moistures ranged between 14.8%
and 16.9% moisture with filter cycle times ranging between 11 minutes and 13.5
minutes;
•
The Abra Zone detox tailings sample filter cake moistures ranged between 14.1% and
20.6% moisture with filter cycle times ranging between 8.25 minutes and 12 minutes.
The filter cake moisture for the 40 mm chamber thickness were unexpectedly high.
These results were confirmed by Metso and likely caused by incorrect set-up and
operation of the test unit. Moisture values comparable to or lower than those achieved
for the 50 mm chamber tests would be expected at a 40 mm chamber thickness;
•
The New Abra Zone detox tailings sample filter cake moistures ranged between 15.2%
and 17.2% moisture with filter cycle times ranging between 10 minutes and 13 minutes;
•
The Old Abra Zone detox tailings sample filter cake moistures ranged between 15.5%
and 18.8% moisture with filter cycle times ranging between 10.5 minutes and 14
minutes;
•
The filter cake moistures were generally lowest in the 40 mm chamber tests, except for
the misleading Abra Zone sample results described above;
•
Overall, the cakes released off the filter cloth easily and crumbled readily. Filter cloths
were easy to wash and there was no cloth blinding noticed during either testing for the
Eagle and Abra Zone samples; and
•
During the test runs, filtrate initially had some ultra-fines pass through the cloth while
the cakes were forming but it became clear approximately 5 seconds later.
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Vacuum Filtration
Vacuum filtration testing was conducted on the same Eagle and Abra Zone composite samples to
compare against the pressure filtration results. Again, each sample was first leached and then
cyanide detoxified with air/SMBS to simulate process conditions. The tests were completed using
Metso’s Larox Pannevis belt (buchner) vacuum technology to examine filtration characteristics.
For all samples, the following testwork parameters were used:
•
•
•
•
Feed solids content ranged between 54% and 56% w/w solids;
Maro S30 or M60 filter cloths;
Vacuum of 0.08 MPa; and
Slurry temperature of 20°C.
The vacuum filtration test results for the two eagle and three Abra Zone detox tailings samples are
summarized in Table 13.55, Table 13.56, Table 13.57, Table 13.58 and Table 13.59. The term
kgD.S/m2.h is kg of dry solids per square metre per hour. The impact of filtration and dry drying
time on cake thickness, cake moisture and filtration rate can be observed.
TABLE 13.55
2024 EAGLE ZONE DETOX TAILINGS VACUUM FILTER RESULTS
Run No.
Units
Filter Cloth
Test #3
Test #4
Test #5
Test #6
Test #7
Test #8
Test #9
Maro
S30
Maro
S30
Maro
S30
Maro
S30
Maro
S30
Maro
S30
Maro
S30
Total
Cycle
Time
s
47
72
101
102
129
161
222
Air Drying
s
30
30
30
60
90
120
180
Cake
Thickness
mm
6.5
12.2
16.4
12.7
11.7
11.9
11.5
Cake
Moisture
% w/w
16.8
18.3
19.2
17.6
14.8
14.3
13.5
Filtration
Rate
kgD.S/
m2.h*
715
952
942
658
503
412
305
*kgD.S/m2.h is kg of dry solids per square metre per hour
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Page 206 of 577
TABLE 13.56
2024 EAGLE ZONE (NEW EAGLE) DETOX TAILINGS VACUUM FILTER RESULTS
Run No.
Units
Test #2
Test #4
Test #5
Test #6
Test #7
Test #8
Test #9
Filter Cloth
Maro
S60
Maro S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Total Cycle
s
Time
101
161
224
193
227
258
311
Air Drying
s
30
30
30
60
90
120
180
Cake
Thickness
mm
7.2
10.3
12.8
10.7
10.8
10.5
10.1
Cake
Moisture
% w/w
20.6
21.0
21.0
19.2
18.2
17.4
17.1
Filtration
Rate
kgD.S/
m2.h*
403
358
311
286
244
221
177
*kgD.S/m2.h is kg of dry solids per square metre per hour
TABLE 13.57
2024 ABRA ZONE DETOX TAILINGS VACUUM FILTER RESULTS
Run No.
Units
Test #2
Test #4
Test #5
Test #6
Test #7
Test #8
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Total Cycle
s
Time
161
314
225
247
291
314
Air Drying
s
30
30
30
60
90
120
Cake
Thickness
mm
11.0
17.0
14.0
13.0
14.0
13.0
Cake
Moisture
% w/w
20.1
21.9
21.1
20.0
19.3
18.8
Filtration
Rate
kgD.S/
m2.h*
355
298
343
295
270
240
Filter Cloth
*kgD.S/m2.h is kg of dry solids per square metre per hour
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Page 207 of 577
TABLE 13.58
2024 ABRA ZONE (NEW ABRA) DETOX TAILINGS VACUUM FILTER RESULTS
Run No.
Units
Filter Cloth
Test #2
Test #4
Test #5
Test #6
Test #7
Test #8
Test #9
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Total
Cycle
Time
s
108
157
217
197
241
262
301
Air Drying
s
30
30
30
60
90
120
180
Cake
Thickness
mm
7.3
10.8
13.4
11.3
10.7
10.8
11.0
Cake
Moisture
% w/w
20.8
20.9
21.2
19.6
18.5
17.9
17.1
Filtration
Rate
kgD.S/
m2.h*
464
373
350
308
264
236
186
*kgD.S/m2.h is kg of dry solids per square metre per hour
TABLE 13.59
2024 ABRA ZONE (OLD ABRA) DETOX TAILINGS VACUUM FILTER RESULTS
Run No.
Units
Filter Cloth
Test #2
Test #4
Test #5
Test #6
Test #7
Test #8
Test #9
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Maro
S60
Total
Cycle
Time
s
200
294
493
362
409
411
476
Air Drying
s
30
30
30
60
90
120
180
Cake
Thickness
mm
7.6
12.0
13.2
11.1
11.2
10.9
10.3
Cake
Moisture
% w/w
22.4
22.3
22.1
21.4
20.9
20.0
19.2
Filtration
Rate
kgD.S/
m2.h*
214
195
158
167
148
146
122
*kgD.S/m2.h is kg of dry solids per square metre per hour
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Conclusions from the testwork were:
•
The Eagle Zone detox tailings sample filter cake moistures, filtration rates and filter
cake thickness ranged between 13.5% and 19.2% moisture, 305 kg/m2.h and 943
kg/m2.h and 6.5 mm and 16.4 mm respectively;
•
The New Eagle Zone detox tailings sample filter cake moistures, filtration rates and
filter cake thickness ranged between 17.1% and 21.0% moisture, 117 kg/m2.h and 403
kg/m2.h and 7.2 mm and 12.8 mm respectively;
•
The Abra Zone detox tailings sample filter cake moistures, filtration rates and filter
cake thickness ranged between 18.8% and 21.9% moisture, 240 kg/m2.h and 355
kg/m2.h and 11.0 mm and 17.0 mm respectively;
•
The New Abra Zone detox tailings sample filter cake moistures, filtration rates and
filter cake thickness ranged between 17.1% and 20.9% moisture, 186 kg/m2.h and 464
kg/m2.h and 7.3 mm and 13.4 mm respectively;
•
The Old Abra Zone detox sample filter cake moistures, filtration rates and filter cake
thickness ranged between 19.2% and 22.4% moisture, 122 kg/m2.h and 214 kg/m2.h
and 7.6 mm and 13.2 mm respectively;
•
The filter cake moistures achieved were generally 4.0% higher than those achieved in
the pressure filtration testwork;
•
Less than 20.0% moisture was achieved on the New Eagle and three Abra samples at
filtration rates between 122 kg/m2.h and 286 kg/m2.h. The Eagle Zone sample
dewatered very well compared to the other samples and achieved a moisture less than
20.0% at a high filtration rate of 952 kg/m2.h.; and
•
The Old Abra sample was the worst dewatering sample with higher moistures and
considerably lower filtration rates than the other four samples.
13.2.1.10
Cyanide Detoxification Testwork
SGS-Lakefield Program (2023-2024)
Batch and continuous cyanide detoxification tests, using the SO2/air process were completed on
Eagle and Abra composite post-leach slurry samples. The aims of the testwork were to confirm
the slurry samples were amenable to the SO2/air cyanide destruction process and to confirm
operating conditions at which a discharge CNWAD level of <3-5 mg/L could be achieved.
The cyanide detox conditions and results for the Eagle and Abra Zones are shown in Table 13.60
and Table 13.61, respectively. All tests were conducted in a temperature controlled 1, 2, 7 or 10 L
continuously fed agitated Makreactor at 55% w/w solids and a pH of 11.0. The Eagle feed solution
measured 956 mg/L CNT, 799 mg/L CNWAD, 459 mg/L Cu and 52.5 mg/L Fe. The Abra feed
solution measured 1,150 mg/L CNT, 799 mg/L CNWAD, 219 mg/L Cu and 53.9 mg/L Fe. Copper
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sulphate was added to some tests as a catalyst and sodium metabisulphite was added as the source
of SO2. Tests 1-6 and 2-9 were continuous bulk tests.
TABLE 13.60
2024 EAGLE ZONE SO2/AIR CYANIDE DETOXIFICATION RESULTS
Test
No.
Test
Test
Retention
Size Duration
Time
(L)
(mins)
(mins)
Feed
1-1
1-2
1-3
1-4
1-5
1-6
1
1
2
2
2
7
109
133
119
120
60
121
109
133
119
120
60
121
CNWAD
SO2
CaO
Cu
Cu
(mg/L)
Equiv
pH
(g/g
(g/g
(mg/L) Product
(g/g
CNWAD) CNWAD)
Solution CNWAD)
10.9
459
799
8.7
154
4.21
1.96
0.00
8.5
1.57
5.86
4.13
0.05
8.5
0.34
5.63
2.69
0.01
8.5
0.2
0.05
5.17
2.40
0.00
8.6
59.9
5.24
2.39
0.00
8.5
<0.1
<0.1
5.07
1.90
0.00
TABLE 13.61
2024 ABRA ZONE SO2/AIR CYANIDE DETOXIFICATION RESULTS
Feed -
-
-
2-1
1
360
117
(CNT) / (CNT) /
CNWAD CNWAD
pH (mg/L)
(mg/L)
Product Product
Solution Solution
(1,150) /
11.0
219
799
8.6
2.02
2-2
1
360
111
8.7
-
21.5
4.82
2.18
0.00
2-3
2
360
122
8.5
-
5.83
5.79
2.42
0.00
2-4
2
360
123
8.5
-
6.78
6.54
2.39
0.00
2-5
2
270
119
8.5
-
5.45
5.25
2.19
0.01
2-6
2
360
121
8.5
-
7.73
5.15
1.76
0.03
2-7
2
360
127
8.5
-
5.56
5.48
2.39
0.07
2-8
2
360
120
8.4
<0.1
<0.1
8.13
3.24
0.07
2-9
10
727
142
8.5
3.4
<0.1
9.13
3.50
0.08
Test
Test
Test
Size Duration
No.
(L)
(mins)
Retention
Time
(mins)
SO2
Equiv
(g/g
CNWAD)
CaO
(g/g
CNWAD)
Cu
(g/g
CNWAD)
4.79
1.71
0.00
The results indicate that for the Eagle Zone, the addition of 5.07 g/g CNWAD of sulphur dioxide
(SO2) in combination with 1.9 g/g CNWAD of lime was effective in reducing the cyanide
concentration to <0.1 mg/L CNWAD at a retention time of 120 minutes (Test 1-6). This treatment
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enabled the sample to reach a cyanide level significantly below the targeted threshold of 3-5 mg/L.
The ratio of 5.07:1 SO2:CNWAD is typical of industry practice but the retention time is 200% of the
typical 60 minutes.
The Abra Zone results showed that that the addition of 8.13 g/g CNWAD of SO2 in combination
with 3.24 g/g CNWAD of lime was effective in reducing the cyanide concentration to <0.1 mg/L
CNWAD at a retention time of 120 minutes (Test 2-8). Copper was added as a catalyst to enhance
the overall efficiency of the detoxification process, with 0.07 g/g CNWAD added. Some of the test
results in the Abra dataset appear misleading and are not as expected as CNWAD levels appear to
increase between tests as reagent dosage strengths increase.
The CNWAD measurements for tests 1-1 to 1-3, 1-5 and 2-1 to 2-7 were determined using picric
acid, as opposed to analytical solution assays used for tests 1-4, 1-6, 2-8 and 2-9. In all cases, the
picric acid CNWAD values were higher that the corresponding solution assay value. In the case of
test 2-9, the analytical assay CNWAD measurement was <0.1 mg/L and the corresponding picric
acid CNWAD measurement was 5.62 mg/L. This suggests that some of the tests, i.e. 2-1, 2-3, 2-5
and 2-8 could potentially have been under the 3-5 mg/L testing target.
The testwork demonstrated the efficiency of the SO2/air detoxification process in achieving the
desired detoxification limit. Some further optimization testing could be completed to optimize the
reagent dosing and test conditions prior to detailed design, especially on the Abra sample.
13.3
METALLURGICAL VARIABILITY
The mine consists of three Zones; Eagle, Abra and CC. Eagle and Abra will be underground mined
and fed exclusively to the plant for the first nine years of production. The 2024 test program
completed by SGS included three master composite grade samples each from the Eagle and Abra
Zones and seven and five variability samples from the Eagle and Abra Zone, respectively. The
program also included one master composite and two variability samples from the CC Zone, which
will be mined as part of a future open pit development and won’t be introduced into the plant feed
until year 10 of production.
The samples selected represent a range of grade, depth and zone parameters from the target mining
areas. The samples were selected from a preliminary FS mine plan available at the commencement
of the study, but it may not fully reflect the subsequent final FS mine plan.
The 2024 Eagle and Abra metallurgical composite, variability and comminution sample locations
relative to the ultimate Eagle and Abra underground mine and open pit shells are shown in the 3D isometric image in Figure 13.4. As shown, there is good spatial sampling coverage for both the
Eagle and Abra underground deposits. The future open pit appears to have limited sampling spatial
coverage as it was not a major focus of the study and this ore source would only be introduced into
the mill feed in year 10 onwards of the mine plan.
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FIGURE 13.4
2024 ABRA AND EAGLE MET COMPOSITE AND VARIABILITY AND COMMINUTION SAMPLE LOCATIONS
(LOOKING NORTH)
Source: GoGold, (2025)
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FIGURE 13.5
CC MASTER COMPOSITE, VARIABILITY AND COMMINUTION SAMPLE LOCATIONS
(ISOMETRIC 3-D IMAGE - LOOKING NORTH)
Source: GoGold, (2025)
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The 2024 CC metallurgical composite, variability and comminution sample locations relative to
the ultimate CC open pit shell is shown in the 3-D isometric image in Figure 13.5, above. As
shown, both CC comminution samples sit within the CC pit development. Roughly half of the
master composite is within the CC pit shell and the variability samples generally fit within the pit
shell. The spatial sampling coverage of the CC open pit is acceptable for this level of study due to
the mine plan timing of the CC Zone being introduced into the mill feed.
Boxplots of the annual mine plan and metallurgical testwork head grades have been developed for
each Zone and are published in Figure 13.6. The metallurgical head grades from all relevant met
testwork programs have been included in the plots. Observations when comparing the box plot
data include:
•
Eagle Zone; the gold, silver and copper underground mine plan quarterly head grades
during the first nine years of mine production range between 0.7 g/t to 3.4 g/t, 54 g/t to
356 g/t and 0.03% to 0.43%, respectively;
o The mine plan gold head grades are within the metallurgical testwork head grades
tested, with the mean of the testwork values falling between the mean and 75th
percentile of the mine plan grades. The max gold head grade tested is three times
the max mine plan head grade;
o The mine plan silver head grades are within the metallurgical testwork head grade
range, however, there were few sample head grades between the 200 g/t to 350 g/t
range which is the upper mill feed head grade range expected in the first two years
of production; and
o Copper levels in the Eagle samples were generally higher than what is expected in
the mine plan. The max Eagle feed copper head grade expected is 0.43%, which
equals the approximate mean of the met testwork dataset. The maximum copper
head grade tested was approximately 1.4% copper.
•
Abra Zone; the gold, silver and copper underground mine plan quarterly head grades
during the first nine years of mine production range between 0.7 g/t to 2.1 g/t, 121 g/t
to 321 g/t and 0.02% to 0.14%, respectively;
o The mine plan gold head grades are within the metallurgical testwork head grades
tested, with the mean of the testwork values aligning with mine plan grade mean.
The max gold head grade tested is one and half times the max mine plan head grade;
o The mine plan silver head grades are within the range of metallurgical testwork
head grades, with the mean of the testwork values slightly higher than the mine plan
grade mean. The max silver head grade tested is one and half times the max mine
plan head grade; and
o Copper levels in the met testwork samples were generally higher than what is
expected in the mine plan. The max feed copper head grade expected is 0.14%,
which equals the approximate mean of the met testwork dataset. The maximum
copper head grade tested was over 1% copper.
•
CC Zone; the gold, silver and copper underground mine plan annual head grades from
year 10 to 13 of mine production range between 0.59 g/t to 0.85 g/t, 26 g/t to 50 g/t and
0.01% to 0.02%, respectively;
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o The mine plan gold head grades are lower than the range of metallurgical testwork
head grades, with the lowest head grade of the testwork samples higher than the
mine plan max grade;
o The mine plan silver head grades are lower than the metallurgical testwork head
grades tested, with the lowest head grade of the met testwork samples being very
close to the maximum head grade expected; and
o Copper levels in the met testwork samples were generally aligned with what is
expected in the mine plan. The testwork and mine plan copper grades were low and
ranged between 0.01% to 0.02%.
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FIGURE 13.6
EAGLE, ABRA AND CC TESTWORK AND MINE PLAN HEAD GRADE BOX PLOTS
Source: Ausenco (2025)
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13.4
DELETERIOUS ELEMENTS
Based on the testwork performed to date, two deleterious elements, mercury and arsenic have been
identified. Arsenic levels were below detection levels in the 2021, 2022 and 2023 testwork
programs, but did reach up to 52 g/t for the 014/035 Eagle hole composite in the 2023 testwork.
Arsenic levels were generally under detection levels for the 2024 testwork samples, but arsenic
did measure up to 112 and 182 g/t in the Eagle samples and 44 g/t in the Abra. The measured
arsenic levels in all CC samples were below detection limits (<30 mg/t). The arsenic demonstrated
no adverse effect on gold and silver recoveries.
Mercury levels were not tested in the previous 2021, 2022 or 2023 testwork campaigns. In the
2024 campaign, mercury was measured between 0.8 g/t to 2.9 g/t and 0.7 g/t to 2.4 g/t for the
respective Eagle and Abra Zone samples. The CC Zone samples had lower measured mercury head
grades ranging between <0.3 g/t to 0.5 g/t. It is anticipated mercury will be extracted to a small
extent during leaching and that which has leached is expected to report directly to the zinc
precipitate. A mercury retort facility will be incorporated into the plant design to recover the
mercury attendant with the zinc precipitate and minimize its deportment to the doré product.
Copper levels were observed to be high in previous testwork programs, which coincided with high
cyanide consumptions. The copper head grades in the annual mine plan range between 0.02% and
0.19%, with the higher grades expected in early production, and typically 65% to 80% of the
copper will leach into solution. The plant design will incorporate both an acid digestion circuit in
the gold room to reduce the copper reporting to the doré and a SART plant to precipitate a copper
sulphide concentrate from the Merrill Crowe barren solution.
There are no other deleterious elements in mineralization which would require special processing
techniques or compromise doré marketability.
13.5
RECOVERY ESTIMATES
Testwork indicated that the preferred flowsheet for the LRS Deposit should include whole ore
cyanide leaching, precious metal recovery using the Merrill Crowe process, cyanide and copper
sulphide recovery using the SART process, dewatering and pressure filtration.
Bottle roll and bulk leach test results for the Eagle and Abra Deposits from the SGS 2021 met
program through to 2024 were analyzed at the target grind k80 size of 75 µm, 2 and 3 g/L cyanide
addition, 4 hours of air pre-aeration (zone composite samples) and no pre-aeration (zone variability
samples) and a leaching time of 96 hours to establish a recovery model for use with the mine
production schedule to provide metal recovery estimates.
Table 13.62 provides the list of samples and the corresponding test ID/conditions used in the
development of the gold, silver and copper recovery model. Composite and variability samples
which had both gold and silver head grades greater than the highest quarterly FS mine plan grades,
or less than lowest quarterly FS mine plan grades, were excluded from the recovery model dataset.
Where either the gold or silver grade of a sample was within the FS mine plan head grade range,
then all gold, silver and copper leach extraction and residue grade data for that sample and test was
used in the dataset. A total of 31 samples,13 from the Eagle Zone and 18 from the Abra Zone,
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were used in the recovery model build-up. Of the 18 Abra samples, 8 were from the 2022 master
composite and variability testing campaign and the remainder was sourced from the 2024 testwork.
Of the 13 Eagle samples, 3 were from the 2023 hole composite campaign with the remainder of
the dataset made up from the 2024 composite and variability testwork. In total, 4 out of 18 Abra
samples and 4 out of 13 Eagle samples were tested without pre-aeration.
Individual Eagle and Abra Zone recovery models were initially developed, however, a combined
Eagle and Abra recovery model was used to estimate the metal recoveries for the following
reasons:
•
Leaching results indicated the Abra and Eagle leaching behaviour was similar, with
and without pre-aeration;
•
A combined recovery model created a single dataset which mitigates potential over and
under recovery estimation with the pre-aeration; and
•
The combined model provides a comparable metal production to the individual models.
Eagle variability sample test CN-50 was excluded from the recovery model dataset even though
the gold and silver head grades were within the expected mine plan range. CN-50 displayed poor
gold (76-79%) and minimal silver extraction (2-3%) and requires further investigation. The CN50 test results were excluded because the elemental head assay analysis of variability sample No.5
sample shows total copper to be 3.24 times the average of the Eagle Zone samples. The copper
grade of this sample was considerably higher than the maximum copper head grade expected in
the mine plan.
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TABLE 13.62
EAGLE AND ABRA BOTTLE ROLL AND BULK LEACH TESTS USED IN THE DEVELOPMENT OF THE RECOVERY MODEL
Zone
Test
Program
Sample Grade
Test ID
Grind
Residue
k80
(µm)
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Abra
Eagle
Eagle
Eagle
Eagle
Eagle
2022
2022
2022
2022
2022
2022
2022
2022
2024
2024
2024
2024
2024
2024
2024
2024
2024
2024
2023
2023
2023
2024
2024
Abra Master Comp
Abra Master Comp
Abra Variability 1
Abra Variability 1
Abra Variability 3
Abra Variability 4
Abra Variability 5
Abra Variability 6
Abra High
Abra Low
Abra Med
Abra High
Abra Low
Abra Medium
Variability 2
Variability 3
Variability 4
Variability 5
012
014/035
031
Eagle Low
Eagle Low
CN-7
BL-1
CN-15
CN-16
CN-17
CN-18
CN-19
CN-20
BCN-5
BCN-3
BCN-4
CN-25
CN-13-R2
CN-19
CN-54
CN-55
CN-56
CN-57
CN-13
CN-15
CN-14
BCN-7
BCN-1
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74
85
67
71
72
62
75
73
81
76
80
68
67
71
77
77
72
71
69
64
66
77
86
Preaeration
(h)
4
4
4
4
4
4
4
4
4
4
4
4
4
4
0
0
0
0
4
4
4
4
4
Initial
NaCN
Conc
(g/L)
2.0
2.0
2.0
2.0
2.0
2.0
2.0
2.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
NaCN
Conc
(h)
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
24
Page 219 of 577
NaCN
Conc
Decay
(h)
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
72
NaCN
Cons.
(kg/t)
CaO
Cons.
(kg/t)
5
4.07
0.87
0.77
1.07
1.26
1.52
2.07
8.12
6.27
4.81
8.43
6.09
2.48
7.46
11.7
4.85
13.8
6.1
7.7
11.5
9.84
7.97
0.97
0.83
0.39
0.23
0.26
0.15
0.21
0.24
0.00
0.00
0.00
0.29
0.22
0.33
0.13
0.10
0.14
0.12
0.0
0.0
0.0
0.00
0.00
TABLE 13.62
EAGLE AND ABRA BOTTLE ROLL AND BULK LEACH TESTS USED IN THE DEVELOPMENT OF THE RECOVERY MODEL
Zone
Test
Program
Sample Grade
Test ID
Grind
Residue
k80
(µm)
Eagle
Eagle
Eagle
Eagle
Eagle
Eagle
Eagle
Eagle
2024
2024
2024
2024
2024
2024
2024
2024
Eagle Medium
Eagle Medium
Low
Medium
Variability 3
Variability 4
Variability 6
Variability 7
BCN-2
BCN-8
CN-1
CN-7
CN-48
CN-49
CN-51
CN-52
81
76
72
69
CN-77
CN-69
67
71
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Preaeration
(h)
4
4
4
4
0
0
0
0
Initial
NaCN
Conc
(g/L)
3.0
3.0
3.0
3.0
3.0
3.0
3.0
3.0
NaCN
Conc
(h)
24
24
24
24
24
24
24
24
Page 220 of 577
NaCN
Conc
Decay
(h)
72
72
72
72
72
72
72
72
NaCN
Cons.
(kg/t)
CaO
Cons.
(kg/t)
7.38
9.09
8.62
7.23
8.46
12.90
10.6
13.8
0.00
0.00
0.99
0.93
0.15
0.3
0.51
0.53
Plots of gold, silver and copper leach extraction residue grades versus calculated head grade were
used to develop residue tailings and head grade regression formulas. The corresponding plots for
the combined Abra and Eagle Zones gold, silver and copper are shown in Figure 13.7, Figure 13.8
and Figure 13.9, respectively.
FIGURE 13.7
EAGLE AND ABRA ZONES LEACH RESIDUE GOLD GRADE VERSUS
CALCULATED HEAD GRADE PLOT
0.90
Leach Residue Grade (g/t Au)
0.80
0.70
0.60
0.50
y = 0.0685x0.9171
R² = 0.6396
0.40
0.30
0.20
0.10
0.00
0.00
1.00
2.00
3.00
4.00
5.00
6.00
Au Calculated Head Grade (g/t)
Source: Ausenco (2025)
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7.00
FIGURE 13.8
EAGLE AND ABRA ZONE LEACH RESIDUE SILVER GRADE VERSUS
CALCULATED HEAD GRADE
80
Leach Residue Grade (g/t Ag)
70
y = 0.111x + 4.4614
R² = 0.5522
60
50
40
30
20
10
0
0
50
100
150
200
250
300
350
400
450
500
Ag Calculated Head Grade (g/t)
Source: Ausenco (2025)
FIGURE 13.9
EAGLE AND ABRA ZONE LEACH RESIDUE COPPER GRADE VERSUS
CALCULATED HEAD GRADE PLOT
0.30
Leach Residue Grade (% Cu)
0.25
y = 0.0229e3.5986x
R² = 0.9213
0.20
0.15
0.10
0.05
0.00
0.00
0.10
0.20
0.30
0.40
0.50
Cu Calculated Head Grade (%)
Source: Ausenco (2025)
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0.60
The following regression formulas for the combined Eagle and Abra Zones were established:
•
Leach extraction for gold: % Au Extraction = [(Au in feed - {0.0685 x Au in
feed^0.9171}) ÷ Au in feed] x 100
•
Leach extraction for silver: % Ag Extraction = [(Ag in feed - {0.111 x Ag in feed +
4.4614}) ÷ Ag in feed] x 100
•
Leach extraction for copper:
o When Cu in feed >0.0376%; % Cu Extraction = [(Cu in feed - {0.0229*exp (3.5986
x Cu in feed)}) ÷ Cu in feed] x 100
o When Cu in feed ≤0.0376% Cu; % Cu Extraction = 30.8%
The CCD solution recoveries below, calculated from METSIM simulations, were applied to the
extraction formulas to determine the individual metal recoveries:
•
•
•
CCD recovery for Au = 99.3%
CCD recovery for Ag = 99.7%
CCD recovery for Cu = 97.3%
Due to the limited CC dataset and the timing of the CC Zone in the mine plan, a fixed CC gold,
silver and copper recovery has been applied. An average metal extraction was calculated using the
master composite CN-31 and variability CN-58, CN-59 and CN-59P tests.
The same CCD solution recovery factors used in the combined Eagle and Abra regressions were
applied to calculate average gold, silver and copper recoveries of 92.5%, 87.6% and 42.6%,
respectively. The CC samples and the corresponding test ID and recoveries used in the CC
recovery estimates are included in Table 13.63.
TABLE 13.63
CC RECOVERY ESTIMATE SAMPLE TEST IDS AND RECOVERIES
Test
Zone
Program
CC
2024
CC
CC
CC
2024
2024
2024
Sample
Grade
Master
Composite
Variability 1
Variability 2
Variability 2
Test ID
Grind
Au
Ag
Cu
Residue
Extraction Extraction Extraction
k80
(%)
(%)
(%)
(µm)
CN-31
69
93
90
47
CN-58
CN-59
CN-59P
70
73
73
95
94
90
87
87
86
34
48
49
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The estimated gold, silver and copper recoveries at the LOM nominal annual mine plan head
grades are shown in Table 13.64.
TABLE 13.64
NOMINAL LOM GOLD, SILVER AND COPPER RECOVERY
ESTIMATES FOR EAGLE AND ABRA ZONES
Product
Au
Ag
Cu
Average
LOM Head
Grade
Extraction
(%)
1.39 g/t
145 g/t
0.1 %
93.3
85.8
67.2
Average
Extraction
(%) at
Average LOM
Head Grade
92.6
85.5
65.4
Based on the selected flowsheet for the plant operation and over the range of grades (variability
samples) tested, the estimated recoveries for gold, silver and copper are 92.6%, 85.5% and 65.4%,
respectively.
Further comparative testwork of Eagle and Abra leach bottle roll testing with and without preaeration is recommended to further support the recovery estimates and confirm the final leaching
configuration of the flowsheet in detailed design.
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14.0
MINERAL RESOURCES ESTIMATE
The Mineral Resource Estimate presented herein for the Los Ricos South Project has been prepared
following the guidelines of the Canadian Securities Administrators’ National Instrument 43-101
and Form 43-101F1, and in conformity with generally accepted “CIM Estimation of Mineral
Resource and Mineral Reserves Best Practices” guidelines. Mineral Resources have been
classified in accordance with the “CIM Standards on Mineral Resources and Reserves: Definition
(2014) and Best Practices (2019)” as adopted by CIM Council.
A Measured Mineral Resource is that part of a Mineral Resource for which
quantity, grade or quality, densities, shape, and physical characteristics are
estimated with confidence sufficient to allow the application of Modifying Factors
to support detailed mine planning and final evaluation of the economic viability of
the deposit. Geological evidence is derived from detailed and reliable exploration,
sampling and testing and is sufficient to confirm geological and grade or quality
continuity between points of observation. A Measured Mineral Resource has a
higher level of confidence than that applying to either an Indicated Mineral
Resource or an Inferred Mineral Resource. It may be converted to a Proven
Mineral Reserve or to a Probable Mineral Reserve.
An Indicated Mineral Resource is that part of a Mineral Resource for which
quantity, grade or quality, densities, shape and physical characteristics are
estimated with sufficient confidence to allow the application of Modifying Factors
in sufficient detail to support mine planning and evaluation of the economic
viability of the deposit. Geological evidence is derived from adequately detailed
and reliable exploration, sampling and testing and is sufficient to assume
geological and grade or quality continuity between points of observation. An
Indicated Mineral Resource has a lower level of confidence than that applying to a
Measured Mineral Resource and may only be converted to a Probable Mineral
Reserve.
An Inferred Mineral Resource is that part of a Mineral Resource for which quantity
and grade or quality are estimated on the basis of limited geological evidence and
sampling. Geological evidence is sufficient to imply but not verify geological and
grade or quality continuity. An Inferred Mineral Resource has a lower level of
confidence than that applying to an Indicated Mineral Resource and must not be
converted to a Mineral Reserve. It is reasonably expected that the majority of
Inferred Mineral Resources could be upgraded to Indicated Mineral Resources
with continued exploration.
Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability.
There is no guarantee that all or any part of the Mineral Resource will be converted into a Mineral
Reserve. Confidence in the estimate of Inferred Mineral Resources is insufficient to allow the
meaningful application of technical and economic parameters or to enable an evaluation of
economic viability worthy of public disclosure. The Authors are not aware of any known
permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that
could materially affect the Mineral Resource Estimate.
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Page 225 of 577
The Authors consider that the block model Mineral Resource Estimate and Mineral Resource
classification represent a reasonable estimation of the global mineral resources for the Los Ricos
South Property with regard to compliance with generally accepted industry standards and
guidelines, the methodology used for estimation, the classification criteria used and the actual
implementation of the methodology in terms of Mineral Resource estimation and reporting. The
Mineral Resources have been estimated in conformity with the requirements of the CIM
“Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines (2019) as
required by the Canadian Securities Administrators’ National Instrument 43-101. Mineral
Resources are not Mineral Reserves and do not have demonstrated economic viability.
Mineral Resource wireframes were developed by GoGold using Seequent Leapfrog GeoTM
software. Mineral Resource estimation and variography was performed by GoGold Seequent
Leapfrog EdgeTM software. Open-pit optimization was performed by the Authors using NPV
SchedulerTM software.
14.1
DATA SUPPLIED
GoGold supplied drill hole data as csv format tables for the Abra, Eagle and Cerro Colorado Veins.
The supplied drill hole tables included collar, survey, assay, lithology and bulk density data. Assay
data included Au g/t, Ag g/t and Cu % grades, and ICP assay grades for the recent drill programs.
The database contains a total of 615 drill hole records. A total of 61 drill holes were excluded from
the Mineral Resource Estimate, due to unreliable historical records, incomplete sampling or
outside the modelled mineralization. A total of 554 drill holes were used for Mineral Resource
modelling (Table 14.1). The drilling extends approximately four km along strike (Figure 14.1 and
Appendix A).
TABLE 14.1
DRILL HOLE SUMMARY
Type
Diamond Drill Hole
Reverse Circulation
Total
Count
518
36
554
Total
(m)
87,241.35
2,853.30
90,094.65
The coordinate system used is WGS84 UTM Zone 13N and elevations are reported as height above
the EGM2008 geoid. GoGold supplied a 1.0 m resolution Digital Terrain Model for the Los Ricos
South Project, and AutoCADTM format wireframes of the historical workings. GoGold geologists
also supplied interpreted wireframes of the identified mineralized structures at Los Ricos South,
as well as a block model for the Los Ricos South Project.
Industry standard validation checks were carried out by the Authors on the supplied databases, and
minor corrections made where necessary. The Authors typically validate a Mineral Resource
database by checking for inconsistencies in naming conventions or analytical units, duplicate
entries, interval, length or distance values less than or equal to zero, blank or zero-value assay
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results, out-of-sequence intervals, intervals or distances greater than the reported drill hole length,
inappropriate collar locations, and missing interval and coordinate fields.
No significant errors were noted with the supplied databases. The Authors consider that the drill
hole database is suitable for Mineral Resource estimation.
14.2
ECONOMIC CONSIDERATIONS
For the Mineral Resource model, GoGold selected the economic parameters listed in Table 14.2.
TABLE 14.2
ECONOMIC PARAMETERS
Item
Au
Ag
Cu
Au Process Plant Recovery
Ag Process Plant Recovery
Cu Process Plant Recovery
Mining Cost
Process + G&A Cost
Calculated AgEq Cut-off
Equivalent Factor Ag:Au
Equivalent Factor Ag:Cu
14.3
Unit
$/oz
$/oz
$/lb
%
%
%
$/t
$/t
AgEq
g/t
%
PitConstrained
$1,850
$23.75
$4.00
95
86
77
$2.10
$28
40
86.0
103.4
Out-of-Pit
$1,850
$23.75
$4.00
95
86
77
$60
$30
130
86.0
103.4
MINERALIZED DOMAINS
Interpreted mineralized wireframes were developed by GoGold geologists for the Eagle and Abra
Veins based on logged drill hole lithology, assay grades and historical records. GoGold identified
continuous zones of mineralization from assay grades equal to or greater than 0.30 g/t AuEq with
observed continuity along strike and down-dip, using a Ag:Au equivalent factor of 75:1. Highgrade domains (e.g., Abra and Eagle Plums) were developed from assay grades equal to or greater
than 3.0 g/t AuEq with observed continuity along strike and down-dip, using a Ag:Au equivalent
factor of 75:1. The selected intervals include lower-grade material where necessary to maintain
continuity between drill holes. Three-dimensional wireframes linking drill hole cross-sections
were subsequently constructed using the LeapfrogTM Radial Basis Function, with hanging wall and
footwall surfaces snapped directly to the selected drill hole intercepts.
A total of eight individual mineralized veins were defined (Figure 14.2 and Appendix B).
The Authors reviewed the mineralized veins and confirms that they are suitable for Mineral
Resource estimation. The mineralized vein wireframes were used to back-tag the assay, bulk
density and composite tables with unique rock codes (Table 14.3).
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FIGURE 14.1
DRILL HOLE PLAN VIEW
Source: P&E (2024)
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TABLE 14.3
MINERALIZED VEINS
Vein
Abra Main
Abra Plum North
Abra Plum South
Eagle Main
Eagle Plum FW
Eagle Plum HW
Cerro Colorado 1
Cerro Colorado 2
FIGURE 14.2
Rock Code
Strike Length
(m)
100
110
120
200
210
220
400
500
1,200
900
200
725
325
550
1,250
650
Average
True Width
(m)
15.3
3.6
3.3
16.1
5.4
6.1
4.7
5.7
MINERALIZED VEINS
Source: P&E (2024)
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14.4
EXPLORATORY DATA ANALYSIS
The average nearest neighbour collar distance for the Los Ricos South Project drilling is 15.3 m.
The average length of the diamond drill holes is 168.42 m, and the average length of the reverse
circulation drill holes is 79.26 m. Summary statistics for the drill hole assay data are listed in Table
14.4.
TABLE 14.4
ASSAY SUMMARY STATISTICS
Assay Element
Count
Diamond Drill Hole
RC
Total
29,750
624
30,374
StdDev
(g/t)
Au (g/t)
0.41
3.36
0.50
1.19
0.41
3.33
Diamond Drill Hole
RC
Total
Diamond Drill Hole
RC
Total
Mean
(g/t)
CoV
Minimum
(g/t)
Maximum
(g/t)
8.23
2.37
8.11
0.001
0.001
0.001
266.00
11.11
266.00
29,750
624
30,374
Ag (g/t)
44.1
561
66.4
212
44.6
556
12.7
3.20
12.5
0.001
0.001
0.001
55,684
3,600
55.684
29,632
41
29,673
Cu (%)
0.05
0.19
0.90
0.30
0.05
0.19
3.67
0.33
3.64
0.00005
0.001
0.00005
7.87
1.00
7.87
Bulk density measurements were determined by GoGold using the water immersion method on
diamond drill core samples. Samples were collected at 20 m intervals downhole and data from all
the major units in the hanging wall, mineralized zones and footwall were logged. The average bulk
density is 2.53 t/m3, with a slightly higher bulk density observed for the higher-grade domains
(Table 14.5).
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TABLE 14.5
SUMMARY OF BULK DENSITY STATISTICS
Item
Waste
Abra Main
Abra Plum North
Abra Plum South
Eagle Main
Eagle Plum FW
Eagle Plum HW
Cerro Colorado 1
Cerro Colorado 2
Total
14.5
Count
2,807
527
169
10
384
199
404
5
10
4,515
Avg. Bulk
Density
(t/m3)
2.46
2.55
2.70
2.54
2.57
2.71
2.79
2.55
2.50
2.53
Median Bulk
Density
(t/m3)
2.50
2.54
2.67
2.55
2.57
2.65
2.69
2.54
2.49
2.54
COMPOSITING
Constrained assay sample lengths within the defined mineralized veins range from 0.15 to 12.0 m,
with an average sample length of 1.07 m and a mode of 1.00 m. A total of 28% of the constrained
assay sample lengths equal 1.00 m (Figure 14.3).
FIGURE 14.3
SAMPLE LENGTHS
Source: P&E (2024)
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Based on the predominance of 1.00 m sample lengths, all constrained assay samples were
composited to this length in order to ensure equal sample support. Length-weighted composites
were calculated within the defined mineralized veins. A small number of un-sampled intervals in
the data were treated as null intervals due to the presence of mining voids, backfill and surface
rock piles. Backfill and surface rock assays were not included in the compositing process.
The compositing process started at the first point of intersection between the drill hole and the
mineralized vein intersected, and halted upon exit from the mineralization. Downhole residual
composites that were less than half the compositing length were discarded in order to not introduce
a short sample bias into the composite sample population. The wireframes that represent the
mineralized veins were used to back-tag a rock code variable into the composite workspace.
The composite data were visually validated against the mineralized wireframes, and extracted for
analysis and grade estimation. Composite summary statistics are listed in Table 14.6.
TABLE 14.6
COMPOSITE SUMMARY STATISTICS
Veins
Count
Mean
Abra Main
Abra Plum
Abra Plum S
Eagle Main
Eagle Plum FW
Eagle Plum HW
4,855
586
142
3,937
522
646
78.38
257.20
450.70
28.70
129.58
533.00
StdDev
Ag (g/t)
174.25
477.60
1,031.50
73.52
184.65
3,033.00
Abra Main
Abra Plum
Abra Plum S
Eagle Main
Eagle Plum FW
Eagle Plum HW
Abra Main
Abra Plum
Abra Plum S
Eagle Main
Eagle Plum FW
Eagle Plum HW
4,855
586
142
3,937
522
646
4,448
577
104
3,911
522
636
CoV
Minimum
Maximum
2.22
1.86
2.29
2.56
1.43
5.69
0.2500
6.1000
0.7000
0.2500
3.8200
5.0000
5,080.30
6,203.30
7,352.30
3,027.01
1,515.20
51,013.00
0.56
2.11
3.08
0.30
3.49
4.42
Au (g/t)
1.76
2.95
5.73
1.32
8.83
11.93
3.17
1.40
1.86
4.33
2.53
2.70
0.0025
0.0110
0.0070
0.0025
0.0090
0.0120
66.40
24.72
45.90
41.29
110.14
159.50
0.05
0.19
0.05
0.07
0.42
0.35
Cu (%)
0.12
0.43
0.06
0.14
0.50
0.40
2.38
2.34
1.26
1.91
1.20
1.15
0.0002
0.0020
0.0005
0.0001
0.0029
0.0006
2.36
4.03
0.30
2.83
3.78
2.74
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14.6
TREATMENT OF EXTREME VALUES
Capping thresholds were determined by the analysis of individual composite
log-probability distributions (Figures 14.4 and 14.5). Composites were capped to the defined
threshold prior to grade estimation (Tables 14.7 to 14.9).
TABLE 14.7
SILVER COMPOSITE CAPPING THRESHOLDS
Veins
Abra Main
Abra Plum
Abra Plum S
Eagle Main
Eagle Plum FW
Eagle Plum HW
Threshold
(g/t)
10,000
2,500
900
500
1,200
7,000
Average
Uncapped Value
(g/t)
78.38
257.17
450.66
28.70
129.58
533.31
Number
Capped
0
6
14
6
4
9
Average
Capped Value
(g/t)
78.38
242.50
241.20
27.25
127.94
340.56
TABLE 14.8
GOLD COMPOSITE CAPPING THRESHOLDS
Veins
Abra Main
Abra Plum
Abra Plum S
Eagle Main
Eagle Plum FW
Eagle Plum HW
Threshold
(g/t)
90
14
16
6
30
40
Average
Uncapped
Value
(g/t)
0.56
2.11
3.07
0.30
3.48
4.42
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Number
Capped
Average
Capped Value
(g/t)
0
8
5
16
5
10
0.56
2.05
2.65
0.26
3.01
3.72
Page 233 of 577
TABLE 14.9
COPPER COMPOSITE CAPPING THRESHOLDS
Veins
Abra Main
Abra Plum
Abra Plum S
Eagle Main
Eagle Plum FW
Eagle Plum HW
Threshold
(%)
2.0
3.0
1.0
1.2
2.5
2.5
Average
Uncapped
Value
(%)
0.05
0.19
0.05
0.07
0.42
0.35
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Number
Capped
Average
Capped Value
(%)
1
3
0
8
4
2
0.05
0.18
0.05
0.07
0.41
0.35
Page 234 of 577
FIGURE 14.4
LOG-PROBABILITY PLOTS FOR ABRA DOMAIN COMPOSITES
Source: GoGold (2024)
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FIGURE 14.5
LOG-PROBABILITY PLOTS FOR EAGLE DOMAIN COMPOSITES
Source: GoGold (2024)
14.7
CONTINUITY ANALYSIS
Three-dimensional continuity analyses (variography) were conducted on the domain-coded
uncapped composite data. In general, an acceptable semi-variogram could only be developed for
the Abra, Eagle and associated high-grade domains, primarily due to the small number of data
points available for the other domains (i.e., Cerro Colorado 1 and 2, and Abra Plum South). The
downhole variogram was viewed at a 1.0 m lag spacing (equivalent to the composite length) to
assess the nugget variance contribution. Standardized spherical models were used to model the
experimental semi-variograms (Figures 14.6 and 14.7). The search ellipsoid orientation can be
seen in Figure 14.8.
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FIGURE 14.6
SEMI-VARIOGRAMS FOR AG COMPOSITES
Source: GoGold (2024)
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FIGURE 14.7
SEMI-VARIOGRAMS FOR AU COMPOSITES
Source: GoGold (2024)
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FIGURE 14.8
LONGITUDINAL PROJECTION SHOWING THE ABRA AND EAGLE VEINS
WITH HISTORICAL WORKINGS AND SEARCH ELLIPSOID ORIENTATIONS
Source: GoGold (2024)
14.8
BLOCK MODEL
GoGold developed a rotated block model in Leapfrog with the block model limits selected in order
to cover the extent of the mineralized structures, the potential open pit dimensions, and to reflect
the general nature of the mineralized zones (Table 14.10). A three-dimensional sub-blocked model,
with 3 m x 3 m x 3 m parent and minimum 1 m x 1 m x 1 m sub-blocks, was utilized to represent
the narrow vein nature of the mineralized domains. The block model consists of estimated Au, Ag
and Cu grades, bulk density, domain codes and classification criteria. Au and Ag equivalent block
grades were subsequently calculated from the estimated Au, Ag and Cu grades.
TABLE 14.10
BLOCK MODEL SETUP
Direction
Minimum X
Minimum Y
Minimum Z
Rotation
Origin
611,600
2,325,200
632
Number of
Parent
Parent
Block
Blocks
Size
768
3m
1,380
3m
346
3m
40° counter-clockwise
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Sub-Block
Minimum
Size
1m
1m
1m
Page 239 of 577
14.9
GRADE ESTIMATION, BULK DENSITY AND MINERAL RESOURCE
CLASSIFICATION
Mineralized domain bulk density values were assigned for the Abra and Eagle Veins based on the
median value. A bulk density of 2.54 t/m3 used for Abra Main, 2.67 t/m3 for Abra Plum, 2.55 t/m3
for Abra Plum South, 2.57 t/m3 used for Eagle Main, 2.65 t/m3 for Eagle Plum FW, and 2.69 t/m3
for Eagle Plum HW.
Block grades for gold and silver were estimated by Inverse Distance Cubed (“ID3”) anisotropic
estimation of capped composites using a minimum of one and a maximum of 12 composites.
The orientation of the search ellipsoid was defined by the modeled variography, observed grade
trends and historical mining. Composite samples were selected within a 120 m x 60 m x 30 m
ellipsoid either plunging -53 >265° for Abra domains or -49 >284° for Eagle domains.
Sample selection was restricted to a maximum of two composite samples from a single drill hole.
Search and grade estimation were constrained by the individual mineralized veins, which define
hard boundaries for grade estimation. Due to insufficient points to generate acceptable semivariograms for the Abra Plum South, the Abra Main Vein semi-variogram was used. AgEq block
model cross-sections and plans can be seen in the Appendix C. Block grades for copper were
estimated by Inverse Distance Squared (“ID2”) isotropic estimation of uncapped composites using
a minimum of one and a maximum of 12 composites.
The parameters used to define the classification limits included spatial analysis, drill hole spacing,
and the observed continuity of the mineralization. Mineral Resources were classified
algorithmically based on the local drill hole spacing within each individual mineralization domain.
For the purposes of classification, historical underground channel sampling was treated as drill
holes; however, underground channel sampling was not used for the grade estimation process. All
blocks within 30 m of three or more drill holes/channel samples were classified as Measured,
blocks within 60 m of two or more drill holes were classified as Indicated, and all additional
estimated blocks were classified as Inferred. Classification block model cross-sections and plans
can be seen in Appendix D.
14.10 MINERAL RESOURCE ESTIMATE
Open-pit Mineral Resources reported herein have been constrained within an optimized pit shell.
The results from the optimized pit shell are used solely for the purpose of reporting Mineral
Resources and include Measured, Indicated and Inferred Mineral Resources. Historical mining has
been depleted from the Abra Vein. Pit-constrained Mineral Resources are reported using a cut-off
of 40 g/t AgEq. The optimized pit shell is shown in Appendix E.
Out-of-pit Mineral Resources have been reported beneath the pit shell; however, they are restricted
to the Eagle and Abra mineralized veins, which exhibit historical continuity and reasonable
potential for extraction by cut-and-fill and longhole mining methods. Out-of-pit Mineral Resources
are reported using a cut-off of 130 g/t AgEq.
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The Mineral Resource has an effective date of January 14, 2025 (Table 14.11).
The sensitivity of the Mineral Resource to changes in cut-off grade was also calculated across a
range of potentially economic AgEq cut-offs (Tables 14.12 and 14.13).
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TABLE 14.11
LOS RICOS SOUTH MINERAL RESOURCE ESTIMATE – PIT-CONSTRAINED AND OUT-OF-PIT (1-9)
Mining Method
Pit-Constrained5
Pit - Cerro Co.6
Out-of-Pit7,8
Total
Classification
Measured
Indicated
Meas & Ind
Inferred
Indicated
Inferred
Measured
Indicated
Meas & Ind
Inferred
Measured
Indicated
Meas & Ind
Inferred
Tonnes
(M)
2.9
2.0
4.9
0.7
0.6
0.3
2.4
3.1
5.5
1.1
5.3
5.6
11.0
2.2
Au
(g/t)
1.13
0.74
0.97
0.66
0.90
0.87
2.00
2.25
2.14
1.48
1.53
1.59
1.56
1.11
Average Grade
Ag
Cu
AuEq
(g/t) (%)
(g/t)
150
0.03
2.96
107
0.03
2.07
133
0.03
2.60
108
0.03
2.00
43
0.01
1.43
34
0.01
1.28
218
0.14
4.79
187
0.26
4.80
201
0.20
4.79
127
0.51
3.61
181
0.08
3.79
144
0.15
3.50
162
0.12
3.64
106
0.27
2.70
AgEq
(g/t)
250
174
219
168
121
109
405
407
406
306
320
296
308
229
Au
(koz)
106
47
153
16
17
10
156
223
379
53
262
287
549
79
Contained Metal
Ag
Cu
AuEq
(koz) (Mlb) (koz)
14,065
1.7
278
6,747
1.4
130
20,812
3.1
408
2,552
0.5
47
787
0.1
26
377
0.1
14
17,025
7.5
373
18,525 17.4
476
35,550 24.9
849
4,544
12.4
130
31,090
9.2
651
26,059 18.9
632
57,150 28.1
1284
7,473
13.0
191
AgEq
(koz)
23,446
10,974
34,420
3,979
2,243
1,217
31,567
40,331
71,898
10,995
55,013
53,549
108,562
16,191
Notes:
1.
Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The estimate of Mineral Resources
may be materially affected by environmental, permitting, legal, title, taxation, socio-political, marketing, or other relevant issues.
2.
The Inferred Mineral Resource in this estimate has a lower level of confidence than that applied to an Indicated Mineral Resource and
must not be converted to a Mineral Reserve. It is reasonably expected that the majority of the Inferred Mineral Resource could be upgraded
to an Indicated Mineral Resource with continued exploration.
3.
The Mineral Resources were estimated in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM
Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve
Definitions and adopted by the CIM Council.
4.
Historically mined areas were depleted from the Mineral Resource model.
5.
The pit constrained cut-off grade of 40 g/t AgEq was derived from US$1,850/oz Au price, US$23.75/oz Ag price, 86% Ag and 95% Au
process recovery, US$28/tonne process and G&A cost. The constraining pit optimization parameters were $2.10/t mineralized material
and waste mining cost, and 45-degree pit slopes.
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6.
7.
8.
9.
Cerro Colorado Resource constrained to open pit mining methods only; Out-of-pit Mineral Resources are restricted to the Eagle and
Abra mineralized veins, which exhibit historical continuity and reasonable potential for extraction by cut and fill and longhole mining
methods.
The out-of-pit cut-off grade of 130 g/t AgEq was derived from US$1,850/oz Au price, US$23.75/oz Ag price, 86% Ag and 95% Au process
recovery, US$30/tonne process and G&A cost, and a $60/tonne mining cost. The out-of-pit Mineral Resource grade blocks were quantified
above the 130 g/t AgEq cut-off, below the constraining pit shell and within the constraining mineralized wireframes. Out–of-Pit Mineral
Resources are restricted to the Eagle and Abra Veins, which exhibit historical continuity and reasonable potential for extraction by cut
and fill and longhole mining methods
AgEq and AuEq were calculated at an Ag/Au ratio of 86:1 for pit constrained and out-of-pit Mineral Resources.
Totals may not sum due to rounding.
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TABLE 14.12
AGEQ CUT-OFF SENSITIVITIES: PIT-CONSTRAINED MINERAL RESOURCE (1-2)
Pit
Constrained
Measured
Indicated
Inferred
Cut-off
(AgEq)
(g/t)
250
150
80
60
40
30
250
150
80
60
40
30
250
150
80
60
40
30
Tonnes
(M)
0.9
1.6
2.3
2.6
2.9
3.0
0.3
0.7
1.4
1.7
2.0
2.1
0.2
0.2
0.6
0.7
0.7
0.7
Au
(g/t)
2.51
1.84
1.39
1.25
1.13
1.09
2.2
1.46
0.94
0.83
0.74
0.71
1.61
1.4
0.73
0.69
0.66
0.66
Average Grade
Ag
Cu
AuEq
(g/t) (%)
(g/t)
311
0.03
6.29
232
0.03
4.65
180
0.03
3.58
164
0.03
3.25
150
0.03
2.96
145
0.03
2.86
301
0.08
5.91
198
0.05
3.9
132
0.04
2.57
118
0.03
2.29
107
0.03
2.07
103
0.03
1.97
263
0.05
4.82
220
0.05
4.1
117
0.03
2.17
111
0.03
2.06
108
0.03
2.00
107
0.03
1.99
AgEq
(g/t)
531
393
302
274
250
241
499
329
217
193
174
167
406
345
183
173
168
167
Au
(koz)
74
92
102
105
106
106
22
32
43
45
47
47
8
10
15
16
16
16
Contained Metal
Ag
Cu
AuEq
(koz) (Mlb) (koz)
9,147
0.6
185
11,552
1.0
232
13,238
1.5
263
13,716
1.7
272
14,065
1.7
278
14,161
2.0
279
2,985
0.5
59
4,391
0.8
86
6,033
1.3
118
6,454
1.1
125
6,747
1.4
130
6,839
1.4
131
1,281
0.2
23
1,500
0.2
28
2,429
0.4
45
2,519
0.5
47
2,552
0.5
47
2,556
0.5
47
Notes:
1.
Assumptions listed under Table 14.11.
2.
Cerro Colorado not included due to lower grade than Eagle and Abra domains.
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AgEq
(koz)
15,595
19,580
22,225
22,934
23,446
23,583
4,952
7,297
9,934
10,558
10,974
11,102
1,979
2,359
3,794
3,931
3,979
3,986
TABLE 14.13
AGEQ CUT-OFF SENSITIVITIES: OUT-OF-PIT MINERAL RESOURCE (1-2)
Out-of-Pit
Measured
Indicated
Inferred
Cut-off
(AgEq)
(g/t)
250
200
150
140
130
120
250
200
150
140
130
120
250
200
150
140
130
120
Tonnes
(M)
1.4
1.7
2.2
2.3
2.4
2.6
1.9
2.3
2.8
2.9
3.1
3.3
0.6
0.8
1.0
1.0
1.1
1.2
Au
(g/t)
2.91
2.57
2.17
2.09
2.00
1.9
3.09
2.75
2.41
2.33
2.25
2.16
2.07
1.84
1.6
1.55
1.48
1.43
Average Grade
Ag
Cu
AuEq
(g/t)
(%)
(g/t)
310
0.19
6.86
274
0.17
6.06
235
0.15
5.17
227
0.15
4.98
218
0.14
4.79
209
0.14
4.58
246
0.3
6.4
222
0.29
5.75
198
0.27
5.11
193
0.26
4.96
187
0.26
4.80
180
0.25
4.62
178
0.57
4.89
156
0.54
4.36
136
0.52
3.86
132
0.52
3.76
127
0.51
3.61
122
0.49
3.49
AgEq
(g/t)
580
512
437
421
405
387
542
487
433
420
407
392
415
370
327
319
306
296
Au
(koz)
127
139
151
154
156
158
191
205
218
220
223
226
40
46
51
52
53
54
Contained Metal
Ag
Cu
AuEq
(koz)
(Mlb)
(koz)
13,586
5.7
300
14,809
6.3
328
16,315
7.1
359
16,658
7.6
366
17,025
7.5
373
17,448
8.0
382
15,184
12.7
395
16,539
14.8
429
17,901
16.7
462
18,207
16.8
469
18,525
17.4
476
18,884
17.9
484
3,410
7.5
94
3,892
9.3
109
4,340
11.4
123
4,428
11.9
126
4,544
12.4
130
4,658
12.8
133
Notes:
1.
Assumptions listed under Table 14.11.
2.
Cerro Colorado not included as it was deemed too low grade for extraction via underground mining methods.
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AgEq
(koz)
25,409
27,690
30,374
30,959
31,567
32,237
33,481
36,371
39,109
39,703
40,331
41,023
7,943
9,250
10,445
10,672
10,995
11,271
14.11 VALIDATION
The block model was validated visually by the inspection of successive cross-sections in order to
confirm that the block models correctly reflect the distribution of high-grade and low-grade values
(see Appendix C).
The average estimated block grades were compared to the average Nearest Neighbor block
estimate at a 0.001 g/t AgEq cut-off (Table 14.14). The results fall within acceptable limits for
grade estimation.
TABLE 14.14
COMPARISON OF ID AND NN AVERAGE BLOCK GRADES
3
Rock Code
ID3
(g/t)
NN
(g/t)
Ratio ID3/NN
51.4
159.0
125.2
28.3
112.4
240.4
119.4
16%
24%
24%
1%
5%
15%
16%
0.32
1.40
1.21
0.26
2.59
3.04
0.32
16%
26%
23%
1%
0%
-1%
16%
Ag
Abra Main
Abra Plum
Abra Plum South
Eagle Main
Eagle Plum FW
Eagle Plum HW
Average
59.9
197.1
154.9
28.7
117.7
275.6
139.0
Au
Abra Main
Abra Plum
Abra Plum South
Eagle Main
Eagle Plum FW
Eagle Plum HW
Average
0.37
1.77
1.49
0.26
2.60
3.03
0.37
The volume estimated was also checked against the reported volume of the individual mineralized
veins. Estimated volumes are based partial block volumes at a 0.001 g/t AgEq cut-off (Table
14.15). The results fall within acceptable limits for grade estimation.
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TABLE 14.15
VOLUME COMPARISON
Veins
Rock Code
Abra Main
Abra Plum
Abra Plum South
Eagle Main
Eagle Plum FW
Eagle Plum HW
100
110
120
200
210
220
Wireframe
Volume
(k m3)
8,357
932
121
4,275
417
536
Estimate
Volume
(k m3)
8,356
932
121
4,274
417
536
A check for local estimation bias was completed by plotting swath plots of the estimated ID3 block
grade and the Nearest Neighbour grade. The results demonstrate a reasonable level of smoothing
for the ID3 estimate. The results fall within acceptable limits for linear estimation (Figures 14.9
and 14.10).
An additional validation check was completed by comparing the average grade of the composites
in a block to the associated model block grade estimate within the aforementioned swath plots.
The results fall within acceptable limits for grade estimation.
FIGURE 14.9
SWATH PLOTS -ABRA
Abra Ag
200.0000
180.0000
160.0000
Mean (g/t)
140.0000
120.0000
100.0000
80.0000
60.0000
40.0000
20.0000
0.0000
0
200
400
600
800
Strike Distance (m)
BM_ag_1
1000
1200
BM_agnn_1
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Abra Au
1.0000
0.9000
0.8000
Mean (g/t)
0.7000
0.6000
0.5000
0.4000
0.3000
0.2000
0.1000
0.0000
0
200
400
600
800
Strike Distance (m)
BM_au_1
1000
1200
BM_aunn_1
Source: P&E (2024)
FIGURE 14.10
SWATH PLOTS – EAGLE
Eagle Au
0.6000
Mean (g/t)
0.5000
0.4000
0.3000
0.2000
0.1000
0.0000
0
100
200
300
400
500
Strike Distance (m)
BM_au_1
600
700
800
BM_aunn_1
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Eagle Ag
90.0000
80.0000
70.0000
Mean (g/t)
60.0000
50.0000
40.0000
30.0000
20.0000
10.0000
0.0000
0
100
200
300
400
500
Strike Distance (m)
BM_ag_1
600
700
800
BM_agnn_1
Source: P&E (2024)
14.12 CERRO COLORADO
The Mineral Resource Estimate presented herein for the Cerro Colorado Deposit has been prepared
following the guidelines of the Canadian Securities Administrators’ National Instrument 43-101
and Form 43-101F1, and in conformity with generally accepted “CIM Estimation of Mineral
Resource and Mineral Reserves Best Practices” guidelines. Mineral Resources have been
classified in accordance with the “CIM Standards on Mineral Resources and Reserves: Definition
(2014) and Best Practices (2019)” as adopted by CIM Council. The effective date of this Mineral
Resource Estimate is January 14, 2025.
Wireframe modelling utilized Seequent Leapfrog GeoTM software. Mineral Resource estimation
was performed using GEOVIA GEMSTM software program. Variography was performed using
Snowden SupervisorTM. Open-pit optimization was performed using the NPV SchedulerTM
software.
14.12.1
Data Supplied
The Authors developed the Cerro Colorado Mineral Resource models from information supplied
by GoGold. The supplied drill hole tables included collar, survey, assay, lithology and bulk density
data. Assay data included Au g/t, Ag g/t and Cu % grades. A total of 40 drill holes were available
for Mineral Resource modelling for Cerro Colorado (Table 14.16). The drilling extends
approximately 1,200 m along strike.
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TABLE 14.16
DRILL HOLE SUMMARY
Type
Count
Diamond Drill Hole
Reverse Circulation
Total
38
6
44
Total
(m)
3,210.4
443.0
3,653.4.0
The coordinate system used is WGS84 UTM Zone 13N and elevations are reported as height above
the EGM2008 geoid.
GoGold supplied a 1.0 m resolution Digital Terrain Model for the Los Ricos South Project, and
AutoCADTM format wireframes of the historical workings. GoGold geologists also supplied
interpreted wireframes of the identified mineralized structures.
Industry standard validation checks were carried out on the supplied databases, and minor
corrections made where necessary. The Authors typically validate a Mineral Resource database by
checking for inconsistencies in naming conventions or analytical units, duplicate entries, interval,
length or distance values less than or equal to zero, blank or zero-value assay results, out-ofsequence intervals, intervals or distances greater than the reported drill hole length, inappropriate
collar locations, and missing interval and coordinate fields.
No significant errors were noted with the supplied databases. The Authors consider that the Cerro
Colorado drill hole database is suitable for Mineral Resource estimation. The drill hole data were
imported into a GEMSTM format MS-Access database.
14.12.2
Economic Considerations
Based on knowledge of similar projects, review of available historical data, and consideration of
potential mining scenarios, the economic parameters listed in Table 14.17 were deemed
appropriate for the Cerro Colorado Mineral Resource Estimate.
TABLE 14.17
ECONOMIC PARAMETERS
Item
Au
Ag
Au Process Plant Recovery
Ag Process Plant Recovery
Mining Cost Pit
Process + G&A Cost
Open Pit Cut-off
Unit
$/oz
$/oz
%
%
$/t
$/t
AgEq g/t
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Cost
(US$)
1,850
23.75
95
86
2.10
28.00
40
Page 250 of 577
TABLE 14.17
ECONOMIC PARAMETERS
14.12.3
Item
Unit
Equivalent Factor
Equivalent Factor
Ag/Cu
Ag/Au
Cost
(US$)
103.4
86.0
Mineralization Domains
Interpreted mineralization wireframes were developed by GoGold geologists based on logged drill
hole lithology, assay grades and historical records. The Authors identified continuous zones of
mineralization within the GoGold-supplied wireframes from assay grades equal to or greater than
400 g/t AgEq with observed continuity along strike and down-dip. The selected intervals include
lower-grade material where necessary to maintain continuity between drill holes. Threedimensional wireframes linking drill hole cross-sections were subsequently constructed using the
LeapfrogTM Radial Basis Function, with hanging wall and footwall surfaces snapped directly to
the selected drill hole intercepts.
Two individual mineralized veins were defined for the Cerro Colorado. The mineralized vein
wireframes were used to back-tag the assay, bulk density and composite tables with unique rock
codes (Table 14.18).
TABLE 14.18
MINERALIZED VEINS
14.12.4
Veins
Rock Code
Cerro Colorado 1
Cerro Colorado 2
400
500
Strike
Length
(m)
1,250
650
Average
True Width
(m)
4.7
5.7
Exploratory Data Analysis
The average nearest neighbour collar distance for Cerro Colorado drilling is 27.6 m. The average
length of the diamond drill holes is 72.4 m, and the average length of the reverse circulation drill
holes is 66.4 m. Summary statistics for the assay data are listed in Table 14.19.
TABLE 14.19
ASSAY SUMMARY STATISTICS
Element
Ag (g/t)
Au (g/t)
Cu (%)
Count
1,263
1,263
1,330
Average
10.6
0.14
0.006
StdDev
52.5
0.55
0.01
CoV
4.95
3.79
1.96
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Minimum
0.15
0.0025
0.0002
Maximum
1,189.0
7.37
0.236
Page 251 of 577
Bulk density measurements were determined by GoGold using the water immersion method on
diamond drill core samples. A total of seven measurements were taken for the Cerro Colorado
Deposits. The average bulk density was 2.49 t/m3, and the median bulk density was 2.50 t/m3.
14.12.5
Compositing
Constrained assay sample lengths within the defined mineralized veins range from 0.15 m to 4.85
m, with an average sample length of 1.26 m and a mode of 1.50 m. A total of 20% of the
constrained assay sample lengths equal 1.00 m and 85% are between 0.50 m and 1.50 m. All
constrained assay samples were composited to 1.00 m in order to ensure equal sample support.
Length-weighted composites were calculated within the defined mineralized veins. A small
number of un-sampled intervals in the data were treated as null intervals due to the presence of
mining voids, backfill and surface rock piles. Backfill and surface rock assays were not included
in the compositing process.
The compositing process started at the first point of intersection between the drill hole and the
mineralized vein intersected, and halted upon exit from the mineralization. Downhole residual
composites that were less than half the compositing length were discarded in order to not introduce
a short sample bias into the composite sample population. The wireframes that represent the
mineralized veins were used to back-tag a rock code variable into the composite workspace.
The composite data were visually validated against the mineralization wireframes, and extracted
for analysis and grade estimation. Composite summary statistics are listed in Table 14.20.
TABLE 14.20
SUMMARY COMPOSITE STATISTICS
Element
Ag (g/t)
Au (g/t)
Cu (%)
14.12.6
Count
296
296
237
Average
34.3
0.57
0.011
StdDev
72.3
0.99
0.014
CoV
2.11
1.73
1.26
Minimum
0.001
0.0001
0.0006
Maximum
688.9
7.36
0.085
Treatment of Extreme Values
Capping thresholds were determined by the analysis of individual composite
log-probability distributions (Figure 14.11). Composites are capped to the defined threshold prior
to estimation (Tables 14.21 and 14.22), and capped composites are further restricted to a maximum
range of influence of 40.0 m. No capping was applied to copper composites.
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TABLE 14.21
SILVER COMPOSITE CAPPING THRESHOLDS
Veins
Cerro Colorado 1
Cerro Colorado 2
Threshold
(g/t)
120
60
Average
Uncapped
Value
(g/t)
43
23
Number
Capped
14
5
Average
Capped
Value
(g/t)
29
19
TABLE 14.22
GOLD COMPOSITE CAPPING THRESHOLDS
Veins
Threshold
(g/t)
Cerro Colorado 1
Cerro Colorado 2
3.00
2.00
Average
Uncapped
Value
(g/t)
0.64
0.53
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Number
Capped
10
6
Average
Capped
Value
(g/t)
0.56
0.48
Page 253 of 577
FIGURE 14.11
LOG-PROBABILITY PLOTS FOR COMPOSITES
Source: P&E (2024)
14.12.7
Continuity Analyses
Three-dimensional continuity analysis (variography) was conducted on the domain-coded
uncapped Ag and Au composite data using isotropic median indicator semi-variograms.
Standardized spherical models were used to model the experimental semi-variograms in order to
establish a reasonable classification range (Figure 14.12), with observed ranges of approximately
20 to 30 m.
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FIGURE 14.12
CERRO COLORADO SEMI-VARIOGRAMS
Source: P&E (2024)
14.12.8
Block Model
A rotated block model was established with the block model limits selected in order to cover the
extent of the mineralized structures, the potential open pit dimensions, and to reflect the general
nature of the mineralized zones (Table 14.23). The block model consists of separate variables for
estimated grades, volume percent wireframe block inclusion, rock codes, bulk density and
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classification attributes. A volume percent block inclusion attribute was used in order to ensure
that mineralized domain volumes are accurately reported.
TABLE 14.23
BLOCK MODEL SETUP
Direction
Minimum X
Minimum Y
Minimum Z
Rotation
14.12.9
Number of
Blocks
611,600
460
2,325,200
800
450
250
40° counter-clockwise
Origin
Block
Size
5m
5m
5m
Grade Estimation, Bulk Density and Classification
A bulk density value of 2.60 t/m3 was used for the Cerro Colorado Mineral Resource Estimate,
based on the overall average mineralized bulk density for the Los Ricos South Project.
Block grades for gold and silver were estimated by Inverse Distance Cubed (“ID3”) estimation of
capped composites using a minimum of four and a maximum of 12 composites. The orientation of
the search ellipsoid was defined by the modeled variography, observed grade trends and historical
mining. Composite samples were selected within a 240 m x 240 m x 40 m ellipsoid oriented
parallel to the modeled mineralization Sample selection was restricted to a maximum of three
composite samples from a single drill hole. Search and grade estimation were constrained by the
individual mineralized veins, which define hard boundaries for grade estimation.
Block grades for copper were estimated by Inverse Distance Squared (“ID2”) estimation of
uncapped composites using a minimum of four and a maximum of twelve composites.
The parameters used to define the classification limits included spatial analysis, drill hole spacing,
and the observed continuity of the mineralization. All blocks within 60 m of two or more drill
holes were classified as Indicated, and all additional estimated blocks were classified as Inferred.
14.12.10 Mineral Resource Estimate
Open pit Mineral Resources reported herein have been constrained within an optimized
constraining pit shell. The results from the optimized pit shell are used solely for the purpose of
reporting Mineral Resources and include Indicated and Inferred Mineral Resources. Pitconstrained Mineral Resources are reported using a cut-off grade of 40 g/t AgEq. The optimized
pit shell is shown in Appendix E.
The Mineral Resource has an effective date of January 14, 2025 (Table 14.24).
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TABLE 14.24
CERRO COLORADO MINERAL RESOURCE ESTIMATE (1-8)
Mt
Ag
(g/t)
Au
(g/t)
5.8
43
0.90
347
34
0.87
Cu
AgEq
(%)
(g/t)
Indicated
0.01
121
Inferred
0.01
109
AgEq
(koz)
AuEq
(g/t)
AuEq
(koz)
2,243
1.43
26.4
1,217
1.28
14.3
Notes:
1.
Mineral Resources, which are not Mineral Reserves, do not have demonstrated economic viability. The
estimate of Mineral Resources may be materially affected by environmental, permitting, legal, title, taxation,
socio-political, marketing, or other relevant issues.
2.
The Inferred Mineral Resource in this estimate has a lower level of confidence than that applied to an Indicated
Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority
of the Inferred Mineral Resource could be upgraded to an Indicated Mineral Resource with continued
exploration.
3.
The Mineral Resources were estimated in accordance with the Canadian Institute of Mining, Metallurgy and
Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared
by the CIM Standing Committee on Reserve Definitions and adopted by the CIM Council.
4.
Historically mined areas were depleted from the Mineral Resource model.
5.
The pit-constrained cut-off grade of 40 g/t AgEq was derived from US$1,850/oz Au price, US$23.75/oz Ag
price, 85% Ag and 95% Au process recovery, US$28/t process and G&A cost. The constraining pit optimization
parameters were $2.10/t mineralized material and waste mining cost, and 45° pit slopes.
6.
Cerro Colorado Mineral Resource is constrained to open pit mining methods only.
7.
AgEq and AuEq were calculated at an Ag/Au ratio of 86:1 for pit-constrained Mineral Resources.
8.
Totals may not sum due to rounding.
14.12.11 Validation
The block model was validated visually by the inspection of successive cross-sections in order to
confirm that the block models correctly reflect the distribution of high-grade and low-grade values
(see Appendix C).
The average estimated block grades were compared to the average Nearest Neighbour (“NN”)
block estimate at a 0.001 g/t AgEq cut-off for Measured and Indicated Mineral Resources (Table
14.25). The results fall within acceptable limits for grade estimation.
TABLE 14.25
GRADE COMPARISON
Veins
Cerro Colorado 1 Ag
Cerro Colorado 2 Ag
Cerro Colorado 1 Au
Cerro Colorado 2 Au
Estimated
25
15
0.46
0.41
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NN
25
14
0.41
0.43
Page 257 of 577
The volume estimated was also checked against the reported volume of the individual mineralized
veins. Estimated volumes are based partial block volumes (Table 14.26). The results fall within
acceptable limits for grade estimation.
TABLE 14.26
VOLUME COMPARISON
Veins
Cerro Colorado 1
Cerro Colorado 2
Total
Wireframe Volume
(k m3)
1,095.5
543.8
1,639.3
Estimated Volume
(k m3)
1,096.0
542.9
1,638.9
A check for local estimation bias was completed by plotting vertical swath plots of the estimated
ID3 block grade and the Nearest Neighbour grade. The results demonstrate a reasonable level of
smoothing for the ID3 estimate and are within acceptable limits for linear estimation (Figure
14.13).
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FIGURE 14.13
SWATH PLOTS
Ag
300
250
Ag g/t
200
150
100
50
0
40
80
120
160
200
240
280
320
360
400
440
480
520
560
600
640
680
720
760
800
840
880
920
960
1000
1040
1080
1120
0
Strike Distance (m)
AGID
AGNN
Au
2.50
Au g/t
2.00
1.50
1.00
0.50
0
40
80
120
160
200
240
280
320
360
400
440
480
520
560
600
640
680
720
760
800
840
880
920
960
1000
1040
1080
1120
0.00
Strike Distance (m)
AUID
AUNN
Source: P&E (2024)
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15.0
MINERAL RESERVE ESTIMATE
15.1
MINERAL RESERVE SUMMARY
The Los Ricos South Project will consist of both underground (“UG”) and open pit mining
operations. Hence, the Mineral Reserve is sub-divided into underground and open pit Mineral
Reserves. The Mineral Reserve for the Los Ricos South Project is summarized in Table 15.1. The
Mineral Reserve Estimate has an effective date of January 14, 2025. Approximately 44% of the
Mineral Reserve is classified as Proven.
All currency in this Technical Report section is in US dollars unless otherwise indicated.
15.2
FACTORS THAT MAY AFFECT THE MINERAL RESERVE ESTIMATE
Information that will affect the cut-off value used for estimating the Mineral Reserve includes
metal prices, the US Dollar:Mexican Peso exchange rate, dilution, overall mine and process
variable and fixed costs, process recoveries, copper concentrate quality targets, and management
of the operation. The Authors consider that the underground mining and open pit methodologies,
design criteria and parameters used are appropriate.
External factors, such as permitting, social, environmental governance (“ESG”) and political
issues could affect the viability of the Project. Therefore, these factors could materially impact the
Mineral Reserve estimate for the Project.
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TABLE 15.1
LOS RICOS SOUTH MINERAL RESERVE ESTIMATE (1-8)
Classification
Proven
Probable
Subtotal Underground
Tonnes
(k)
3,902
3,611
7,512
187
152
170
Average Grade
Au
Cu
(g/t)
(%)
Underground
1.61
0.09
1.70
0.18
1.65
0.13
159
130
136
1,768
4,961
6,728
13
44
57
0.3
1.4
1.7
2,965
8,924
11,889
312
248
276
25,240
22,607
47,847
215
241
457
7.7
16.1
23.8
44,904
45,819
90,723
Ag
(g/t)
Proven
Probable
Subtotal Open Pit
580
2,140
2,720
95
72
77
Open Pit
0.72
0.02
0.64
0.03
0.66
0.02
Proven
Probable
Total Proven & Probable
4,482
5,751
10,233
175
122
145
Total
1.49
0.08
1.31
0.13
1.39
0.10
Contained Metal
Au
Cu
(koz)
(Mlb)
AgEq
(g/t)
Ag
(koz)
AgEq
(koz)
334
318
326
23,472
17,647
41,119
202
197
399
7.5
14.6
22.1
41,939
36,895
78,834
Notes:
1.
Mineral Reserves are based on Measured and Indicated Mineral Resource Classifications only.
2.
Mineral Reserves are reported using the 2014 CIM Definition Standards and 2019 Best Practices Guidelines and have an effective date
of January 14, 2025.
3.
Mineral Reserves are defined within mine plans and incorporate mining dilution and ore losses.
4.
Open Pit Mineral Reserves are based on metal prices of US$23.75/oz Ag, US$1,850/oz Au and US$4.00/lb Cu, and are constrained within
optimized pit shells and designs that use 45–48º overall wall slopes, and process recoveries of 86% Ag, 95% Au and 51% Cu.
5.
An Open Pit cut-off grade of 46.4 g/t AgEq is estimated to differentiate ore from waste and is based on cost assumptions of US$26.22/t
processing, US$4.11/t site general and administrative, and 0.5% government mining tax on net revenue. Mining costs are estimated at
US$2.10/t of ore and waste rock.
6.
Underground Mineral Reserves are based on metal prices of US$23.75/oz Ag, US$1,850/oz Au and US$4.00/lb Cu, and are constrained
within a mine design, and use process recoveries of 86% Ag, 95% Au and 77% Cu.
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7.
8.
An Underground marginal cut-off grade of 150 g/t AgEq is estimated to differentiate ore from waste, and is based on cost assumptions of
US$34.93/t processing, US$4.46/t site general and administrative, and mining costs of US$41.93/t. An Underground economic cut-off
grade of 210 g/t AgEq is estimated to account for sustaining capital development costs of US$53.86/t in addition to those used to calculate
the marginal cut-off grade.
Totals may not sum due to rounding.
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15.3
15.3.1
UNDERGROUND MINERAL RESERVE ESTIMATE
Block Model Modifications from Mineral Resource Estimate to Mineral
Reserve Estimate
Mineral Reserves, as defined by NI 43-101, consist of only converted Measured and Indicated
Mineral Resources. The block model utilized to estimate the underground Mineral Reserve has
been slightly modified from the original block model to account for mineralization within the
Historical Mine Workings (“HMW”) and to reduce edge effects in stope design along the limits of
the wireframes. The changes between the two block models are presented in Table 15.2.
TABLE 15.2
COMPARISON OF ORIGINAL MINERAL RESOURCE MODEL
AND REVISED MODEL
Field
Block Size (W x L x H, m)
M&I Mineral Resource2 (Mt)
AgEq (Moz)
Au (Moz)
Ag (Moz)
Cu (kt)
HMW AgEq3 (Moz)
HMW Au3 (Moz)
HMW Ag3 (Moz)
HMW Cu3 (kt)
Original Model
3x3x3
27.154
134.12
0.63
72.54
22.28
4.90
0.02
3.21
0.44
Revised Model
0.5 x 0.5 x Variable1
27.158
134.11
0.63
72.52
22.28
0.74
0.00
0.47
0.02
Notes:
1
Maximum dimension of sub-block is 0.5 m
2
Measured and Indicated Mineral Resource at zero cut-off grade
3
HMW Mineral Resources are separate from M&I Mineral Resources. See Section 15.3.2 for how these Mineral
Resources are accounted for in Mineral Reserve calculations.
As can be seen, no appreciable change to the Mineral Resources has been made during the
conversion of the block models. The reduction in the Mineral Resources in the HMW model from
the original model to the revised model is explained in the following section.
15.3.2
Accounting for Historical Mine Workings Material
Significant amounts of historical mining have been undertaken at the Property. HMW extents
have been reconstructed by GoGold personnel based on historical surveys, and drill hole intercepts
of backfill and voids have been used to verify their locations. Previous studies (P&E, 2023 PEA)
have used the assumption that material within these areas is unconsolidated, free-flowing, and
mineralized, and these assumptions are continued in this Technical Report. Updates to the density
and grades of the material have been performed for this Technical Report.
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15.3.2.1
Bulk Density
The bulk density of the HMW material was based on the drill hole intercepts located within the
provided HMW shapes. These intercepts returned one of three values: Air, Broken Rock or Solid
Rock. While Air and Broken Rock are expected and self-explanatory results, Solid Rock is
indicative of either extensive slabbing around the HMW shapes (large sheets of rock spalled off
of stope walls), or that the HMW shapes do not precisely align with the actual mined-out areas
(which is more likely). Since Solid Rock makes up a very small percentage of the overall
intercepts, all three intercept types were utilized to generate a weighted average of the density of
the material within the HMW shapes. Broken Rock was assumed to have the same original in-situ
bulk density as Solid Rock, and to have expanded in volume after breaking up by a factor (Swell
Factor) of 40%. Table 15.3 shows the calculated bulk density of the HMW material.
TABLE 15.3
OLD WORKINGS INTERCEPT DATA AND BULK DENSITY CALCULATIONS
1
15.3.2.2
Material
Intercept
(m)
Air
Solid Rock
Broken Rock
Total
111.25
2.27
85.46
198.98
Bulk Density
of Material
(t/m3)
0
2.541
2.54 / 1.4 = 1.81
0.81
HMW
Volume
(%)
55.91
1.14
42.95
100
2.54 t/m3 is the background waste density for the Underground block model.
Grade
Additional grade testing was performed on samples intercepting the HMW shapes. Box-andwhisker plots of all samples were generated for each metal to determine outliers in the data sets,
and all samples outside of the Q1-Q4 range were discarded. After discarding outliers, 112
intercepts remained for analysis. The results of the analysis are shown in Table 15.4.
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TABLE 15.4
HISTORICAL WORKINGS INTERCEPT DATA AND
METAL GRADE CALCULATIONS
Item
Number of Intercepts
Total Length of Intercepts (m)
Au g/t – Minimum
Au g/t – Maximum
Au g/t – Average
Au g/t - Weighted Average
Ag g/t – Minimum
Ag g/t – Maximum
Ag g/t – Average
Ag g/t - Weighted Average
Cu % – Minimum
Cu % – Maximum
Cu % – Average
Cu % – Weighted Average
Blanket Au g/t for HMW material
Blanket Ag g/t for HMW material
Blanket Cu % for HMW material
Blanket AgEq g/t for HMW material
HMW material value per tonne at $0.6567/g AgEq
Value
112
199.0
0.003
1.204
0.286
0.293
0.80
167.16
44.75
46.67
0.0000
0.0366
0.0047
0.0047
0.29
44.75
0.005
70.2
$46.10
AgEq contents were calculated using the following formula:
𝐴𝑔𝐸𝑞𝑔𝑝𝑡 = 𝐴𝑔𝑔𝑝𝑡 + 𝐴𝑢𝑔𝑝𝑡 ∗ 86.1 + 𝐶𝑢𝑝𝑐𝑡 ∗ 103.4
15.3.2.3
Treatment of HMW Mineral Resources
Material derived from HMW backfill is not classified as Measured or Indicated; however, is
surrounded entirely by Mineral Resources of one or both of those two classifications. This backfill
material has been included in the Mineral Reserve as a Probable Mineral Reserve since it is known
to:
1. Contain mineralization;
2. Be an undesirable yet unavoidable result of mining operations (similar to other forms
of dilution); and
3. Make up an insignificant portion of overall recovered metal.
HMW Mineral Resources make up approximately 2.7% of UG Mineral Reserve tonnes and 0.6%
of UG Mineral Reserve value (0.44 Moz of 78.83 Moz AgEq).
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15.3.3
Cut-off Grade Calculations
Cut-off Grades (“COGs”) for this Technical Report were derived from a combination of contractor
estimates, first-principles calculations, metal price estimates, and cost estimates from previous
studies.
15.3.3.1
Metal Prices, Recoveries and Equivalency Calculations
Mineral Reserves for the Los Ricos South UG are reported in Silver Equivalent (“AgEq”). This
method divides the value of a non-silver metal by a unit value of silver to simulate an amount of
silver equivalent to the value of the non-silver metal. To generate these equivalencies, it is
necessary to know the process recoveries, metal prices, and transport/refining/sale costs of each
metal component. Since no smelter contract existed when these equivalency calculations were
performed, only process recoveries and metal prices are included in the AgEq calculations, as
shown in Table 15.5.
TABLE 15.5
MINERAL RESERVE METAL PRICES AND PROCESS RECOVERIES
Metal
Gold (Au)
Silver (Ag)
Copper (Cu)
Silver Equivalent
(AgEq)
oz
oz
lb
Process
Recovery
(%)
95
86
77
Value per
Metal Grade
(US$/unit)
56.50
0.6567
67.90
oz
86
0.6567
Price
(US$/unit)
Price
Unit
1,850
23.75
4.00
23.75
Grade
Unit
g/t
g/t
%
g/t
Value per metal grade is calculated by multiplying the price by the recovery and converting to the
grade unit utilized in the block model (grams per tonne for precious metals, percent by mass for
base metals). Dividing the relative values per metal grade by the base silver value provides the
following formula for silver equivalent:
𝐴𝑔𝐸𝑞𝑔𝑝𝑡 = 𝐴𝑔𝑔𝑝𝑡 ∗
0.6567
56.50
67.90
+ 𝐴𝑢𝑔𝑝𝑡 ∗
+ 𝐶𝑢𝑔𝑝𝑡 ∗
0.6567
0.6567
0.6567
This yields the same formula previously shown in Section 15.3.2.2:
𝐴𝑔𝐸𝑞𝑔𝑝𝑡 = 𝐴𝑔𝑔𝑝𝑡 + 𝐴𝑢𝑔𝑝𝑡 ∗ 86.1 + 𝐶𝑢𝑝𝑐𝑡 ∗ 103.4
Alternatively, the same formula re-written for gold equivalency is:
𝐴𝑢𝐸𝑞𝑔𝑝𝑡 = 𝐴𝑢𝑔𝑝𝑡 ∗
56.50
0.6567
67.90
+ 𝐴𝑔𝑔𝑝𝑡 ∗
+ 𝐶𝑢𝑔𝑝𝑡 ∗
56.50
56.50
56.50
𝐴𝑢𝐸𝑞𝑔𝑝𝑡 = 𝐴𝑢𝑔𝑝𝑡 + 𝐴𝑔𝑔𝑝𝑡 ∗ 0.0116 + 𝐶𝑢𝑝𝑐𝑡 ∗ 1.2017
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15.3.3.2
Contractor Estimates
Estimates of unit rates for development, production, construction and rehabilitation activities were
provided by three different mining contractors with significant presence in Mexico. Estimates were
based on quantities and heading sizes derived from the previous PEA (P&E, 2023), and were
received in Q1 of 2024. Details of manning, machine hours, productivity, consumables and rental
rates were provided, and used to benchmark first-principles calculations for the same activities.
15.3.3.3
First-Principles Costs Estimate
To avoid reliance on any individual contractor’s rates for the ongoing success of the underground,
the Authors generated benchmarked first-principles calculations for all required underground
activities. Estimates included allowances for contractor overheads and profits in addition to the
items mentioned in Section 15.3.3.13.2, and generated costs within the boundaries of the provided
contractor bids with similar machine hours, consumption rates and manning.
15.3.3.4
Additional Site Cost Estimates
Additional operating costs at site include processing, tailings storage, cemented paste backfill
(“pastefill”), Company and contractor indirect costs (staff salaries outside of the unit rates) and
G&A. For the purposes of generating cut-off grades, estimates of these costs were derived from
the previous PEA (P&E, 2023) with appropriate modifications.
15.3.3.5
Economic, Marginal and Break-Even Cut-off Grade Estimates
In-situ mining shapes were designed around a marginal mineralized shell with an economic core.
An overall Economic Cut-off Grade (“ECOG”) was calculated, and then a Marginal Economic
Cut-off Grade (“MCOG”) was calculated to determine what additional tonnes adjacent to
economic shapes could be added to the Mineral Resource base without incurring additional indirect
costs. Finally, a Break-Even (Development) Cut-off Grade (“DCOG”) was calculated to determine
whether material that was necessary to mine as a result of development activities could incidentally
support the additional costs required to process that material versus sending it to the waste rock
storage area.
Initial COG calculations are shown in Table 15.6.
TABLE 15.6
INITIAL CUT-OFF GRADE CALCULATIONS
Item
Units
OPEX
Ore
$/t
Development
Production Drill &
$/t
Blast
ECOG
MCOG
Component Component
HMW
DCOG
MCOG
Component
Component
11.53
11.53
-
-
6.30
6.30
-
-
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TABLE 15.6
INITIAL CUT-OFF GRADE CALCULATIONS
Item
Load and Haul
Backfill
Indirect and G&A
Processing
OPEX Subtotal
Capital Development
All Other Costs and
Contingencies
CAPEX Subtotal
Total1
Calculated
Cut-off
2
Grade
Rounded
Cut-off
2,3
Grade
Final Cut-off Grade2
Units
ECOG
MCOG
Component Component
HMW
DCOG
MCOG
Component
Component
9.87
4.46
4.46
34.93
34.93
49.26
39.39
-
$/t
$/t
$/t
$/t
$/t
$/t
14.23
9.87
4.46
34.93
81.33
20.58
14.23
9.87
4.46
34.93
81.33
-
$/t
33.28
-
-
-
$/t
$/t
53.86
135.20
81.33
49.26
39.39
g/t AgEq
205.9
123.9
75.0
60.0
g/t AgEq
210
130
80
60
g/t AgEq
210
150
80
65
Notes:
1
Totals may not sum due to rounding.
2
Uses $0.6567 value per in-situ gram AgEq.
3
Rounded up to nearest 10 g/t increment
Economic COG
The underground ECOG was estimated at 210 g/t AgEq. At this COG, stopes would be expected
to generate a profit after covering all costs for the underground, both operating and capital. This
COG was used to generate stoping targets throughout the mine.
Marginal COG
The underground MCOG was estimated at 130 g/t AgEq. At this grade, a stope will cover all
directly attributable costs of production (OPEX development, drill and blast, excavate and fill, and
applicable salaries/G&A costs). Since some stopes may require a portion of capital development
(footwall drift (“FWD”) or infrastructure) in addition to operating development, a slight
contingency was applied to the true marginal cost and 150 g/t AgEq was used as the final MCOG.
Stopes generated at this MCOG were used to add adjacent and contiguous stopes to the stope set
generated at ECOG where discontinuities (either longitudinal or transverse) existed in stoping
areas on a level.
Subsequently, an arbitrary mid-point COG between the selected ECOG and MCOG of 180 g/t
AgEq was used to generate a third set of stopes to replace stopes generated at the MCOG to
generate improved margin.
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HMW Marginal COG
The COG for the HMW areas was estimated at 80 g/t AgEq, which would cover all items in the
MCOG, with the removal of drill and blast processes and their associated services and electrical
power costs as well as all OPEX development costs. It is possible that the material will also be
slightly less expensive to process, given its particle sizes; however, no change to process costs has
been included.
Given that the material is expected to be free-flowing and is expected to be completely recovered
to its full extents without the need for demarcation, delineation, or any other ancillary operations
other than excavation, processing, and filling, the HMW MCOG is not applied for the generation
of any stopes: historical survey voids are used as stope limits instead. Since HMW stope grades
are below the HMW MCOG value, they are classified as diluting material.
Development COG
The COG for development material was estimated at 60 g/t AgEq. For this development material,
since tailings storage space is at a premium and there is a desire to fill all available underground
voids, it is expected that all development material would be transported to surface regardless of
grade, and only the additional costs associated with processing and indirects/G&A would be
attributable, since the material would need to be excavated and transported (and backfilled in some
cases) regardless. Since it is planned to stockpile significant portions of the material mined at this
COG, a contingency of 5 g/t AgEq was added to account for stockpile management and rehandle
costs, resulting in a final DCOG of 65 g/t AgEq.
15.3.4
Dilution
Dilution describes the necessary inclusion of sub-economic material in the mining process to
recover economic material. For the Los Ricos South UG, there are two primary sources of dilution:
waste dilution (internal and external) and backfill dilution resulting from overbreak into adjacent
filled stopes and over-digging (floor gouging) backfill on stope floors.
15.3.4.1
Internal Dilution
Internal dilution is the deliberate inclusion of sub-economic material in the design of a mining
shape for the purposes of extracting economic material in a practical manner. This dilution may or
may not be mineralized.
Internal dilution is calculated by:
𝐷𝑖𝑙𝑢𝑡𝑖𝑜𝑛𝐼𝑛𝑡𝑒𝑟𝑛𝑎𝑙 = 1 −
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝐴𝑏𝑜𝑣𝑒 𝐶𝑂𝐺
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝑇𝑜𝑡𝑎𝑙
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Internal dilution on stopes is estimated at 21.2% by mass. Internal dilution on development is
estimated at 25.6% by mass. Combined, internal dilution on all UG material in the process plant
feed averages 21.8% by mass.
15.3.4.2
External Dilution
External dilution on mining shapes is the result of overbreak of the shape outside of the planned
limits. External dilution for the Los Ricos South UG can either be into previously unmined rock
(whether mineralized or unmineralized), into pastefill (unmineralized) or into the HMW material
(mineralized). Where the overbreak is into previously unmined rock, it is possible for the external
dilution to carry small pockets of material that exceed the MCOG grade, but whose inclusion in
the mining shape requires addition of excess material that makes the pockets uneconomical to
include in the mining plan.
Since the vast majority of mineralized development included in the Mineral Reserves will
overbreak into stopes and the net balance of stope tonnes and overbroken development tonnes will
be near zero, overbreak on mineralized development is treated as 0%.
For stopes, overbreak into the hanging wall (“HW”) is estimated to progress 0.8 m on the Abra
Vein and 0.6 m on the Eagle Vein. Overbreak into the footwall (“FW”) is estimated to progress
0.4 m on all veins. In addition to this material, a quantity of pastefill (“PF”) dilution is expected
from end walls and side walls of stopes adjacent to previously filled areas, along with some dilution
from gouging the floor (for overhand mining) or overbreak into the back (for underhand mining).
A general value of 3.5% PF dilution by mass is used for all stopes in previously unmined ground,
with 5.0% PF dilution by mass being used in areas adjacent to or including HMW shapes.
Stope external dilution is calculated as:
𝐷𝑖𝑙𝑢𝑡𝑖𝑜𝑛𝐸𝑥𝑡𝑒𝑟𝑛𝑎𝑙 =
𝑀𝑎𝑠𝑠𝑂𝑣𝑒𝑟𝑏𝑟𝑒𝑎𝑘,𝑇𝑜𝑡𝑎𝑙
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝑇𝑜𝑡𝑎𝑙 + 𝑀𝑎𝑠𝑠𝑂𝑣𝑒𝑟𝑏𝑟𝑒𝑎𝑘,𝑇𝑜𝑡𝑎𝑙
External dilution on stopes is estimated at 20.3%. External dilution on development is nominally
0% for reasons previously stated. Combined, external dilution on all underground material in the
process plant feed averages 18.1% by mass.
15.3.4.3
Total Dilution and Additional Considerations
Total dilution, utilizing the components of internal and external dilution defined above, is
calculated as:
𝐷𝑖𝑙𝑢𝑡𝑖𝑜𝑛 𝑇𝑜𝑡𝑎𝑙 = 1 −
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝐴𝑏𝑜𝑣𝑒 𝐶𝑂𝐺
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝑇𝑜𝑡𝑎𝑙 + 𝑀𝑎𝑠𝑠𝑂𝑣𝑒𝑟𝑏𝑟𝑒𝑎𝑘,𝑇𝑜𝑡𝑎𝑙
The total dilution, calculated in this manner, is estimated at 35.9% by mass. Since some material
that passes cut-off is incidentally mined as a component of external dilution, the actual total
economic dilution is somewhat less than that, at 32.1%, using the formula below:
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𝐷𝑖𝑙𝑢𝑡𝑖𝑜𝑛𝐸𝑐𝑜𝑛𝑜𝑚𝑖𝑐 = 1 −
15.3.5
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝐴𝑏𝑜𝑣𝑒 𝐶𝑂𝐺 + 𝑀𝑎𝑠𝑠𝑂𝑣𝑒𝑟𝑏𝑟𝑒𝑎𝑘,𝐴𝑏𝑜𝑣𝑒𝐶𝑂𝐺
𝑀𝑎𝑠𝑠𝑃𝑙𝑎𝑛𝑛𝑒𝑑 𝑆ℎ𝑎𝑝𝑒,𝑇𝑜𝑡𝑎𝑙 + 𝑀𝑎𝑠𝑠𝑂𝑣𝑒𝑟𝑏𝑟𝑒𝑎𝑘,𝑇𝑜𝑡𝑎𝑙
Mining Loss
Mining loss is defined as a portion of material left behind in an excavation due to any or all of
drilling, blasting, excavating, or ground support issues. For the Los Ricos South UG, mining loss
is assumed to be:
•
•
•
•
•
1% for development. Easy access and ability to reblast results in high recovery;
5% for longitudinal retreat stopes mined overhand with downhole drilling;
7.5% for transverse stopes mined overhand with downhole drilling. This is an average
of 5% loss for primary stopes and 10% loss for secondary stopes, applied to all stopes;
10% for longitudinal retreat stopes mined overhand with uphole drilling; and
15% for transverse stopes mined underhand with uphole drilling.
Overall average underground mining loss for the Deposit is estimated at 8.8% by mass.
15.3.6
Underground Mineral Reserve Estimate
Mineral Reserves consist of Measured and Indicated Mineral Resources. As mentioned above,
mined HMW material is classified as Probable Reserves. HMW Mineral Resources make up
approximately 2.7% of underground Mineral Reserve tonnes and 0.6% of underground Mineral
Reserve value.
15.3.7
Mining Plan Summary
Table 15.7 shows the contents of the mine plan by Mineral Resource classification. As stated in
Section 15.3.2.32, HMW material was not classified as Measured or Indicated Resources in the
block model, and has been identified separately in Table 15.7 for clarity.
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TABLE 15.7
UNDERGROUND MINE PLAN SUMMARY BY MINERAL RESOURCE CLASS1
Item
Fully
Diluted
Stopes
Mining
Losses
Mine
Plan
Classification
Tonnes
(k)
Measured
Indicated
Subtotal M&I
Other
Mine Plan Total
Measured
Indicated
Subtotal M&I
Other
Mine Plan Total
Measured
Indicated
Subtotal M&I
Other
Mine Plan Total
4,369
3,631
8,000
235
8,235
468
221
689
34
723
3,902
3,410
7,312
201
7,512
Ag
Grade
(g/t)
187
159
175
44
171
189
169
183
44
176
187
158
174
44
170
Au
Grade
(g/t)
1.60
1.79
1.68
0.28
1.64
1.51
1.81
1.61
0.28
1.54
1.61
1.79
1.69
0.28
1.65
Cu
Grade
(%)
0.084
0.194
0.134
0.005
0.130
0.064
0.175
0.099
0.005
0.095
0.087
0.195
0.137
0.005
0.134
AgEq
Grade
(g/t)
333
333
333
68
326
326
343
331
68.
319
334
333
334
68
326
Notes:
1
Totals may not sum due to rounding.
15.3.8
Underground Mineral Reserve Estimate
Table 15.8 presents the Mineral Reserve Estimate for the Los Ricos South UG. As stated in Section
15.3.2.32, HMW material is classified as Probable Mineral Reserve.
TABLE 15.8
UNDERGROUND MINERAL RESERVE ESTIMATE (1-6)
Classification
Proven
Probable
Total Proven
Probable
and
Tonnes
(k)
3,902
3,611
Ag Grade
(g/t)
187
152
Au Grade
(g/t)
1.61
1.70
Cu Grade
(%)
0.09
0.18
AgEq Grade
(g/t)
334
318
7,512
170
1.65
0.13
326
Notes:
1.
Mineral Reserves are based on Measured and Indicated Mineral Resource Classifications only.
2.
Mineral Reserves are reported using the 2014 CIM Definition Standards and 2019 Best Practices Guidelines
and have an effective date of January 14, 2025.
3.
Mineral Reserves are defined within mine plans and incorporate mining dilution and ore losses.
4.
Underground Mineral Reserves are based on metal prices of US$23.75/oz Ag, US$1,850/oz Au and US$4.00/lb
Cu, and are constrained within a mine design, and use process recoveries of 86% Ag, 95% Au and 77% Cu.
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5.
6.
An Underground marginal cut-off grade of 150 g/t AgEq is estimated to differentiate ore from waste, and is
based on cost assumptions of US$34.93/t processing, US$4.46/t site general and administrative, and mining
costs of US$41.93/t. An Underground economic cut-off grade of 210 g/t AgEq is estimated to account for
capital development costs of US$53.86/t in addition to those used to calculate the marginal cut-off grade.
Totals may not sum due to rounding.
Table 15.9 shows the underground Mineral Reserve Estimate breakdown by grade bin.
TABLE 15.9
UNDERGROUND MINERAL RESERVE BY GRADE BIN
Classification
Proven
Probable
Total
Item
High Grade
Mid-Grade
Low-Grade
Dev-Grade
High Grade
Mid-Grade
Low-Grade
Dev-Grade
High Grade
Mid-Grade
Low-Grade
Dev-Grade
Bin Min
Bin Max
(g/t AgEq)
450
>450
300
450
150
300
65
150
450
>450
300
450
150
300
65
150
450
>450
300
450
150
300
65
150
Tonnage
Portion
(%)
19
26
50
5
15
29
49
7
17
27
49
6
AgEq Oz
Portion
(%)
37
28
34
2
32
32
34
2
35
30
34
2
Approximately 65% of all AgEq ounces in the Mineral Reserve are from mining areas grading
>300 g/t AgEq. Approximately 6% of all mined tonnes are from development areas grading below
the stope MCOG but above DCOG, containing 2% of the Mineral Reserve AgEq ounces.
15.4
OPEN PIT MINERAL RESERVE ESTIMATE
The upper portion of the Los Ricos South Deposits are near surface and lend themselves to
conventional open pit mining methods. Two open pits areas will be developed towards the end of
the life of underground mining operations. These two pit areas are the Abra Pit and Cerro Colorado
Pit (North and South).
The open pit Mineral Reserve is the portion of the Mineral Resource that has been deemed
economically mineable by the open pit production plan and within designed open pits.
The design of the open pits requires several steps:
1. Complete pit optimizations to select the optimal pit shells to be used for the mine
designs;
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2. Design operational pits (with ramps and benches) based on the optimal pit shells; and
3. Develop a life-of-mine (“LOM”) open pit production schedule, based on supplying up
to 2,000 tonnes of ore per day (0.72 million ore tonnes per year) to the process plant.
15.4.1
Open Pit Optimization
A series of pit optimizations were completed on the Mineral Resource block models using
Datamine NPV SchedulerTM software. This optimization process produces a series of nested pit
shells each containing ore that is economically mineable according to a given set of physical and
economic parameters.
Pit optimizations were run separately for the Abra Veins and the Cerro Colorado Veins.
The pit optimizations were run using the parameters shown in Table 15.10. At the optimization
stage of mine planning, it was assumed that waste rock materials would be hauled to a waste rock
storage facility in close proximity to each pit.
For pit optimization, a base case gold price of US$1,850/oz and a silver price of US$23.75/oz were
used. The pit optimization analysis included only Measured and Indicated Mineral Resources as
ore feed. Inferred Mineral Resource was considered waste material in the pit optimization step.
Metal revenue factors from 1 to 120% were applied in the optimization, with the base case being
the 100% revenue factor.
TABLE 15.10
PIT OPTIMIZATION PARAMETERS
$/oz
$/oz
$/lb
%
Cerro
Colorado
(no UG)
Meas & Ind only
1,850
23.75
4.00
0.5%
Abra and Eagle
(UG stopes
backfilled
Meas & Ind only
1,850
23.75
4.00
0.5%
$/t
$/t
$/t
$/t
$/t
2.10
n/a
2.10
26.22
4.11
2.10
2.10
2.10
26.22
4.11
%
%
%
95
86
51
95
86
51
Item
Classifications
Gold Price
Silver Price
Copper Price
Government tax on Revenue
Operating Costs
Waste Mining & Haul Cost
Backfilled Stope Mining Cost
Ore Mining & Haul Cost
Processing Cost + Dry Stack
G&A
Process Recovery
Gold recovery
Silver recovery
Copper recovery
Units
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TABLE 15.10
PIT OPTIMIZATION PARAMETERS
Item
Units
Cerro
Colorado
(no UG)
Abra and Eagle
(UG stopes
backfilled
Cut-off Grades
Incremental Opex
Cut-off Grade (AgEq)
Pit Optimization Slope Angles
$/t
g/t
degrees
30.33
46.4
45
30.33
46.4
48
Notes: UG = underground, Meas & Ind = Measured and Indicated Mineral Resources.
15.4.1.1
Abra Pit Optimization Result
The Abra Pit is planned to be developed above the underground mining operation, towards the end
of the underground mine life. Hence the Mineral Resource block model used for pit optimization
was depleted for ore that would be mined prior by the underground operation. Much the highergrade ore at depth will have been recovered via underground mining.
The mined-out stopes would be backfilled with cemented paste backfill and could be intersected
by the open pit operation. The depth of the optimized open pit would be driven by mining the
higher-grade crown pillar as well as any unmined lower-grade “halo” that might be surrounding
the upper underground workings. An underground mining cut-off grade is generally higher than
an open pit mining cut-off grade, hence lower-grade ore would not be mined in the stopes.
The results of Abra Pit optimization are shown in Figure 15.1. The NPV curve starts to flatten
above a revenue factor of ~70%. To maximize ore recovered, the revenue factor 96% shell was
selected as the basis for the Abra open pit design.
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FIGURE 15.1
ABRA PIT OPTIMIZATION RESULTS
Source: P&E (2025)
15.4.1.2
Cerro Colorado Pit Optimization Result
No underground mining is planned at the Cerro Colorado Deposit. Hence the objective was to
maximize ore recovery within the open pits.
The results of Cerro Colorado pit optimization are shown in Figure 15.2. The NPV curve flattens
above a revenue factor of 90%. The 100% revenue factor shell was selected as the basis for the pit
design.
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FIGURE 15.2
CERRO COLORADO PIT OPTIMIZATION RESULTS
Source: P&E (2025)
15.4.2
Open Pit Designs
The open pits were designed using the selected optimized shells as a guide for sizing of each pit.
The design of each pit examined preferred access points along the pit periphery, and then added
benches, ramps and haul roads according to the pit slope parameters shown in Table 15.11.
The open pit designs used a 10% ramp at a width of 15 m for double lane traffic and 10 m for
single lane traffic. Single lane haul roads and ramps were used in the bottom benches of the pits to
minimize the addition of excess waste rock from expanding the pit walls outwards more than
required. Cerro Colorado consists of two small pits and were designed using a 10 m ramp width.
The Abra Pit design is shown in Figure 15.3 and the Cerro Colorado Pit is shown in Figure 15.4.
The Abra Pit would be mined in two phases, as described in Section 16.
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FIGURE 15.3
ABRA PIT DESIGN
Source: P&E (2025)
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FIGURE 15.4
CERRO COLORADO PIT DESIGN
Source: P&E (2025)
15.4.2.1
Open Pit Geotechnical Studies
Geotechnical studies have been completed in 2020, 2023, and 2024. The pit slope criteria used in
this Technical Report were provided by WSP in a report titled “Los Ricos South Project – Open
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Pit Design Recommendations” dated February 21, 2025. An Executive Summary of the WSP
report is provided below.
The following is a summary of the open pit slope geometry design recommendations for GoGold’s
proposed open pits at the Los Ricos South Property, in the State of Jalisco, near the town of
Magdelena, in Mexico. This section contains a summary of the review of available and recently
collected geotechnical and hydrogeological data, assessment of this data, development of rock
mass characteristics at the Los Ricos South Property, and stability assessments and assignment of
slope geometry recommendations to different design sectors in the proposed open pits.
Two open pits are included in this study, the Abra Eagle Pit and the Cerro Colorado Pit (Figure
15.5). The Abra Eagle Pit intersects two ore bodies, the Abra and the Aguilas ore bodies. The
Cerro Colorado Pit (separated into North and South pits) intersects a separate ore body.
FIGURE 15.5
PLAN VIEW OF THE LOS RICOS SOUTH OPEN PITS
Source: WSP (2025)
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Summary of Geotechnical Site Characterization
The Los Ricos South geotechnical site characterization has incorporated available geotechnical
drilling and laboratory strength data (Figure 15.6). Geotechnical drilling data was available from
an investigation targeting the Abra Eagle Pit area and a separate investigation targeting the
associated underground deposit at the Los Ricos South Property. No geotechnical drilling was
available near the Cerro Colorado Pit, although drill core photos and point load tests performed on
exploration core near the Cerro Colorado Pit were provided for review.
FIGURE 15.6
ABRA OPEN PIT SHOWING GEOTECHNICAL DRILL HOLE TRACES
Source: WSP (2025)
Note: Not to scale.
The geotechnical characterization described below focused on the available geotechnical data near
the Abra Eagle Pit and did not initially attempt to characterize the rock masses near the Cerro
Colorado Pit (which is discussed later). The resulting characteristics are similar to many
parameters defined as part of the engineering geology model for the underground study at the Abra
and Aguilas Ore Bodies. The resulting rock mass parameters derived from the characterization
include:
•
The definition of five geotechnical domains: Rhyolite Tuff, Andesite FW, Andesite
HW, Abra Ore Zone and Aguilas Ore Zone (Table 15.11).
o Other domains defined as part of the underground study were found to not greatly
intersect or influence the open pit.
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TABLE 15.11
SUMMARY OF THE AVAILABLE
GEOTECHNICAL LOGGING FOR EACH
DOMAIN
Domain
Andesite FW
Andesite HW
Rhyolite Tuff
Abra Ore Zone
Aguilas Ore Zone
•
Metres Logged
(m)
1,577
736
640
165
32
Rock mass quality was evaluated using the RMR76 rock mass classification systems
(Table 15.12). The rock masses are typically found to be Fair to Poor quality. The
footwall rock masses (Andesite FW and the Rhyolite Tuff) represent the best ground
conditions expected in the Abra Eagle Pit; while the hanging wall rock masses and the
Abra Ore Zone represent the worst ground conditions expected in the open pit.
TABLE 15.12
CLASSIFICATION RESULTS FOR EACH DOMAIN
Lithology
Andesite FW
Rhyolite Tuff
Abra Ore Zone
Andesite HW
Aguilas Ore Zone
Mean RMR76
59
55
46
39
55
Note: Design RMR76 Colored by Classification Number: Red (Very Poor Rock [<20]), Orange (Poor Rock [20 40]),
Yellow (Fair Rock [40 – 60]), Light Green (Good Rock [60 – 80]), and Green (Very Good Rock [80 – 100]).
•
Core samples of each domain were strength tested in the laboratory. Hoek Brown intact
strength failure criteria were developed for each geotechnical domain (Table 15.13).
Highly altered or fractured units, although present on the Project site are
underrepresented due to the inability to collect specimens or conduct valid tests in a
laboratory.
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TABLE 15.13
HOEK-BROWN STRENGTH PARAMETERS FOR INTACT ROCK BY
DOMAIN
Domain
[No. of Intact Tests]
Andesite FW [40]
Andesite HW [6]
Aguilas Ore Zone [9]
Abra Ore Zone [13]
Rhyolite Tuff [13]
σci
(MPa)
100
94
74
91
115
mi
r2
21.4
14.7
12.2
16.9
14.5
0.61
0.65
0.33
0.82
0.76
*’Intact Tests’ refers to a strength sample which failed through intact rock and was not influenced by sample defects
or planes of weakness.
•
Characterization of the rock mass fabric was evaluated from structural data collected
in the geotechnical drill holes. Three major sets and two minor sets were identified
including a set perpendicular to the ore body and a set parallel to the orientation of the
ore body (Table 15.14). The third major set (J3A and J3B) is less prominent than J1
and J2 but present in most lithologies and aligns with local faults. The main structural
feature is the parallel joint set (J1) with general orientation of 81° dip and 033° dip
direction; this joint set appears to be dominant throughout the Project site.
o Collected geotechnical data indicated that facture intensity in the FW zone is less
than in the HW zone. The HW zone had a high degree of faulting and broken drill
core zones than the ore zone and footwall domain. Due to this, drilling orientation
data was difficult to collect in the HW and ore bodies. No televiewer data was
collected as part of this study, limiting the understanding of structural orientation
and prevalence of each set.
o Limited information is available on joint persistence and spacing and should be
captured during future study stages or during operations.
•
The collected hydrogeological data from site indicates that the proposed Abra Eagle
Pit is likely to be dry. No information is available for the Cerro Colorado ground water
table.
•
Core intersecting the faults was observed to be damaged and altered. Core intersecting
lithological contacts were observed to have variable alteration and damage; however,
the exact location of the contacts is not well understood and a pattern of damage
variation could not be determined. Contacts were not included in the stability
assessments, which should be better defined and located during future study stages to
understand potential areas of poor ground on the open pit walls.
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TABLE 15.14
SUMMARY OF SELECTED STRUCTURAL SETS
Set
Dip
(°)
Major (blue) and
Discontinuity Sets
Dip
Direction
(°)
Minor
Stereonet
(red)
J1
J2
J3A
J3B
Ja
81
61
65
76
63
033
218
275
099
320
Jb
63
146
•
Seismic hazard peak ground acceleration at the Project site ranges from 0.05 to 0.08 g.
•
Shear strength of the open small-scale jointing and fault structures was assumed to be
32° and 19°, respectively. Direct shear tests should be performed during the next stages
of study to confirm assumptions regarding shear strength of bench scale structures.
o For pit scale stability assessments, the persistence of open jointing was assumed to
be 95%.
Kinematic Stability Assessment Results
The primary kinematic stability assessment for the Abra Eagle Pit was conducted using a
deterministic approach based on the identified structural sets and the following assumptions and
according to the sectors shown in Figure 15.7:
•
•
•
Cohesion: 0 kPa (assumed to account for blast disturbance) for both joints and faults;
Friction Angle (ϕ): 32° for joint sets and 19° for faults; and
Design Factor of Safety (FoS): 1.3 under dry conditions.
Table 15.15 provides an overview of the kinematic design results for Design Sector 1. This design
sector was split into 3 sub-sectors based on potential orientations that GoGold may want to use.
The final pit shell used in this Technical Report only uses Sector 1b recommendations.
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FIGURE 15.7
ABRA OREBODY TRENDING NORTHWEST-SOUTHEAST WITH
AVERAGE DIP DIRECTION OF 230° AND SHOWING DESIGN SECTORS
AN
TABLE 15.15
SUMMARY OF THE KINEMATIC ANALYSIS AND SIMPLIFIED DESIGN
FOR SECTOR 1 – FW DOMAIN
Design
Sector
(Dip
Direction)
Potential Kinematic Failure
Mechanism
Max
BFA
(°)
Bench
Height
(m)
Bench
Width
(m)
Max
IRA
(°)
Sector 1a
(140° - 160°)
Moderate potential for wedge failure
(J1/J3 & J3b); Low potential for
toppling (J3a) and planar failure (J3b)
65
20
8.5
48
Sector 1b
(160° - 300°)
High potential for wedge failure
(J2/J3); High potential for planar
failure (J1 & Pitahaya Fault); High
potential for toppling (J1); Moderate
potential for wedge failure (J2/Ja &
J3b)
60
20
8.5
45
Sector 1c
(300° - 320°)
Moderate potential for wedge failure
(J1/J2 & Ja); Low potential for
toppling (J3b) and planar failure (J3a)
65
20
8.5
48
BFA = Bench Face Angle.
IRA = Inter Ramp Angle.
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Sector 2 (320° - 140°) – HW Domain
The HW domain consists predominantly of Andesite HW, classified as “Poor” rock. The pit walls
in the HW domain are relatively shallow (<50 m). Structural controls include the Pitahaya Fault,
Aguila 3 Fault (Aguila 3), and a dominant joint set (J2) with an average dip of 61° and dip direction
of 218°.
To mitigate most instabilities, a BFA of 65° is recommended. Due to the weathered and poor nature
of the HW host rock (excluding ore zones), a 10 m bench height is recommended. This
configuration consists of 10 m vertical intervals, 6.5 m catch benches, and a BFA of 65°, resulting
in a design IRA of 42°. Potential kinematic instabilities are as follows:
•
High potential: Wedge failure (J2/J3; Aguila 3/J3).
•
Moderate potential: Wedge failure (J3/Ja; Aguila 3/J3b; J2/J3 & J3b; Pitahaya
Fault/J3).
•
Low potential: Toppling (J2, J3a); Planar failure (J3a, J3b, J1, J3, Aguila 3).
•
Multi-bench planar failure: Aguila 3 Fault, mitigated by maintaining an IRA below the
fault dip (54°).
Sector 3 (000° - 360°) – Ore Domain
Sector 3 located in the lower benches of the FW and HW domains. The ore domain rock is
classified as “Poor”. This sector is expected to experience similar kinematic failure mechanisms
as the HW and FW domains. To minimize operational risks, a 10 m bench height, 6.5 m catch
benches, and a BFA of 65°, resulting in a design IRA of 42°; is recommended.
Slope Stability Assessment Results
To assess the potential for overall slope failure at the kinematically designed slope angles, stability
analysis was conducted using the limit equilibrium (“LE”) approach using the Rocscience software
package, Slide2. A series of theoretical cross-sections were considered for the assessment, they
were aligned perpendicular to local slope orientation; the slope was adjusted to match the
recommended pit slope geometry. This assessment considered anisotropic strength aligned with
the main joint sets and A 35th percentile (P35) rock mass Hoek-Brown strengths (Table 15.16).
P35 and 95% persistence of anisotropic strengths was used to help mitigate for the lab testing bias
(towards better rock samples), known blocky conditions, and limited structural confidence. Based
on the results it is not anticipated that deep seated slope failure is unlikely to be a concern for the
open pit under dry conditions.
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TABLE 15.16
SUMMARY OF THE SLOPE STABILITY INPUT PARAMETERS
Anisotropic
Parameters*
35th Percentile of RMR76 = GSI, σci and mi, D = 0.7
Domain
Density
(kg/m3)
P35
RMR76 =
GSI
P35
σci
(MPa)
P35
mi
D
mb
s
a
Cohesion
(kPa)
Friction
(°)
Abra Ore Zone
2550
40
84
16.2
0.7
0.6
0.0002
0.5
50
30
Aguilas Ore Zone
2580
46
65
14.7
0.7
0.8
0.0004
0.5
50
30
Andesite FW
2550
57
88
16
0.7
1.5
0.0020
0.5
70
30
Andesite HW
2540
27
74
13.8
0.7
0.2
0.00003
0.5
50
30
Rhyolite Tuff
2460
44
107
12.9
0.7
0.6
0.0003
0.5
50
30
* Anisotropic parameters represent a strength approximated by combining open structure shear strength (0 kPa cohesion and 32 friction) and rock mass strengths
at a ratio consistent with the assumed persistence (95%).
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Open Pit Design Recommendations
•
When mining through faulted ground or ground damaged from underground (UG)
workings (which is expected to occur from the proposed UG mining activities), the
bench height may need to be reduced to 5 m (single benching). The need for this
reduction depends on the pit configuration and wall performance during operations.
•
The near-surface weathered rock mass zone should use half bench height; i.e., either
10 m or 5 m – depending on design sector (with the remaining geometry parameters
unchanged).
•
Where overburden is encountered, it is recommended that slopes be laid back at a
3 H:1 V slope and a 10 m berm width be incorporated between the overburden and rock
slopes. If overburden thickness is greater than 10 m, benching may be necessary.
•
For the Cerro Colorado Pit, there is limited geotechnical data (RQD and some strength
information) available and no geotechnical specific drill holes.
•
The lithology logged in the exploration drill core suggests that the FW material is
slightly different than the FW material in the Abra Eagle Pit shell.
•
There appear to be more broken drill core and fault zones near to the Cerro Colorado
Pit slope.
•
A review of the drill core photos suggests a weaker / more structurally bound rock mass
at the Cerro Colorado Pit.
•
Given the limited information and the small depth of the proposed Cerro Colorado Pit,
a single bench design was assumed for all sectors of this pit (Table 15.7).
•
For the given quality and strength of the rock masses, the maximum stack height is
between 100 – 120 m after which it requires one geotechnical berm (15 m wide) or a
ramp to provide some additional width to shallow the overall slope and also to ensure
safe and efficient operations.
•
The summary of the design recommendations for provided geometries is presented in
Table 15.17 and Figure 15.8.
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TABLE 15.17
SUMMARY OF THE RECOMMENDED DESIGN CONFIGURATIONS
FOR THE PROPOSED PITS
Pit
Abra-Eagle
Cerro-Colorado
Geotechnical
Domain
Wall Dip
Direction
(°)*
Max
BFA
(°)
Bench
Height
(m)
Bench
Width
(m)
Max
IRA
(°)
FW Domain
160 - 300
60
20
8.5
45
HW Domain
320 - 140
65
10
6.5
42
Ore Domain
000 - 360
65
10
6.5
42
All Domains
0 - 360
65
10
6.5
42
* The wall dip direction refers to the location you are facing if you were standing on the crest and looking downslope.
FIGURE 15.8
SUMMARY OF THE RECOMMENDED DESIGN CONFIGURATIONS FOR THE
PROPOSED PIT GEOMETRIES
Conclusions and Considerations for Future Project Stages
•
There is limited information available regarding the Cerro Colorado Pits. The current
recommendations are based on extrapolating the data which have introduced
uncertainties to the study. It is recommended to collect more data in these regions by
drilling a minimum of four oriented boreholes (two per pit) with full geotechnical
logging and laboratory strength testing.
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15.4.3
•
Fault character and orientation should continue to be updated as more data becomes
available. The pit design recommendations and wireframe geometries in these areas
should be periodically reviewed.
•
Based on a review of drill core photos and logs, there is roughly 1 m to 5 m of poor
ground at some lithological contacts, which could create a localized plane of weakness.
As the location of such contacts was not known, it is recommended that GoGold
develop an understanding of where lithological contacts may be in relation to the FW
so that mitigation activities can be planned (if required).
•
Joint persistence and spacing is largely unknown at this time. Once initial benches are
exposed routine geotechnical mapping is recommended to better quantify joint spacing
and persistence. If jointing is widely spaced and low persistence there may be an
opportunity to adjust the pit slope design.
•
The structural data is limited in certain domains and there are some poorly developed
sets that are in the blind zones (±10°-20° from drill hole orientation). It is recommended
to gather additional structural data with oriented core and downhole televiewer surveys
or pit wall mapping during future study stages.
•
It is recommended to keep monitoring ground water levels, based on installed
instruments, to better understand in situ conditions and seasonal fluctuations.
•
Control blasting and scaling of final pit walls are recommended during operations.
Good blasting and scaling practices will minimize rockfalls, thereby minimizing
material accumulation on the benches.
•
During construction and operations routine geotechnical mapping and geological data
collection will be necessary to verify ground conditions. As additional data related to
large-scale structural features and ground conditions becomes available routine design
reviews should be conducted. There may be an opportunity to optimize bench design
once initial benches are exposed and ground conditions are better observed.
•
It is recommended to review the slope design as each bench is mined. Design reviews
should consider observations of ground conditions, and instrumentation data. If interim
benches will be used, it is recommended that bench performance of the interim slopes
and mapping completed on the interim slopes be used to optimize the final pit slope
design.
•
There is no direct shear test result available. It is recommended to perform direct shear
tests on available drill core and review the inputs to the kinematic assessment.
Hydrogeological Studies
No hydrogeological studies have been completed to assess open pit groundwater conditions. The
Abra open pit will have underground workings below it that will likely drain water away from the
open pit. It is anticipated that most of the Abra open pit will be dry as mining progresses.
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15.4.4
Open Pit Mining Dilution and Losses
Dilution and ore losses will occur during mining. It is assumed that some waste rock surrounding
the mineralized zones would be mixed with the ore during mining, thereby causing dilution.
To address dilution, the Mineral Resource block model was regularized from a percent model to a
whole block selective mining unit (“SMU”) model. Regularizing resulted in waste and ore portions
of a block being combined into a single block value. In some cases, the resulting block grade may
be below the applied cut-off grade, in which case that block would be considered an ore loss. In
other cases, while the block grade may decrease due to waste dilution, the entire block remains
above cut-off grade. That would result in an increase in ore tonnage albeit at a lower grade. Hence,
the SMU model is considered a diluted model and was used for open pit production scheduling
and is the basis for the Mineral Reserve Estimate.
The amount of dilution incorporated varies depending on the block size selectivity and the vein
width. For the narrower vein widths at Cerro Colorado, the block size was reduced from 5.0 m to
2.5 m to represent selective mining. Table 15.18 and Table 15.19 compare the undiluted and
diluted tonnes and grade and summarize the expected net dilution in each Deposit. The net dilution
incorporates both the total dilution and ore losses.
TABLE 15.18
ABRA PIT NET DILUTION
Category
Undiluted1
Diluted2
% dilution
Ore
(kt)
2,583.8
2,844.0
10.1
Au
(g/t)
0.87
0.80
8.6
Ag
(g/t)
102
92
9.2
Cu
(%)
0.06
0.05
5.4
AuEq
(g/t)
2.16
1.97
8.9
AgEq
(g/t)
183
166
8.9
AuEq
(g/t)
1.60
1.28
20.3
AgEq
(g/t)
136
109
20.3
Notes: 1. Cut-off is 38 g/t AgEq.
2. Combined dilution and ore loss.
TABLE 15.19
CERRO COLORADO NET DILUTION
Category
Undiluted1
Diluted2
% dilution
Ore
(kt)
427.8
557.5
30.3
Au
(g/t)
1.02
0.81
21.0
Ag
(g/t)
47
39
18.4
Cu
(%)
0.01
0.01
1.7
Notes: 1. Cut-off is 46 g/t AgEq.
2. Combined dilution and ore loss.
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15.4.5
Open Pit Mineral Reserve Estimate
After the open pit designs were completed, the ore tonnage from the diluted block model was
reported inside each pit. The corresponding tonnes and grades are summarized by Mineral Reserve
classification in Table 15.20. The ore tonnages were used as the planning basis for the open pit
production schedule.
The total quantity of open pit ore sent to the process plant is estimated at 2.72 Mt. An average cutoff grade of 46.4 g/t AgEq was applied for all open pits.
TABLE 15.20
OPEN PIT MINERAL RESERVE ESTIMATE (1-6)
Tonnes
(M)
Ag
(g/t)
Au
(g/t)
Cu
(%)
AgEq
(g/t)
Abra Pit
Proven
Probable
Sub-Total
0.58
1.59
2.16
95
84
87
0.72
0.59
0.62
0.02
0.03
0.03
159
137
143
Cerro Colorado Pit
Probable
Sub-Total
0.56
0.56
39
39
0.81
0.81
0.01
0.01
109
109
Combined Pits
Proven
Probable
Total Proven & Probable
0.58
2.14
2.72
95
72
77
0.72
0.64
0.66
0.02
0.03
0.02
159
130
136
Classification
Notes:
1.
Mineral Reserves are based on Measured and Indicated Mineral Resource Classifications only.
2.
Mineral Reserves are reported using the 2014 CIM Definition Standards and 2019 Best Practices Guidelines
and have an effective date of January 14, 2025.
3.
Mineral Reserves are defined within mine plans and incorporate mining dilution and ore losses.
4.
Open Pit Mineral Reserves are based on metal prices of US$23.75/oz Ag, US$1,850/oz Au and US$4.00/lb Cu,
and are constrained within optimized pit shells and designs that use 45–48º overall wall slopes, and process
recoveries of 86% Ag, 95% Au and 51% Cu.
5.
An Open Pit cut-off grade of 46.4 g/t AgEq is estimated to differentiate ore from waste and is based on cost
assumptions of US$26.22/t processing, US$4.11/t site general and administrative, and 0.5% government
mining tax on net revenue. Mining costs are estimated at US$2.10/t of ore and waste rock.
6.
Totals may not sum due to rounding.
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16.0
MINING METHODS
The Los Ricos South Project will consist of both underground and open pit mining operations.
Production will commence with underground mining. Open pit mining will be initiated in Year 10
and there will be an underground / open pit overlap period of four years. The duration of all mining
activity will be 15 years after the start of commercial production.
The life-of-mine ore processing rate is 2,000 tpd or 720,000 tonnes per year (tpa).
Underground mining is described in Section 16.1 and open pit mining is discussed in Section 16.2.
The combined underground and open pit production schedule is presented in Section 16.3.
16.1
UNDERGROUND MINING
Extraction of ore from the underground (“UG”) portion of the Los Ricos South Project will be
through longhole (“LH”) sublevel mining with cemented paste backfill (“PF”). The majority
(~ 70%) of tonnes are extracted from previously unmined mining areas on the Eagle and Abra
Veins, with the remainder of tonnes (~30%) coming from areas adjacent to Historical Mine
Workings (“HMW”) stopes. The vast majority (95%) of ore extracted adjacent to HMW areas is
located in the Old Abra area. The Cerro Colorado Vein was not evaluated for underground mining
for the purposes of this Technical Report. Previous studies (P&E, 2023) indicated insufficient
tonnes of Measured and Indicated Mineral Resource above COG to justify UG mining at Cerro
Colorado. Figure 16.1 shows an isometric view of the underground mine.
Levels have been spaced at nominal 20 m intervals based on geotechnical recommendations from
WSP (2025). Practically, this spacing varies slightly to avoid intersecting HMW development
wherever possible, however, it never exceeds 20 m. The tight spacing provides benefits to
underhand mining, as top-hammer drills are highly effective under these conditions. For overhand
mining, either top-hammer or in-the-hole (“ITH”) drills can be used.
A crown pillar of a minimum 20 m thickness has been recommended by WSP (2025).
Approximately 3% of the UG Mineral Reserve tonnes are located in this pillar comprising 2.3%
of the UG Mineral Reserve AgEq ounces. Some stopes in the crown pillar area will be eventually
extracted utilizing uphole drilling from the underground and backfilling from surface, and some
stopes are free-flowing HMW areas that do not require blasting or overcut access. The average
grade of ore in the crown pillar is approximately 235 g/t AgEq. The crown pillar generally affects
the Abra Vein to a greater extent than the Eagle Vein.
Due to proximity to surface in conjunction with crown pillar limitations and difficult extraction
geometries, some Mineral Resource that supports underground methods exists within the open pit
area. The few UG stopes that intersect the pit boundary will be extracted and filled with highstrength PF. It is expected that the open pit will eventually expose this backfill in localized portions
of the walls and floor.
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FIGURE 16.1
UNDERGROUND MINE ISOMETRIC VIEW
Source: P&E (2025)
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The underground has been nominally subdivided into three mining areas:
•
Eagle:
Stopes and development on the northern side of the Project located to the
east of the Rascadero Fault;
•
New Abra: Stopes and development on the northern side of the Abra Vein, located to
the west of the Rascadero Fault; and
•
Old Abra: Stopes and development on the southern side of the Abra Vein, located to
the west of the Rascadero Fault and generally contiguous to the majority of HMW
areas.
The division between New and Old Abra is based on convenience: there is no geological difference
between the two areas. Stopes on the Abra Vein to the north of the main HMW areas are generally
assigned to one area or the other based on location relative to the access drift. Figure 16.2 shows
the three mining areas. New Abra and Eagle overlap in cross-section view over a distance of 350
to 400 m. In previous studies the two areas were accessed from the same ramp: to avoid developing
through the Rascadero Fault; and provide improved mining productivity, these areas are now
accessed separately.
Mining is performed by overhand (progressing upwards) and underhand (progressing downwards)
methods at the Project. Most tonnes are extracted using overhand methods: the majority of these
tonnes are located in the Eagle and New Abra mining areas. In general, underhand mining is
restricted to areas contiguous with HMW voids or the extraction of the top levels of overhand
mining blocks underneath artificial pillars. Therefore, the majority (79%) of tonnes from the Old
Abra area are extracted with underhand methods, whereas the minority of Eagle (10%) and New
Abra (17%) tonnes are extracted using those methods.
In overhand mining areas, the stopes are segregated into a series of mining “blocks”, where mining
progresses from the bottom of the block to the top. The position and extents of the blocks are
largely a function of scheduling, and the blocks vary significantly in lateral and vertical extents.
The purpose of the blocks is to allow overhand mining without requiring development to progress
to the deepest extents of the mine before production commences.
All stopes on the lowest level in each block (excluding the very lowest blocks in the mine, or
stopes with no mining planned below them) will be filled with high-strength PF to facilitate
eventual undermining by future stopes. Where no stopes are planned underneath, lower-strength
PF can be used. Drilling for most stopes (depending on geometry) will be downhole, improving
accuracy and recovery compared to uphole drilling. When mining commences on the top level of
a block, upholes will be drilled to the PF contact, after which the stope will be blasted and
extracted. Filling for these stopes will require drilling holes from the undercut to the top of the
stope void through which PF will be pumped. Extraction proceeds on retreat, repeating the process
until the top of the block is complete. Lower recoveries are applied to these areas versus normal
downhole overhand mining.
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FIGURE 16.2
THREE UNDERGROUND MINING ZONES
Source: P&E (2025)
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For underhand mining in Old Abra, mining blocks are not particularly relevant, as the progression
of all stopes around the HMW is top-down due to the need to remove free-flowing material and
provide an engineered fill to excavate around. As portions of the Old Abra extend to the northwest
of the HMW, overhand mining and underhand mining can be present in the same nominal block,
however, sequencing is used to ensure that the two methods maintain stable pillars (initially insitu rib or sill, but eventually artificial PF) between them to prevent negative interactions.
Both uphole and downhole drilling are used to drill the stopes. In general, downholes are preferred
due to ease of QA/QC and explosive loading, however, the relatively tight level spacing (nominally
20 m) improves the efficacy of uphole drilling. Uphole and downhole drilling make up roughly
equivalent components of the planned stopes (46% uphole, 49% downhole, with the rest of the
tonnes made up of free-flowing HMW areas).
Two methods of access are used to extract the stopes: longitudinal and transverse mining.
Longitudinal mining uses sills driven down the strike of the ore, with stope extraction starting at
the far end of a sill and retreating to the start. Transverse mining is utilized where the thickness
of the ore lens is wider than a nominal 15 m, or where multiple lenses are encountered in close
proximity and improved productivity is desired. Overhand and underhand methods can be utilized
with either method of access. In general, transverse mining is most common in the Old Abra area,
however, it is present in significant quantities in the Eagle area, and in a small minority of New
Abra. Overall, transverse access stopes make up 62% of all mining areas. Figure 16.3 shows the
distribution of transverse and longitudinal mining in the underground.
Initial mine development will prioritize development in the Eagle area, with additional
development directed towards the New Abra area. Total mine lateral development is 60.9 km
lateral (64.8 km of 4.0 mW x 4.0 mH equivalent) and 1.9 km vertical (1.9 km of 3.0 mW x 3.0 mH
equivalent). Early mine production is predominately in the Eagle Vein, where the best ground
conditions and highest grades are expected. Over the first 5 years of the mine plan (2 years preproduction and 3 years production), Eagle makes up slightly more than 60% of mined tonnes and
70% of AgEq contained ounces. New Abra and Old Abra are then brought online when the number
of available faces at Eagle decreases as transverse mining areas there are exhausted. The output
of the Eagle and New Abra areas decrease until they are exhausted near the end of YR 10, at which
point only the deepest parts of Old Abra are still active, and the open pit is contributing the majority
of ore to the process plant. Underground mine production ceases at the end of YR 13.
Over the life-of-mine (“LOM”), a total of 7.51 Mt of ore will be recovered from the UG, containing
41.1 Moz Ag, 399 koz Au and 10.0 kt Cu, equivalent to 78.8 Moz AgEq. Table 16.1 shows the
breakdown of mined tonnes and metal by mining location over LOM from the underground.
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FIGURE 16.3
LONGITUDINAL PROJECTION SHOWING STOPE TYPES
Source: P&E (2025)
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TABLE 16.1
LOM TOTALS BY MINING AREA
Mining Area
Eagle
New Abra
Old Abra
Total1
1
Mined Mass
(kt)
2,541
1,315
3,656
7,512
Mined AgEq
(Moz)
32.1
13.2
33.5
78.8
Mined Ag
(Moz)
13.7
7.1
20.3
41.1
Mined Au
(koz)
188.6
62.3
148.3
399.2
Mined Cu
(kt)
6.4
2.5
1.2
10.0
Totals may not sum due to rounding.
All UG mining will be performed using contract labour and equipment, with Company technical
services and management. A request for proposals was issued to selected mining contractors in
Mexico in December 2023, with three qualified responses received in Q1 2024.
16.1.1
16.1.1.1
Design Methodology
Modified Hill of Value Analysis
During initial stope design, an array of scenarios was evaluated to produce a plot of the estimated
economic impacts of varying Hill of Value (“HOV”) parameters as shown in Table 16.2.
TABLE 16.2
MODIFIED HOV ANALYSIS PARAMETERS
Item
Level Spacing
Cut-Off Grade in AgEq
Stope Length
Units
m
g/t
m
Eagle
20, 30
New Abra
Old Abra
20, 30
20
150 to 300 in increments of 30
10, 15, 20
15
Each mining area (Eagle, New Abra, Old Abra) was modelled separately and allowed to vary
parameters independent of the other areas. The combination of these data sets produced 1,944
data points. A mining plan targeting Eagle, New Abra and Old Abra in sequence was then
automatically generated to produce a simplified cashflow model, and correlations between varying
parameters were established. The most significant correlation was between increased COG in
Eagle versus improved IRR. Optimizing the other parameters indicated that the best scenarios
occurred when the Eagle COG was 210 g/t AgEq, while the selection of COG for the New Abra
and Old Abra Veins had little overall impact on the results. In addition, it was found that utilizing
a 20 m H x 15 m L stope in the Abra Vein produced the best outcome relative to all other stope
geometries. Varying the geometry in Eagle had little overall impact, and the peak results occurred
when the same stope geometries were used in Eagle as in Abra, which simplifies the overall design
and operation of the mine.
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16.1.1.2
Contractor Versus Owner
Contractor mining was selected to reduce initial capital costs of acquiring a mining fleet, and to
allow flexibility in yearly development requirements: when peak development rates of ~6.4 km/yr
are no longer required after YR 5 there is no need to retrain or lay off Company personnel.
Additionally, utilizing contractors generally provides more experienced personnel at reduced effort
versus hiring personnel directly by the Company. It is envisioned that the Company will provide
all necessary supervisory, technical, management and support personnel to assist, oversee and
direct the contractor operations with a high degree of QA and QC. All actual mining operations
(excluding surveying and grade control) will be directly performed by the contractor.
16.1.1.3
Historical Mine Workings
Significant HMW exist at Los Ricos South, primarily within the Abra Vein, and to lesser extents
in the Eagle Vein. Utilizing historical mine schematics and production reports from the early
1900s, a 3-D primitive solid was created by GoGold to approximate the extents of the underground
workings for the 2021 and 2023 PEA reports on the Project. This solid was updated by GoGold
with drill hole void and backfill intercepts to improve spatial accuracy for this Technical Report.
In addition, a campaign of LIDAR surveys by drone was undertaken in May 2024 to improve the
accuracy of surveys in accessible workings, namely the upper works (“Windows”) areas of the
Abra Vein and the “El Troze” portal and workings on the Eagle Vein. Additional surveys were
undertaken at other known historical adits around the Property, however, these areas do not
intersect the planned mining areas for the Project. Historical reports indicate there are additional
unsurveyed exploration works on the Eagle Vein, with no significant historical mining in the area.
Figure 16.4 shows the current estimated extents of the HMW for the Project.
The HMW stoping areas in Abra are assumed to be fully flooded below the 1,150 m EL, and full
of free-flowing material at an average grade of 70 g/t AgEq and an average density of 0.81 t/m3.
The El Troze area of Eagle, having been surveyed and shown to be an open void, is treated
accordingly. HMW development areas are assumed to be open void.
Extraction of the Abra Vein requires significant mining through the HMW stopes, whereas
extraction of the Eagle Vein requires significantly less interaction with HMW areas. Additional
information on the extraction plan for these areas is provided below in Section 16.1.2.55.
Increased mining loss is planned around HMW areas, as is increased dilution. The majority of
mining around HMW areas utilizes upholes and underhand mining, with a minority using overhand
mining where the extents of the intersected HMW areas are minimal. HMW areas generally range
from 3 to 5 m in width, however, areas can be locally wider, or can have multiple parallel openings
on a level, particularly around structural features.
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FIGURE 16.4
LONGITUDINAL PROJECTION SHOWING HISTORICAL MINE WORKINGS
Source: P&E (2025)
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16.1.1.4
Stope Design
Deswik Stope Optimizer (“DSO”) was utilized to generate stope sets for various combinations of
parameters. These parameters were utilized in a modified HOV analysis to determine the
parameters for the final stope set utilized for the mine plan. This final set of parameters is shown
in Table 16.3.
TABLE 16.3
LOM TOTALS BY MINING AREA
Item
Stope Height
Stope Length Along Strike
Minimum Mining Width
Minimum Waste Pillar
ELOS (total HW + FW)1
FW Dip Angle2
HW Dip Angle
Units
m
m
m
m
m
degrees
degrees
Azimuth Limits
degrees
Dip and Strike Source
Points per Ring
Sub-Shapes3
Cut-off Grades4
No.
g/t AgEq
Eagle
New Abra
Old Abra
20
15
2.4
10
1.0
1.2
1.2
65° ± 15°
90° ± 40°
90° ± 35°
30° stope-to-stope variance
±45° max variance from model strike (40° NW)
Vein wireframe
6
5 m increments on height and length along strike
210
150
Notes:
1
Ground conditions are expected to be better on the Eagle Vein, with reduced overbreak on HW versus the Abra
Vein. ELOS = equivalent linear overbreak slough
2
Due to the impacts of the HMW voids, FW dip for shapes on the Abra Vein can vary significantly.
3
The most economic individual, or combination of, sub-shapes was selected from 13 generated sub-sets.
4
See Section 16.1.1.5 for further detail.
Access method (transverse versus longitudinal) is determined by a combination of factors, most
prominent of which is thickness (HW to FW) of the stope. The access method was not known until
after the DSO work was completed, therefore the stope length along strike for DSO was limited to
the expected crosscut spacing of transverse accesses. In practice, stope lengths for longitudinal
access stopes will vary slightly based on local ground conditions.
The longitudinal method has been planned for stopes up to 15 m wide, and transverse for greater
than 15 m. Where HW to FW spans of greater than 15 m are found in isolated areas otherwise
mined by longitudinal retreat, two parallel longitudinal drifts are utilized to limit the HW to FW
span of any individual stope to 15 m or less. Extraction takes place on the FW side of the stope
first, and then on the HW side after the FW is backfilled and cured. This occurs only in isolated
portions of the Eagle Vein.
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Partial-height stopes will be extracted using uphole drilling from the undercut, or by downhole
drilling from the overcut and blasting only a portion of the hole before backfilling. Partial-height
stopes are utilized only where they are contiguous with the excavation horizon (undercut).
Finally, a manual revision of stope shapes was done where geometries and thicknesses generated
by DSO were not conducive to real-world mining around the HMW areas. Additionally, manual
shapes at true marginal COG (“MCOG”) of 130 g/t AgEq were generated where gaps appeared
along strike and development was needed to reach stopes further down strike (e.g. where
development cost would not be applicable to the stope since it was already sunk in reaching stopes
further along). No manual stopes below MCOG were generated or included in the mine plan.
16.1.1.5
Cut-off Grades
Different COGs were used in different areas of the mine to achieve the economic goals of the
Project. While the actual MCOG was determined to be 130 g/t AgEq, a contingency was added
and 150 g/t AgEq was utilized. Additionally, as increased COG in the Eagle area was correlated
with improved Project economics, and minimal correlation was seen between COG in the New or
Old Abra areas, a middle-ground set of stopes at 180 g/t AgEq was generated and used to replace
stopes generated at 150 g/t AgEq wherever the two sets overlapped to generate a higher margin
per tonne.
As mentioned previously, a manual set of stopes at 130 g/t AgEq was generated to infill gaps along
strike where development costs were already sunk in reaching stopes further along strike. These
marginal-value tonnes total less than 0.3% of the Mineral Reserve tonnes. Finally, a manual
revision of stope shapes around the HMW voids to improve extraction resulted in a series of shapes
with no specific COG. DSO shapes generated at 150 g/t AgEq adjacent to HMW areas (no specific
COG, with nominally 80 g/t AgEq) were combined to generate a single mining shape comprised
of two zones with no specific COG.
16.1.1.6
Production Rate
Production rate for the mine was set using Long’s Modification to Taylor’s Rule, allowing the UG
mine to set the process plant rate. The formula used is:
𝑅𝑎𝑡𝑒𝑡𝑝𝑑 = 0.297 ∗ 𝑀𝑖𝑛𝑒𝑎𝑏𝑙𝑒 𝑇𝑜𝑛𝑛𝑒𝑠 0.562
Initial modified HOV work indicated a mine plan tonnage of approximately 6.4 Mt in-situ, plus an
estimated 700 kt of HMW material for a total of 7.1 Mt. Eventual technical revisions resulted in
an increase to mined tonnage of ~300 kt for a final Mineral Reserve of 7.5 Mt, which yields a
theoretical production rate of 2.2 ktpd. A 2.0 ktpd rate was utilized for this study to allow
conservatism on reliance of mining rates around the HMW.
Initial stope productivity estimates were performed, generating estimated production rates for
transverse and longitudinal stopes of varying geometries. These were utilized to confirm the target
production rate for the mine through manual scheduling. For the final mine plan, however,
Deswik.Sched was utilized to generate individual derived tasks for each individual stope shape for
the following items:
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•
•
•
•
•
•
•
•
Drilling raise (either Dice-5 or V30);
Drilling slot;
Drilling stope;
Loading stope;
Excavating stope;
Installing PF barricade;
Filling stope with PF; and
Curing PF.
In addition to the above activities, a further 4 days delay is incurred for the following operations
(1 day each):
•
•
•
•
Prepping stope area;
Loading and blasting raise;
Loading and blasting slot; and
Surveying stope.
These activities result in variable production rates per stope due to variable geometries, however,
on average the mine requires roughly 10 active stopes to support the 2 ktpd production rate.
16.1.1.7
Fleet Sizing and Selection
The fleet was selected to achieve the following goals:
•
•
•
•
•
•
3-pass loading of trucks in level accesses;
Any load-haul-dump (“LHD”) equipment can be used to load any truck at any loadout,
therefore an LHD excavating a stope can be repurposed to load a truck at the level
access when the remuck bay is full;
Mechanized installation of ground support tendons up to 2.4 m in length;
Rapid ramp development using twin-boom jumbos;
Minimizing spare parts warehousing and specialized training; and
Minimized development profile.
Two possible avenues were identified to achieve the above:
1. 7-t class LHDs with small LH drills and 30 t trucks using 3.5 mW x 3.5 mH sills and a
4.0 mW x 4.0 mH ramp; and
2. 10-t class LHDs with mid-sized LH drills and 30 t trucks using 4.0 mW x 4.0 mH sills
and a 4.5 mW x 4.0 mH ramp.
Option 2 was selected, since Option 1 did not provide a significantly smaller cross-sectional drift
area, required the drills to operate at the high end of the operating envelope, and did not provide
the option of expanding level spacing in Eagle while retaining the same equipment if favourable
conditions are encountered after the start of mining.
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The contractor will provide equipment comparable to Sandvik units as shown in Table 16.4.
TABLE 16.4
COMPARABLE EQUIPMENT TYPES
Type
LHD
Truck
Development Jumbo
Development Bolter
ITH LH Drill
Top-Hammer LH Drill
Cable Bolter
Service Unit
Comparable Unit
LH410
TH430
DD421
DS312 w/ SBH81
DU412 w/ V302
DL321
DS421
Non-Sandvik3
Notes:
1
Bolting head for 8 ft (2.4 m) rebar, hydro-bolts or split-sets.
2
The 30 in (0.762 m) reamer head can be installed for raise drilling, particularly uphole raises in Old Abra.
3
Several manufacturers produce lines of vehicles suitable for use: the contractor may select one at their discretion.
16.1.1.8
Backfill
PF was selected for underground backfill since longitudinal retreat mining results in poor
geometries for filling when trackless equipment is utilized. There is no turn-around or re-muck
bay infrastructure within the access drift, since it is driven in ore along strike of the vein. Long
hauls for LHD units along drifts that follow changes in vein azimuth result in poor productivities
for rock or aggregate fill, with a pumpable fill being a better choice. Additionally, pumpable fill
will better seal any connections between old development workings and stopes, limiting water
inflow into stopes below and reducing ventilation short-circuiting. Finally, surface storage limits
necessitate the return of a significant portion of mine tailings back into the underground. Based
on availability of suitable fill components (water, tailings, waste rock material source), PF was
selected as the optimal fill for the underground.
Several different PF recipes will be in common use in the UG. A high-strength recipe will be used
for areas requiring eventual undermining (artificial sill pillars), while a lower-strength recipe will
be used for areas that will only be exposed laterally. A minimum-strength recipe will be used
where tailings emplacement is the primary goal of PF, rather than excavation stability.
16.1.1.9
Water Inflows
Little is known about the hydrogeological conditions that will be experienced underground.
Historical evidence suggests minimal water inflow rates that can easily be managed with pumping,
since HMW areas were excavated over a century ago using methods and technologies of the era.
Areas that can currently be accessed (El Troze, the Windows, etc.) are all dry. The Farmyard
portal area is currently covered by a small agricultural water retention dam with minimal flow (see
Figure 16.5), however, it is not known if there are other drainage channels out of the HMW areas.
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FIGURE 16.5
FARMYARD PORTAL WATER RETENTION DAM
Source: P&E (2024)
Estimates of the flow at the Farmyard portal are 2.5 L/s. Machine service water is calculated at
4.8 L/s, while diesel condensation and leakage are calculated at 0.2 L/s for a total of 7.5 L/s inflow.
The HMW voids are assumed to be saturated and will require dewatering before mining.
16.1.2
16.1.2.1
Geological and Geotechnical Considerations
WSP Study and Recommendations
WSP provided level spacing, stope sizing and footwall infrastructure offset recommendations
based on a drilling program completed to support the FS. An Executive Summary of the WSP
report titled “Underground Geotechnical Recommendations – Los Ricos South Project” dated
February 4, 2025 is provided below.
The geometry of GoGold’s Los Ricos South ore body is long, narrow and subvertical, and ranges
in thickness (hanging wall to footwall) from approximately 5 to 35 m. The dip of the mineralization
is approximately 65 to 70 degrees with a dip towards the southeast. The ore bodies are planned to
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be mined by both open pit and underground methods with the open pit following the underground
mining.
The geotechnical and hydrogeological investigation program used to support this study (WSP,
2025) was originally planned around a different underground and open pit layout, meaning there
are areas in the underground mine layout where no geotechnical dedicated drilling has yet been
completed (Figure 16.6). The field investigation that was completed was carried out during the last
quarter of 2023, and consisted of:
•
Drilling of 17 geotechnical holes with a total length of 4,069.5 m;
•
Geotechnical drill core logging including description and orientation of discontinuities;
•
Downstage hydraulic conductivity testing (packer testing) at varying intervals in three
drill holes and Vibrating Wire Piezometer (“VWP”) installation into those same drill
holes;
•
Collection of rock samples for laboratory testing; and
•
Point Load Testing (“PLT”) along every drill hole.
To supplement the field investigation data and provide some data in areas where geotechnical
drilling was not available, photo logging of specific intervals in 19 exploration drill holes was
completed.
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FIGURE 16.6
VIEW LOOKING NORTHEAST OF PLANNED UNDERGROUND DEVELOPMENT SHOWING AREAS OF THE MINE
THAT HAVE NO DATA FROM THE GEOTECHNICAL INVESTIGATION
Source: WSP (2025)
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Engineering Geology Model (EGM) Conclusions
The logged lithologies were combined into geotechnical domains based on lithologies that exhibit
similar rock mass characteristics. The domains in Table 16.5 were established.
TABLE 16.5
SUMMARY OF THE LOGGED LITHOLOGIES COMBINED INTO THEIR RESPECTIVE
DOMAINS
Domain¹
Logged
Lithologies
Metres Logged
(m)
Andesite Footwall (FW)
Andesite (from FW), Andesitic Tuff (from
FW), Andesitic Lithic Tuff
1,577
Andesite
(HW)
Andesite (from HW and ore zone (OZ)),
Andesitic Tuff (from HW and OZ), Aphanitic
Andesitic Tuff
736
Rhyolite Tuff
Rhyolite Tuff
640
Porphyritic Andesite
Porphyritic Andesite
320
Abra Ore Zone
Los Ricos Vein
165
Abra Ore Zone Breccia
Tectonic Breccia (from Abra Zone),
Hydrothermal Breccia (from Abra Zone),
Mineralized Breccia (from Abra Zone)
45
Aguilas Ore Zone
El Aguila Vein
32
Aguilas Ore Zone Breccia
Tectonic Breccia (from Aguilas Zone),
Hydrothermal Breccia (from Aguilas Zone),
Mineralized Breccia (from Aguilas Zone)
88
Hanging
Wall
¹ The following lithologies were excluded due to a lack of drilling data or because they fall outside the underground
design area: Diorite (68 m), Andesite Dyke (12 m), Volcano-sedimentary (2.19 m), Calcite Vein (1.07 m)
•
Characterization of the rock mass fabric was evaluated from structural data collected
in the geotechnical drill holes. Three major and two minor discontinuity sets were
identified including a set perpendicular to the ore body and a set parallel to the
orientation of the ore body (Table 16.6). The third major set (J3A and J3B) is less
prominent than J1 and J2 but present in most lithologies and aligns with local faults.
The main structural feature is the parallel joint set (J1) with general orientation of 81°
dip and 033° dip direction; this joint set appears to be dominant throughout the Project
site.
•
Collected geotechnical data indicated that facture intensity in the FW zone is less than
in the HW zone. The HW zone had a high degree of faulting and broken drill core zones
than the ore zone and FW domain. Due to this, drilling orientation data was difficult to
collect in this area.
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•
Core samples of different domains were strength tested in the laboratory. Hoek-Brown
failure criteria were developed for each geotechnical domain (Table 16.7). Highly
altered or fractured units, although present on the Project site are underrepresented due
to the inability to collect specimens or conduct valid tests in a laboratory.
•
Rock mass quality was evaluated using the NGI-Q and RMR76 rock mass classification
systems (Table 16.8).
•
Hydrogeologic data, collected primarily in the more competent FW rock masses,
suggest a hydraulic conductivity ranging from 1 x10-7 to 1 x10-10 m/s representing
relatively low conductivity conditions. The vibrating wire piezometers have indicated
that the static water level is approximately at 1,160 m above sea level (“masl”) in the
center and north areas of the mine, while is it slightly higher (~1,190 masl) at the south
edge of the mine.
•
No in situ stress data has been collected and was assumed based on the nearest data in
the World Stress Map to be:
ơv = γ * z
ơ1 = KH1 * ơv = 1.5 * ơv
ơ2 = KH2 * ơv = 1.3 * ơv
•
Based on the EGM, the types of rock mass behaviour at Los Ricos South that is
assumed to be encountered during mining activities and form the basis for some of the
rock engineering assessments are:
o Unravelling of Poor ground (e.g., fault zones, ore zones, HW rock masses, etc.) in
low to zero confinement areas;
o Gravity driven block failure in Fair to Good ground conditions under low
confinement; and
o Induced stress damage to areas of Good rock character near to the excavated stopes
(this is likely to be limited and mostly focused in the FW rock masses).
TABLE 16.6
SUMMARY OF SELECTED STRUCTURAL SETS
Set
J1
J2
J3A
J3B
Ja
Jb
Dip
(°)
81
61
65
76
63
63
Dip Direction
(°)
033
218
275
099
320
146
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TABLE 16.7
SUMMARY OF HOEK-BROWN INTACT ROCK STRENGTH ENVELOPES FOR THE
LITHOLOGICAL UNITS AT PROPOSED PROJECT UNDERGROUND MINE
Domain
[No. of Intact Tests]
Andesite FW [40]
Andesite HW [6]
Aguilas Ore Zone [9]
Abra Ore Zone [13]
Rhyolite Tuff [13]
σci
(MPa)
100
94
74
91
115
mi
r2
21.4
14.7
12.2
16.9
14.5
0.61
0.65
0.33
0.82
0.76
*’Intact Tests’ refers to a strength sample which failed through intact rock and was not influenced by sample defects
or planes of weakness.
TABLE 16.8
SUMMARY OF RMR76 MEAN AND CALCULATED Q' PARAMETERS
BY GEOTECHNICAL DOMAIN
Q’
Mean RMR76
Andesite FW
10-14
59
Rhyolite Tuff
6-13
55
Porphyritic Andesite
16
59
Main Ore Zone – Above 200 m
1.6-2.7
46
Main Ore Zone – Below 200 m
(Based on Photo Logging)
5-9
64
Andesite HW – Above 200 m
1-1.6
39
Andesite HW – Below 200 m
(Based on Photo Logging)
3.5-6
54
Aguilas Ore Zone
5-12
55
Aguilas Ore Zone Breccia
3-5
46
Main Ore Zone Breccia
5-12
48
Lithology
Note: Design RMR76 Colored by Classification Number: Red (Very Poor Rock [<20]), Orange (Poor Rock [20 - 40]),
Yellow (Fair Rock [40 – 60]), Light Green (Good Rock [60 – 80]), and Green (Very Good Rock [80 – 100]).
Q’ Colored by Classification Number: Dark Red (Extremely Poor [<0.01]), Red (Very Poor [0.01 – 0.1]),
Orange (Poor [0.1 – 1]), Yellow (Fair [1 – 4]), Light Green (Good [4 – 10]), Dark Green (Very Good [10 –
40]), and Grey (Extremely Good [>40]).
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Geotechnical and Hydrogeological Data Coverage and Potential Gaps
The following potential gaps in the geotechnical and hydrogeological data coverage have been
identified during development of the EGM:
•
No geotechnical drilling and hydrogeological testing was completed below a depth of
approximately 200 m (~1,050 masl).
o Data below 1,050 masl is limited to photo logging of exploration drill core and
available PLT data from exploration core.
•
Some zones above 1,050 masl had limited or no geotechnical data collection.
•
Orientation of discontinuities from drill core for interpretation of structural fabric was
possible over approximately 50% of the geotechnical drilling meterage due to the
generally low rock mass quality.
o No downhole televiewer surveys were attempted due to low rock mass quality and
high likelihood of drill hole blockage.
o The structural interpretation of joint sets was based on limited pole concentrations
for certain apparent structural orientations, resulting in some assumptions being
applied to how the data was interpreted and used.
•
No in situ stress testing or publicly available in situ stress information is available for
the Project site and region.
•
RQD logging data from the geotechnical drill holes indicate there may be some bias
involved in the data collection as each run had either >80% RQD or <20% RQD with
little cases of runs being in between.
Stope Dimensions and Dilution Estimates
Stable spans for open stopes can be estimated based on an empirical stability graph, developed by
Mathews et al. (1981) and updated by Potvin (1988) amongst others. Stope dimensions were
assessed by geotechnical domain composing the HW, FW, stope back and stope end walls and the
stopes for Abra and Aguilas Zones. Depth range (above or below 200 m depth) and hanging wall
dip (55-70° or >70°) were also included in the assessment. The range of rock mass quality in the
mineralized zones, ore body HW and FW varies from Poor (Q’=1.0) to Good (Q’=12) ground. The
recommended HW design was based on Potvin (1988) unsupported curve in the stability graph.
The empirical results indicated that stope HW, FW, stope backs and stope end walls are forecasted
to be stable unsupported for 20 mH x 15 mL stope dimensions (for spans up to 15 m). Larger stope
sizes were also found to be stable unsupported in areas with stronger rock masses; however, the
Authors decided to apply a single stope size of 20 mH x 15 mL (for spans up to 15 m) across the
Project site.
An estimate of HW dilution was completed using the same stope size on the Clark and Pakalnis
(1998) Equivalent Liner Overbreak/Slough (“ELOS”) chart. The results indicate:
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•
0.5 m to 1.0 m of overbreak for stope HW located in the Aguilas ore zone domains or
the Abra ore zone domain below 200 m depth (Q’=5-12) with a minimum ore zone
thickness of 5 m between the stope HW and the HW rock masses; and
•
1.0 m to >2.0 m of overbreak for stope hanging walls located in Poor quality rock
masses (Andesite HW domain or above 200 m in the Abra ore body domain (Q’=1.02.7))
Stope Sequence, Stand Off Distances and Infrastructure Stability
2-D (Rocscience RS2) and 3-D (MAP3D) finite and boundary element numerical models were
developed to assess the stand off distances of the permanent infrastructure so that they are placed
outside of the influence of stresses induced by stope extraction. Infrastructure (including ramps,
vertical development and other critical infrastructure) are required to be placed a minimum of 40 m
away (laterally) from the ore body, in competent rock.
Vertical Development
Based on the McCracken and Stacey (1989) method and kinematic assessments, vertical
developments of 3.4 m diameter can be unsupported in Fair to Good ground (Q >4). Such vertical
development in Poor ground (1< Q <4) require 1.5 m Swellex (PM12) on a 1.5 m x 1.5 m spacing
in the walls as well as 50 mm of fiber reinforced shotcrete. Vertical develop should not be placed
in ground with a Q <1 (Very Poor ground). Escape ways and surface ventilation shafts should
follow the ground support recommendations for the Poor ground conditions listed above; even
when constructed in Good quality ground.
Crown Pillar Stability Assessment
•
Various effective crown pillar spans were assessed using empirical and numerical
methods, and the results indicated that a crown pillar minimum thickness of 20 m is
required for spans less than or equal to 6 m and a minimum thickness of 30 m for stopes
up to 15 m span.
•
Stopes with spans larger than 6 m and a crown pillar thickness less than 30 m should
be divided into separate backfilled stope panels.
Standard Development Ground Support
Ground support recommendations (Table 16.9) were provided based on the type of excavation,
specifically permanent (ramps, ore drifts, underground shops, etc.) or temporary openings (e.g.,
access drifts), as well as by the quality of the rock mass. For Poor to Good ground conditions,
recommendations were developed using empirical charts (Barton & Grimstad, 1993 and Potvin &
Hadjigeorgiou, 2016), kinematic assessments and support length reviewed against depth of yield
modelled with 2-D finite elements (Rocscience RS2) considering induced stress changes estimated
in 3-D boundary element numerical assessments (MAP3D). A different approach was used for
Very Poor-quality rock masses where soil like behavior is expected and for development through
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cemented paste fill. For both cases, depth of loosening was estimated using Terzaghi’s rock load
methodology (Proctor and White, 1946) to assess the load on the ground support elements.
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TABLE 16.9
GROUND SUPPORT SELECTION
Support
Area
Small
CAPEX
Permanent
Development
(FWD's,
vents)
Rock Quality
Exceptionally
Poor to Very
Poor (Q<1)
Poor (Q=1-4)
Fair to Good
(Q>4)
Exceptionally
Poor to Very
Poor (Q<1)
Large
CAPEX
Permanent
Development Poor (Q=1-4)
(FWD'S &
Fair to Good
Ramp)
(Q>4)
Exceptionally
Poor to Very
Poor (Q<1)
Vertical
Poor (Q=1-4)
Development
Fair to Good
(Q>4)
Support
Application
Area
Back
Walls
Back
Walls
Back
Walls
Surface Support
Support
Element
Max.
Spacing
(m)
Min.
Length
(m)
1.2 × 1.2
2.4
1.2 × 1.2
2.0 × 2.0
1.2 × 1.2
2.4
1.8
1.5
1.2 × 1.2
2.4
1.2 × 1.2
1.2 × 1.2
1.5 × 1.5
1.2 × 1.2
2.4
2.1
1.8
1.8
1.5 × 1.5
1.5
&
Primary
Back
Back
Walls
Back
Walls
Support
Type
(Primary/
Secondary)
Rockbolt
Pattern
and
Specifications
Resin
Grouted
Rebar
&
Primary
Resin
Grouted
Rebar
Inclination
Normal to
Opening
Welded
Wire
Mesh
Gauge
(mm)
NA
Fiber
Reinforced
Shotcrete
Thickness
(mm)
100
50
8 - Back &
Shoulders
NA
100
Normal to
Opening
NA
50
8 - Back &
Shoulders
NA
Not recommended
Walls
Primary
Swellex
(PM12)
No
support
required
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Normal to
Opening
50
NA
NA
NA
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NA
NA
TABLE 16.9
GROUND SUPPORT SELECTION
Support
Area
Rock Quality
Support
Application
Area
Exceptionally
Back
Poor to Very
Walls
Poor (Q<1)
OPEX
Back
Development Poor (Q=1-4) Walls
through
Virgin Rock
Back
Fair to Good
(Q>4)
Walls
OPEX
Development Good Quality
through
Pastefill
pastefill
Exceptionally
Poor to Poor
Underground (Q<4)
Shop
Fair to Good
(Q>4)
Support
Type
(Primary/
Secondary)
Rockbolt
Pattern
and
Specifications
Surface Support
Support
Element
Max.
Spacing
(m)
Min.
Length
(m)
1.2 × 1.2
2.4
1.2 × 1.2
2.4
1.2 × 1.2
1.8
&
&
Primary
Swellex
(PM12)
Inclination
1.5 × 1.5
1.5
1.2 × 1.2
2.1
Normal to
Opening
Welded
Wire
Mesh
Gauge
(mm)
Fiber
Reinforced
Shotcrete
Thickness
(mm)
NA
100
NA
50
8 - Back &
Shoulders
NA
8
NA
Ribs
Back
Primary
Swellex
(PM12)
Normal to
Opening
The shop is assumed to be constructed in Ground Quality with a Q>4
Back
Walls
Passing/Load Poor to Good
Back
out Bays
(Q>1)
Primary
Secondary
Resin
Grouted
Rebar
Cable
Bolts
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1.5 × 1.5
2.4
1.2 × 1.2
2.1
1.5 × 1.5
4.5
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Normal to 8 - Back &
Opening
Shoulders
75
Normal to 8 - Back &
Opening
Shoulders
NA
TABLE 16.9
GROUND SUPPORT SELECTION
Support
Area
Rock Quality
Support
Application
Area
Support
Type
(Primary/
Secondary)
Rockbolt
Pattern
and
Specifications
Surface Support
Support
Element
Max.
Spacing
(m)
Inclination
Min.
Length
(m)
Exceptionally
Poor to Very Not recommended to design portals in Ground Quality with a Q <1
Poor (Q<1)
Portal
Entrance
1.2 × 1.2
2.4
Back
Poor (Q=1-4)
Portal
Entrance
1.2 × 1.2
2.1
Resin
Walls
Normal to
Primary
Grouted
Adit Portal
Opening
Portal
Rebar
General
Entrance
1.5 × 1.5
2.1
Support
Fair to Good Back
(Q>4)
Portal
Entrance
1.2 × 1.2
2.1
Walls
Corrugat
ed Steel
NA
NA
NA
Pipe
Portal Face Secondary
Cable
Horizontal
2.0 × 2.0
4.5
Bolts
into face
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Welded
Wire
Mesh
Gauge
(mm)
Fiber
Reinforced
Shotcrete
Thickness
(mm)
6
75
NA
6
NA
NA
6
50
TABLE 16.9
GROUND SUPPORT SELECTION
Support
Area
General
Brow
Support
(Only
Installed in
Rock Brows)
Intersections
Rock Quality
-
Support
Application
Area
Back
Support
Type
(Primary/
Secondary)
Rockbolt
Pattern
and
Specifications
Surface Support
Support
Element
Max.
Spacing
(m)
Cable
Bolts
1.5 x 1.5
(minimum
of three
bolts per
row)
Cable
Bolts
2.0 × 2.0
Secondary
Poor to Very 3- Way 4x4
Secondary
Good (Q>1)
(5.7 m)
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Inclination
Min.
Length
(m)
4.5
Welded
Wire
Mesh
Gauge
(mm)
Fiber
Reinforced
Shotcrete
Thickness
(mm)
NA
NA
NA
NA
Vertically
into
the
brow
4.5
Page 318 of 577
Normal to
Opening
Backfill Strength
Backfill strength recommendations are provided in (Table 16.10 and 16.11). The strength estimates
were based on empirical methods, considering various failure modes using the Mitchell equations
and limit equilibrium equations. The assessment included free standing stability, flexural failure,
block sliding, caving and rotational failure. Punching shear failure was also assessed, this
mechanism is typically applied for long term or closure applications and is not typically applied to
temporary mining settings such as those at the Los Ricos South Project (e.g., short exposure during
underhand stoping activities).
The in situ backfill strength was assessed to a minimum Factor of Safety (“FoS”) of 1.5. If GoGold
chooses to use a FOS less than 1.5, it is recommended that the effected stopes be operated using
non-entry mining methods.
TABLE 16.10
BACKFILL STRENGTH FOR DRIFTING DRIVING AND OVERHAND APPLICATIONS
Case
Strength Required for
a 1.5 FOS
4.0 m x 4.0 m drifting through
backfilled stopes
600 kPa
Overhand Mining (20 mH x 15 mL
x 5-15 mW) – Any freestanding
stope
400 kPa
Minimum strength required to drift
on top of backfill
200 kPa
(Assumed based on
experience)
Specifications
Ground support must be
installed no more than 1 round
back from the drift face. Strength
assumption only applies to a
maximum unsupported drift
length of 4 m. Assumes
consistent
paste
strengths
throughout the paste volume.
TABLE 16.11
BACKFILL STRENGTH FOR UNDERHAND APPLICATIONS (MPA)
Situation
Case
Underhand
Mining –
>10 m Span
Non-Entry
Stopes (we
understand
Considering Punching Shear
Ignoring Punching Shear
Strength
Strength
Strength
Strength
Strength
Strength
required
required for required for required for required for required for
for a 1.0
a 1.3 FOS
a 1.5 FOS
a 1.0 FOS
a 1.3 FOS
a 1.5 FOS
FOS
1.7
2.2
2.5
1.1
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1.7
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TABLE 16.11
BACKFILL STRENGTH FOR UNDERHAND APPLICATIONS (MPA)
Situation
Considering Punching Shear
Ignoring Punching Shear
maximum
span is 15 m)
Underhand
Mining –
<10 m Span
Non-Entry
Stopes
1.7
2.2
2.5
0.8
1.0
1.1
Conclusions
•
The current engineering geology model is based on limited coverage of geotechnical
and hydrogeological data and incorporates assumptions to address data gaps.
•
The completed geotechnical and hydrogeological program has known data gaps due to
the increased underground mine footprint during the advancement of the Project. To
mitigate data uncertainty and confirm geotechnical design assumptions further
geotechnical drilling is required on the order of a minimum of 6,000 m in the areas
identified above in Figure 16.6. The program should include the following:
o Geotechnical logging of drill holes;
o Laboratory strength and elastic modulus testing;
o Televiewers surveys in available drill holes (specific depths can be targeted to avoid
areas where poor ground may result in loss of the tool);
o VWP installation in the deepest area of the mine; and
o Additional hydraulic conductivity testing in areas of blocky ground and deeper
mining areas.
•
When additional data is available, an RQD block model should be developed to assess
spatial variation in fracturing intensity within domains, to identify weak zones that
would lead to local variation in the mine design.
•
Televiewer surveys and oriented drill core data collection should be aligned to collect
data in the blind zones of J1 and J2.
•
The underground mine geotechnical design guidance and recommendations are based
on a data set with known data gaps and uncertainty. Use of the design guidance and
recommendations for cost estimation and scheduling must be done with understanding
of the limitations of the EGM.
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16.1.2.2
Stope Sizing and Stability
Dimensions for open stopes can be estimated based on an empirical stability graph, developed by
Golder (1981) and updated by Potvin (1988) and Nickson (1992) amongst others. The stability
graph plots the stope Hydraulic Radius (“HR”) versus a Modified Stability Number (N’). The HR
of the exposed stope face is calculated by the ratio of the face area over the face diameter. The
modified stability number is estimated by the following equation:
𝑁 ′ = 𝑄′ 𝑥 𝐴 𝑥 𝐵 𝑥 𝐶
Values for the four sub-components vary as per Table 16.12.
TABLE 16.12
MODIFIED STABILITY NUMBER FACTORS
Factor
1
2
Factor Type
Value Range
A
Rock stress
0.1-1
B
Joint orientation adjustment
0.2-1
C
Gravity adjustment
2-82
Q’
Rock mass quality
0-100
In Eagle “Plum”1
1 in HW,
0.3-1 in back
0.2
(worst case
scenario)
8 in FW,
2 in back
1.4-5
Elsewhere
0.3-1
The “Plum”is a sub-zone of the Eagle Vein located roughly mid-way along strike of the mining target.
Calculated as 8-6*cos(Stope Dip).
Table 16.13 presents stope sizes recommended by WSP.
TABLE 16.13
WSP RECOMMENDED STOPE SIZING
Veins
Sub-Zone
Eagle
Eagle
Abra
Plum
Elsewhere
All
Stope Size
(mH x mL)
30 x 20
20 x 15 or 30 x 10
20 x 15
These dimensions correspond to the “Unsupported Transition” and “Stable with Support” areas of
the modified stability number charts (WSP, 2025). As the joint orientation factor B has been
selected for the worst-case scenario, it is expected that the majority of stopes will not require
additional support, and that atypical localized conditions will be controlled by location-specific
measures. Cable bolts are a common element in the mine ground support standards and can be
used for additional stope support where necessary.
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16.1.2.3
Level Spacing
Level spacing is nominally 20 m, with some levels slightly closer. Improved conditions in specific
parts of the Eagle and New Abra areas appear to support level spacings greater than 20 m, however,
these areas include only a portion of the strike of each area. As a result, the lesser of the two
spacings (20 m versus 30 m) was selected. As 30 m level spacings would allow a reduction in
level development of 33%, the Authors recommend additional geotechnical analysis, particularly
of the Eagle Vein, to evaluate the practicality of increasing the level spacing. In the New Abra
area, as the Old Abra stopes are limited to 20 m spacings and the division between the two zones
is relatively arbitrary, expanding the level spacings will offset every other New Abra level by 10
m from the adjacent Old Abra level, increasing the complexity of access and sequencing, and
limiting the benefit of more widely spaced levels. Therefore, increased level spacing in New Abra
is unlikely.
16.1.2.4
Faults and Breccias
Ongoing geotechnical and geological drilling has improved the understanding and interpretation
of the Deposit appreciably since the previous PEA (P&E, 2023). Three significant faults and two
significant breccia zones are known. All fault and breccia solids were expanded by 5 m in all
dimensions to generate an “impact zone” where adverse conditions could be expected. A minority
of stopes intersect these zones and are expected to experience adverse conditions that can be
mitigated with site-specific modifications to ground support, dilution, or recovery.
Flags were applied to the mining shapes based on the portion of the shape intersecting the 5 m
offset zones of the faults and breccias. If more than 10% of a mining shape intersected the offset
zone of a fault, the shape was flagged for further attention during mining. A total of 8.5% of stope
tonnes are derived from fault shapes flagged in this manner, predominately in the New Abra area.
For breccias, the proportion of the shape in the offset zone was determined and the shape was
flagged for attention. A total of 18.3% of stope tonnes are located within the 5 m offset zones of
the breccia, predominately in the Old Abra area. The breccias have a larger impact than the fault
zones on overall tonnes, however, there is less risk (dilution) than the fault zones since the breccias
are generally mineralized.
Rascadero Fault
Located between the Eagle and Abra Veins, dipping northeast, the Rascadero Fault forms the upper
boundary of mining on the northern side of the New Abra Vein, and the lower boundary of mining
on the southern side of the Eagle Vein. Development is sited to avoid the fault zone wherever
possible, and to cross through it on a perpendicular strike if avoidance is not possible.
Pithaya Fault
The Pithaya Fault is located on the HW side of the Abra Vein and dips southwest. It generally
does not intersect the mining areas. A minority of stopes from the northern end of the New Abra
Vein intersect a 5 m offset from the Pithaya Fault, in an area largely coincident with the intersection
zone of the Rascadero Fault. The Pithaya Fault intersects only one area of new development near
the New Abra ventilation adit portal, and its impacts are likely to be indistinguishable from the
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impacts of weathered overburden on the construction of the portal. Previous HMW development
has intersected and passed through the fault without difficulty.
South Fault
The South Fault is located at the southern end of the Old Abra mining area and dips to the north.
It generally forms the southern limit of mining at Old Abra, with an insignificant portion of mining
occurring on the southern side of the fault. New development to access the Farmyard portal passes
through the fault on a roughly perpendicular strike. HMW development has intersected and passed
through the fault previously without difficulty.
Eagle HW Breccia
A breccia zone is present in the HW of the Eagle Vein, generally located more than 5 m from the
HW of the stopes, that undulates closer and further away with change in depth.
Abra HW Breccia
Similarly to the Eagle Vein, a breccia is located in the HW of the Abra Vein, however, it is
generally located closer to the stopes at Abra than at Eagle and affects a significant number of the
transverse stopes mined in the Old Abra Zone.
16.1.2.5
Historical Mine Workings
HMW areas are defined adequately for the purposes of this FS. As development and mining
progresses, additional drilling will be required to improve the precision of the survey positioning
and extents. A portion of this void definition drilling is likely to be coincidental with grade
definition drilling, however, additional resources and time have been included in the mine plan to
account for excess void definition drilling.
HMW areas below the 1,150 m EL are expected to be saturated with water. HMW areas above
this point drain to the Farmyard Portal and are expected to experience localized pooling of water
only. Inspection of HMW areas at the El Troze and Windows areas show dry ground. Dewatering
of levels below 1,150 m EL in the Old Abra area will be done via a borehole pump utilizing the
historical shaft in the area.
HMW stoping areas in the Abra Vein, particularly the Old Abra area, are contiguous with
significant portions of the underground Mineral Reserve. These areas are assumed to be filled
with free-flowing mineralized material and will be extracted along with the in-situ portion of the
Mineral Reserve. The general sequence for underhand mining (which forms the vast majority of
the mining of this type) around these areas is described below.
1. Develop FW drift and crosscuts until the crosscut intersects the FW of the HMW stope;
2. Extract free-flowing material from the stope until it is empty;
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3. If “rat-holing” or coning occurs, proceed to the level below and draw material down
from there to the floor of the level above;
4. Secure brow and survey stope via cavity monitoring system (“CMS”);
5. Drill off stope using upholes. HMW stope to be used as slot void for first blast;
6. Extract blasted material;
7. Construct PF bulkhead. Rammed rock is likely to be sufficient in most cases, but
engineered barriers may be required in some cases;
8. Fill open stope with PF of appropriate strength from crosscut above and allow to cure;
o If top-cut access is unavailable, drill a hole to intersect the FW of the stope as near
to the corner of the back as possible, and pump fill via the hole. A smaller-diameter
“breather” hole should be drilled in parallel to prevent pressurization during filling;
9. Develop through cured PF to stope on HW side of the HMW;
10. Extract as per normal uphole extraction methods;
11. Fill stope as per 7 and 8 above;
12. If additional stopes remain on the FW side of the original HMW area, extract them on
retreat as per 10 and 11 above; and
13. If mining continues below, redevelop through cured paste on minimum profile to
provide access for PF lines for stopes below.
In very rare cases, there may be two offset HMW voids intersected on the same crosscut. In these
cases, the process from 1-8 above will be repeated until the western-most (HW side) HMW void
is filled, and then mining will take place on retreat as per 9-12 above.
Intersecting HMW development is generally avoided by the mine design where possible, with one
portal (the Farmyard portal) into the Abra Vein planned for re-use (slashing out and re-support of
the tunnel will be required). All other intersections of HMW development areas are incidental.
To the extent possible given the accuracy of location of the HMW areas, it is planned to either
completely intersect the HMW development in the vertical plane (HMW development is nominally
3.0 mH and fits within the profile of the planned 4.0 mH development), or to leave a minimum 1:1
pillar between new development and HMW development. It is expected that probe drilling will
be part of on-level development to better define the pillars. If necessary, PF can be inserted into
HMW development voids via boreholes where thin pillars are discovered if additional ground
support in the pillars is deemed inadequate.
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16.1.2.6
Pillars
Two forms of pillars exist in the UG mine: the Crown Pillar, segregating the UG production
operations from surface topography; and Artificial Sill Pillars, segregating mining blocks and
allowing undermining of filled stopes.
Crown Pillar
To maintain sufficient integrity of the topography above the UG mine, a crown pillar of a nominal
thickness is maintained between UG production operations and the topographic surface in any
direction. Due to the nature of the topography, the vertical distance to surface can be significantly
greater than 20 m where the mountainside is steep. WSP has estimated the thickness at 20 m for
stope spans up to 6 m.
A significant amount of economic mineralization exists within this pillar above the 1,290 m EL
level. While the majority of the tonnes within this pillar on the Abra Vein could be extracted by
eventual open pit mining, it is planned that they will be extracted using underground methods
similar to adjacent stopes, and backfilled in a similar fashion. If necessary, backfill can be
introduced from surface. Any stability issues are expected to manifest as increased dilution, and
if material cannot be recovered via underground methods, it will eventually be recovered by open
pit methods later in mine life. The same methods are expected to be used on the Eagle Vein,
however, open pit mining is not expected in this area and material not recovered from the
underground will largely be lost. Table 16.14 shows an estimate of the portion of UG Mineral
Reserves from each vein located in crown pillar areas.
TABLE 16.14
PORTION OF UG MINERAL RESERVES EXTRACTED FROM
CROWN PILLAR AREAS
Area
Eagle
New Abra
Old Abra
Total
Mineral Reserve in
Crown Pillar
(%)
1.8
7.3
1.1
2.3
Pillar AgEq
(Moz)
0.9
7.2
0.5
1.8
Figure 16.7 shows the location of Crown Pillar areas.
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FIGURE 16.7
LONGITUDINAL PROJECTION SHOWING PILLARS
Source: P&E (2025)
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For overhand mining, mining of the Crown Pillar areas will not commence until at least one level
beneath the pillar has been mined out and backfilled, limiting the influence of any adverse impacts
from pillar recovery to a localized area. For underhand mining in Old Abra, where the crown pillar
areas are extracted first, the intersection of the Crown Pillar and underground mining shapes almost
exclusively occurs in the HMW areas, meaning no blasting is needed to extract the material, which
will limit stability issues. Drill and blast activities in the Crown Pillar in the Old Abra Vein are
limited to a total of two stopes and less than 40 koz AgEq in 8 kt of ore.
Artificial Sill Pillars
The bottom level of each mining block, with the exception of the lowest mining blocks in the mine,
will be backfilled with high-strength PF to create a 20 m thick pillar that can be undermined by
the level below, using remote drilling and excavation. These pillars are spaced at varying intervals
dependent on scheduling constraints. Figure 16.7, above, shows the position and quantity of the
pillars. Material excavated from stopes situated directly below the pillars is recovered at a slightly
increased mining loss due to the use of blind upholes.
16.1.2.7
High-Grade Domains
Narrow higher-grade domains within larger lower-grade domains provide the majority of mining
targets in the UG mine plan. Overbreak from these targets is largely contained within the lowergrade domains, meaning that diluting material generally contains mineralization. While the HMW
shapes do not completely overlap these high-grade domains in the Abra area, they are relatively
contiguous along strike and dip, and it is likely that these domains and the HMW shapes largely
coincide. As a result, it is probable that the currently estimated location of the HMW shapes,
outside of LIDAR-surveyed areas, are slightly different from their true locations.
16.1.2.8
Ground Support Standards
Development ground support design for the UG was performed by WSP (2025). Table 16.9 above
provides the recommendations.
Cablebolts will be installed in stope brows as per Table 16.9 above, and in areas of poor ground
(near faults or breccia zones) it may be necessary to support stope walls with rings of cablebolts.
Due to the significant expected number of cablebolts expected to be required over LOM (in excess
of 20 k tendons), a specific cable bolting unit has been included in the UG mining fleet.
Shotcrete is expected to form part of the ground support system for the UG, and also to be used
for securing rammed-rock PF bulkheads and sealing brows. In excess of 7 k m3 of fibre-reinforced
shotcrete is planned to be used over LOM, equivalent to over 125 k m2 of coverage at 0.05 m
thickness. Specific mixers and shotcrete sprayers are included in the mining fleet.
16.1.3
Cut-off Grades
A detailed breakdown of Cut-Off Grade (“COG”) calculations is provided in Section 15.3.3.
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Table 16.15 shows the portion of the Mineral Reserve generated from stopes at the various AgEq
COGs. Figure 16.8 shows the spatial distribution of stopes by COG.
TABLE 16.15
UG STOPE PROPORTIONS BY COG
COG (g/t AgEq)
210
180
150
130 (Manual)
80/130/150 (Combination
around HMW)
% of Stope Tonnes
26.5
2.3
50.7
0.3
% of Stope AgEq oz
34.6
2.6
46.9
0.2
20.2
15.7
Development material utilizes a COG of 65 g/t AgEq. All development in excess of this grade
generates incremental profit relative to treating it as waste rock. Approximately 31% of all
development ore is material of this type, containing ~12% of all AgEq contained in development
material.
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FIGURE 16.8
LONGITUDINAL PROJECTION OF STOPES BY DESIGNED AGEQ CUT-OFF GRADE
Source: P&E (2025)
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16.1.4
Development
A total of 60.9 km of linear lateral development and 1.9 km of linear vertical development is
planned over the LOM. Approximately 54% of all lateral development is operating development.
All vertical development is capital development.
HMW development workings unavoidably intersect certain lateral portions of the new
underground workings (new vertical development avoids them entirely) although efforts have been
made to avoid them whenever possible. These areas may be open or may be collapsed. No
contingency has been added to development quantities to account for these areas.
16.1.4.1
Lateral Development
Lateral development is sized to accommodate the largest piece of equipment expected to operate
within the mining area. In general, the LH drills drive the limits for operating development
(clearance around the mast during dip/dump and ring drilling), while haul trucks drive the limits
for most capital development, particularly the ramp. Other factors, such as ventilation requirements
or speciality services (e.g. maintenance bays or remuck bays) are also taken into account when
sizing lateral development. Table 16.16 shows lateral development totals by size and category.
TABLE 16.16
LATERAL DEVELOPMENT METRES SUMMARY
Development Type
Ramp
Ventilation Adits
Level Access
FW Drift
Electrical and Dewatering
Maintenance
Other Infrastructure
Vent Accesses and Transfers
Remuck Bays
Longitudinal Sills
Transverse Crosscuts
Transverse Crosscuts through PF
Subtotal
Subtotal
Total1
Nominal Size
Capital /
(mW x mH)
Operating
4.5 x 4.5
Capital
4.5 x 4.5
Capital
4.5 x 4.5
Capital
4.0 x 4.0
Capital
2
4.0 x 4.0
Capital
5.0 x 6.5
Capital
3
4.0 x 4.0
Capital
4.0 x 4.0
Capital
4.5 x 5.5
Capital
4.0 x 4.0
Operating
4.0 x 4.0
Operating
4.0 x 4.0
Operating
Capital
Operating
Capital + Operating
Quantity
(m)1
9,887
209
1,864
9,469
1,373
168
191
2,974
1,617
16,232
14,876
2,050
27,751
33,158
60,909
Notes:
1
Totals may not sum due to rounding.
2
Profiles vary from 3.5 x 4.0 to 4.5 x 4.5 depending on application.
3
Profiles vary from 3.5 x 4.0 to 4.0 x 4.0 depending on application.
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In addition to the above, allowances have been included in the design to account for passing bay
slashes, safety bay cut-outs and electrical distribution bays totalling an additional excavated
volume equivalent to 1,280 m of 4.0 mW x 4.0 mH lateral development.
A small portion of lateral development will involve slashing out existing HMW development for
re-use, mainly on the Farmyard portal drift. This development accounts for less than 0.3% of the
stated 60.9 km of lateral development. Where HMW voids are slashed out, the existing drift will
be used as the slot, with holes in the advancing face drilled at the desired final perimeter and
blasted into the void.
Approximately 4% of all lateral development is located within 5 m of a fault structure and is
expected to require significant additional ground support during excavation. The vast majority
(93%) of these metres are operating development, namely sills and crosscuts.
16.1.4.2
Vertical Development
Vertical development is limited to ventilation raises, which are driven either via raisebore or drop
raising methods, with drop raising being used for slightly more than half of vertical development.
Raiseboring is used where raise segments are longer than 40 m. Due to the relatively short level
interval, raiseboring is utilized to drive a single raise segment connecting three consecutive levels
with minimal lateral offset (raisebored raises are designed to have a dip of 70° or higher), whereas
drop raises are used to connect two consecutive levels where lateral offsets between levels make
raiseboring inefficient. Table 16.17 presents vertical development metres planned over the LOM.
TABLE 16.17
VERTICAL DEVELOPMENT METRES SUMMARY
Development
Type
Raisebore
Drop Raise
Total1
1
Nominal Size
(m)
3.4 diameter
3.0 W x 3.0 H
Capital /
Operating
Capital
Capital
Capital
Quantity
(m)
856
1,001
1,856
Totals may not sum due to rounding.
Escapeways are installed in slightly less than half of all ventilation raises, predominately in drop
raises (682 m of escapeway in drop raises and 185 m in raisebored raises).
16.1.5
16.1.5.1
Production
Mining Methods
All mining of UG production areas utilize a form of longhole stoping with cemented paste backfill.
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Longitudinal Retreat Mining
Longitudinal retreat mining involves driving an access through ore along the strike of the vein to
the end of a group of stopes. The stope is drilled using downholes (or upholes where no top access
exists), loaded and blasted, and ore is recovered from the bottom access using an LHD. Once the
blasted material is excavated and cavity survey is complete, a wall is built to seal the bottom access
of the stope, and the stope is filled with PF. After a curing period (nominally 28 days), the cycle
repeats on the next stope in the sequence (drilling and loading operations can commence prior to
complete curing, blasting cannot). Longitudinal retreat mining progresses from the bottom of a
mining block to the top, working on top of PF. Backfilling can either be from the sill drift above
or through a borehole drilled from nearby development. Boreholes will be the sole option for the
last level in each mining block as there will not be an open sill drift above. Longitudinal retreat
mining minimizes waste development by avoiding offset drifts and crosscuts, at the cost of reduced
mining faces on a level.
Transverse Access Mining
Transverse access mining involves driving an access drift offset from the vein (in this case,
nominally 12 to 15 m from the contact) parallel to strike, and then driving perpendicular crosscuts
across the vein on a set spacing (in this case, nominally 15 m centres). For the Los Ricos South
Project, transverse stopes are generally drilled using upholes due to access limitations (the overcut
level will usually be completely backfilled prior to mining), however, downholes are preferred
where they can be utilized. Mining of individual stopes progresses similarly to longitudinal retreat
mining, however, the sequencing of multiple stopes on a single crosscut can be more complex,
particularly around the HMW areas. Backfilling will generally be from the crosscuts on the level
above, or through boreholes drilled from nearby development. Transverse mining maximizes
productivity at the cost of additional development.
Generally, for transverse access mining, the optimal sequence of stopes on a single crosscut
retreats from HW to FW, as is the case for stopes at the Project where transverse mining is not
contiguous with HMW areas. Due to the positioning of ore and the added complexity of the HMW
areas, the sequence is often reversed, progressing from FW to HW and driving through engineered
backfill.
Alternating crosscuts are extracted in a primary-secondary sequence during transverse access
mining. Stopes on primary (odd-numbered) crosscuts are excavated and filled first prior to stopes
on secondary (even-numbered) crosscuts being excavated. This limits sequencing issues and
exposure to open stopes in adjacent crosscuts.
Uphole Versus Downhole Drilling
Where feasible, stopes will be extracted using downhole drilling. This form of drilling generally
lends itself to simpler QA/QC as more of the holes break through to the undercut and can be more
easily surveyed from above. Additionally, loading of the blast holes from the top is simplified
relative to loading upholes. Finally, for raise and slot blasting, access from above allows more
flexibility in blast sizing and void management. Downhole drilling is the dominant drill type in
longitudinal retreat areas of the mine and is used in some transverse access mining areas
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(particularly in the Eagle Vein) where HMW are not expected to be encountered and the ore is
thick enough to support the method.
Uphole drilling is utilized where the stope geometries or accesses preclude the use of downhole
drilling. In general, this occurs on the upper level of a longitudinal mining block where the overcut
has already been backfilled, or in transverse access stopes adjacent to the HMW areas where
underhand mining is required. In both cases, holes will be drilled until contact with cured PF is
reached. In a minority of cases, uphole stopes will be drilled “blind” in previously unmined ground
where the stopes are not continuous to the level above and there is no physical boundary to limit
drilling. Stopes that are drilled “blind” have higher dilution and mining loss due to more difficult
QA/QC.
Approximately 48% of LH stope drilling is downhole and 52% is uphole. Less than 2% of stopes
are drilled with blind upholes.
Extraction of Historical Mine Workings
Historical mining at the Project predominantly occurred on the Abra Vein, with the majority of
crosscuts in the Old Abra area of the underground expected to intersect HMW areas. These areas
are assumed to be a combination of uncemented backfill with some mineralization, and void (air).
The backfill material is expected to be free-flowing. These HMW areas generally coincide with
the high-grade sub-domains of the Abra Vein, but still have contiguous ore on one or both of the
HW or FW. Less than 3% of the Mineral Reserve tonnes derive from HMW material, however,
approximately 30% of the tonnes are derived from stopes contiguous with HMW areas.
As the backfill is free-flowing, any attempt to excavate in an overhand fashion would require the
extraction of all backfill from the levels above prior to commencement of mining the in-situ ore,
which could destabilize the HMW voids and result in negative geotechnical and economic
consequences. Therefore, underhand mining was selected for the majority of stopes adjacent to
HMW areas (localized small HMW stopes in the New Abra and Eagle Veins are extracted using
overhand methods as their vertical extents are minimal). To safely and economically extract these
areas, the following process is used where ore exists on both sides of the HMW area, and the PF
in the level above has cured to sufficient strength for undermining:
1. Drive a drift parallel to the vein, offset by 12 to 15 m to the FW of the vein;
2. Drive crosscuts perpendicular from the vein on 15 m spacings to just before the HMW
contact, using probe drilling to determine the breakthrough location;
3. Prior to breakthrough, drill the stope rings in previously unmined ground adjacent to
the HMW stope. This prevents drilling near open brows;
4. Break through development to the HMW area and draw down the backfill until the
brow cracks. Ram the brow closed with an LHD using a “rammer-jammer”;
5. Load and prime the drilled holes, pushing the primers into the collars and temporarily
plugging them with bottle brushes or cones;
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6. Utilizing an LHD on remote, excavate the brow until backfill reaches the floor level.
In some instances, drawing down the backfill may be done from the level below if
desired or necessary;
7. Blast the stope, using the HMW void as a slot. If void is insufficient to fire the entire
stope, excavate the blasted material until the brow cracks, and ram the brow closed.
Repeat from Step 5 until the entire stope is blasted and excavated;
8. Install a backfill barricade in the development access;
9. Backfill stope with high-strength PF from the level above and allow to cure to
minimum 600 kPa strength (see Section 0);
10. Excavate through cured PF; and
11. Extract ore on HW side as per normal transverse access stoping practices once adjacent
stope has cured to full 1.4 MPa strength.
Where the HMW void is located on the FW side of the in-situ ore only, the above sequence is
followed with the exception that no drill and blast is required. Filling begins after the HMW void
is drawn down to the floor of the sill and a barricade is placed.
Where the HMW void is located solely on the HW side of the in-situ ore, the above sequence stops
at item 9, and the PF is allowed to cure to full strength for eventual underhand mining on the level
below.
It is expected that some PF introduced into stope voids around HMW areas on primary crosscuts
will be blasted into adjacent HMW voids on secondary crosscuts and increase dilution therein.
The average dilution factor applied to all underhand mining stopes accounts for this.
Some crosscuts are expected to intersect multiple HMW lenses as they progress. In this case, the
above process would be completed for the lens closest to the FW and then, after the PF is
sufficiently cured, a supported drift would be developed through the PF until previously unmined
ground is reached, at which point the process would begin again for the next lens.
Figure 16.9 shows the extraction sequence for a crosscut in the Old Abra mining area that intersects
HMW voids.
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FIGURE 16.9
EXTRACTING STOPES ADJACENT TO HISTORICAL MINE WORKINGS
Source: P&E (2025)
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Where new lateral development crosscuts HMW lateral development, either fencing/barricades
will be installed, or, more practically, the HMW voids will be backfilled with any individual type
or combination of development waste or low-strength PF to limit ventilation recirculation and
improve geotechnical stability.
16.1.5.2
Stope Sizing
Levels are spaced at a nominal interval of 20 m. Longitudinal stopes are generally less than 15 m
wide. Wider areas are experienced in localized portions of the Eagle Vein, and where stope widths
exceed 15 m parallel longitudinal extraction drifts are included for the wider extent. While
geotechnical analysis indicates locally better ground conditions in the Eagle Vein that would
support larger stopes (up to 30 mH x 20 mL), a standardized stope design of 20 mH x 15 mL was
selected for longitudinal areas the study. In practice, longitudinal retreat stope lengths may be
extended further if conditions permit. As maximum stope strikes for the Abra Vein are limited to
15 m, a crosscut spacing of 15 m was selected for transverse access mining throughout the mine
for consistency and simplicity.
Stope thicknesses vary by location, with the Eagle Vein stopes generally being thicker than the
Abra Vein stopes. Table 16.18 shows average thicknesses of stopes in the various areas.
TABLE 16.18
AVERAGE STOPE THICKNESS
Veins
Eagle
New Abra
Old Abra
Total
Mining Type
Transverse
Longitudinal
Transverse
Longitudinal
Transverse
Longitudinal
Transverse
Longitudinal
Average HW-FW
Thickness
(m)
18.7
7.9
17.0
8.8
9.6
7.5
12.1
8.1
The average HW-FW thickness of HMW areas is approximately 3.7 m. Stopes abutting the HMW
workings on average are targeting contiguous ore adjacent to the HMW of approximately 6.3 m
thickness. The average thickness of all stopes is approximately 10.6 m, with stopes on the Eagle
Vein averaging 13.6 m and stopes on the Abra Vein averaging 9.3 m.
Based on geotechnical recommendations, any transverse stopes with a HW-FW span larger than
15 m will be split into multiple panels to limit the span to the recommended 15 m limit.
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16.1.5.3
Backfill
All backfill for the UG mine will be cemented PF. Recipes and strengths will be optimized for
specific backfill applications. Binder will be locally sourced Portland cement. Tailings will be
recovered from the dry stack storage and re-slurried at the paste plant near the Eagle portal prior
to being pumped underground. PF reticulation utilizes gravity feed.
Approximately 40% of all PF will be high-strength fill suitable for eventual underhand mining.
An additional 41% will be PF with sufficient strength for exposure of a wall for adjacent overhand
mining. Finally, approximately 19% of PF will be minimum-strength not suitable for exposure,
whose primary purpose is the emplacement of tailings underground. A total of 3.5 M m3 of PF is
expected to be placed over the LOM, comprised of 2.9 M m3 into production areas and 0.6 M m3
into development areas.
Over LOM, PF will utilize 290 kt of cement and 4.9 Mt of tailings.
If necessary, additional PF can be placed into HMW development openings to further decrease
surface tailings storage requirements.
PF Strength Requirements
WSP (2025) provided the PF strength estimates under various conditions as presented in Table
16.19.
TABLE 16.19
WSP PF STRENGTH REQUIREMENTS (MPA)
Mining Type
Underhand,
>10 m span
Underhand,
<10 m span
Overhand
Development3
Factor of Safety (FoS)1
Including Punching Shear2
Excluding Punching Shear
1.0
1.3
1.5
1.0
1.3
1.5
1.7
2.2
2.5
1.1
1.4
1.7
1.7
2.2
2.5
0.8
1.0
1.1
0.4
0.6
Notes:
1
Factor of Safety is a unitless factor that expresses the strength of a system versus it’s intended load.
2
Punching shear is a failure mechanism typically used in longer term or closure applications. It is not typically
applied to temporary mining settings such as short exposure above underhand stopes.
3
Development through backfilled stopes comprises roughly 3.4% of all lateral development.
Strengths of 1.4 MPa for underhand stopes of >10 m span and 1.0 MPa for underhand stopes of
<10 m span were selected for the Project, corresponding to a FoS of 1.3. From a practical
perspective, all underhand stopes were treated as though their spans were >10 m to provide a
conservative estimate.
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PF Recipes and Cure Times
Paterson and Cooke (P&C, 2025) performed extensive testing on various PF recipes, testing
various combinations of dewatering, admixtures and tailings sources (report titled “GoGold Los
Ricos Backfill Test Work Report, P&C Project No.: GRL-32-1541-00-TW-REP-0001 Rev C
issued January 23, 2025”). A target cure time of 28 days was utilized to enable rapid mining
cycles, particularly later in mine life during the underhand mining period in Old Abra. The recipes
were analyzed by the Authors, in conjunction with P&C, to determine the most economic methods
of generating the required strengths, as shown in Table 16.20.
TABLE 16.20
PASTEFILL RECIPES
Application
Span
(m)
Strength
Target
(kPa)
>10
Any
Any
<10
1,400
400
Minimal
1,000
Binder
(% by
solids
mass)
9.2
3.9
1.5
7.3
5
600
5.2
Underhand Mining
Overhand - Endwall Exposed
Overhand - Not Exposed
Underhand Mining1
Tunneling Through
Pastefill1,2
Tailings Blend
(nominal %)
New
Old
Eagle
Abra
Abra
50
50
0
33
33
33
0
0
100
x
x
x
x
x
x
Notes:
1
These recipes were not utilized in the mine plan, so their portions of each tailings source have not been calculated.
2
Tunnelling through pastefill is only anticipated to be necessary in underhand mining, where 1,400 kPa PF will
already be utilized, negating the need for this recipe to be used.
Although testing indicated that desliming of tailings would lead to marginal improvements in the
PF plant operating cost through the reduction of binder content, the projected capital expenditures
and additional permitting risks necessary for settling of the slimes were determined to outweigh
the operational savings, and whole tailings have been utilized instead. Admixtures were also tested
and were determined to be uneconomic.
Tailings Storage and Segregation
Specific tailings sources are required to achieve the required strengths using the expected quantity
of binder. Since the process plant does not feed tailings directly to the backfill plant, segregated
areas of the dry stack will be utilized to ensure that the correct tailings blend can be excavated and
delivered to the backfill plant as needed.
The Old Abra tailings source is associated with an increased binder requirement to achieve useful
strengths, and therefore is deleterious to the Eagle / New Abra tailings sources. The operational
plan for the process plant to ensure that cross-contamination of the streams is minimized is as
follows:
1. Maintain continuous feed, regardless of ore source;
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2. Where possible, feed one of the following two ore sources:
a. Eagle / New Abra blend (preferably 50/50 blend, but not strictly necessary)
b. Old Abra;
3. When one of the above ore sources is being processed, send dry tailings to designated
segregated storage areas; and
4. During cross-over of sources (Eagle/New Abra to Old Abra, or vice-versa), send
tailings to Old Abra storage.
a. An allowance of 15% loss of total Eagle/New Abra tailings into the Old Abra
storage is included in tailings balancing;
b. Higher usage of Old Abra tailings in the 400 kPa recipe provides further margin for
losses of Eagle/New Abra tailings into the Old Abra tailings stockpile.
Quarter-by-quarter tailings generation from processing, and tailings demand from backfill, has
been modelled to ensure that sufficient supply of the correct tailings source will be available at all
times in the mining plan. Over the UG LOM, 3.85 Mt of Eagle / New Abra tailings will be
generated and 3.23 Mt will be utilized in backfill, with approximately 0.6 Mt planned to be sent to
the Old Abra tailings stockpile due to contamination with Old Abra tailings. Approximately 40 kt
of additional Eagle / New Abra tailings will remain in the dry stack at the end of the UG LOM,
with the remainder made up of Old Abra tailings.
16.1.6
Dilution, Mining Loss and Recovery
Dilution describes the inclusion of sub-economic material in the mining process to recover ore.
Mining loss is defined as a portion of ore left behind in an excavation due to any or all of drilling,
blasting, excavating, or ground support issues. Recovery is the defined as the ratio of ore recovered
to ore planned (the inverse of mining loss). Additional details on these factors can be found in
Sections 15.3.4 and 15.3.5.
The mine-wide average internal dilution of ore to the process plant is estimated at 21.8%. The
average of external dilution on that material is estimated at 18.1%. The average mining loss is
8.8% and average mining recovery is estimated at 91.2%.
16.1.7
Equipment
All equipment is provided by the contractor. Table 16.21 shows the expected number of units
required by type for development and production operations based on the Authors’ estimates. All
nominal fleet estimates are based on 6.6 effective operating hours per shift, derived from a 12hour shift, 75% availability and 73% utilization.
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TABLE 16.21
ESTIMATED UG FLEET REQUIREMENTS
Machine Type
LHD (10 t)
Truck (30 t)
Jumbo Drill
Explosive Loader
Bolter
Scissor Deck
LH Drill
Shotcrete Machine
Grader
Cable Bolter
Flatbed
Fan Handler
Telehandler
Forklift
Backhoe
Fuel/Lube Truck
4WD Pickup
1
Nominal
2.3
5.3
2.0
4.5
2.0
2.1
1.4
0.3
N/A1
Ceiling
3
6
3
5
3
3
2
1
1
1
3
2
2
3
4
2
16
Spare
1
1
1
1
1
1
1
1
0
0
1
0
0
1
0
0
0
Total
4
7
4
6
4
4
3
2
1
1
4
2
2
4
4
2
16
These equipment types were estimated as complete units, not derived from development/production physicals.
It is assumed that the contractor will re-estimate the number of units of each type required to
achieve the mining schedule based on its own expectations of availability and productivity. As
machine rental is included in per-tonne costs, this will not impact the overall cost.
16.1.8
Services and Infrastructure
Mine services include electrical supply and distribution, ventilation, dewatering, compressed air
and communications systems. Major underground infrastructure items include dewatering pump
stations, refuge stations, explosives magazines and maintenance shops (both small maintenance
bays for preventative maintenance work, and a large underground shop located on the 1,150 m EL
ramp in Old Abra).
16.1.8.1
Ventilation
The Los Ricos South UG mine utilizes an exhaust (pull) system, with fresh air being drawn down
the ramps and Fresh Air Raises (“FARs”) onto active levels, across FWDs and into Return Air
Raises (“RARs”). Maximum steady-state airflow for the mine is approximately 240 m3/s (“CMS”)
and utilizes five primary ventilation fans of varying sizes and configurations. All primary
ventilation fans are expected to be equipped with variable frequency drives. All primary fans were
sized at 2.13 m diameter for standardization (other configurations can be slightly more efficient,
however, the difference is negligible and the simplicity of spares for a single size is beneficial).
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Airlocks will be installed in long straight sections of the connector ramps from the Old Abra area
to the New Abra area to ensure sufficient airflow in the active primary haulage routes. These
airlocks will be equipped with overhead ventilation fans, remote switches/light gates, and traffic
management systems to minimize the impact on transiting vehicles.
Table 16.22 shows the types, sizes and locations of the primary ventilation fans over the LOM.
TABLE 16.22
PRIMARY VENTILATION FANS
Fan Location
Fan
blade
diameter (m)
Fan hub diameter
(m)
Motor
Speed
(rpm)
Motor
Power
(kW)
Schedule Year
Setup
1
Eagle
North Vent
Adit
Eagle
South Vent
Adit
New Abra
Vent Adit
Old Abra
South Raise
Old Abra
Ramp
Airlocks1
2.13
2.13
2.13
2.13
1.22
0.91
1.09
1.09
1.09
0.46
900
720
1,200
1,200
1,200
150
25
2 x 200
100
2 x 20
Full blade
Half blade, 2x
in parallel
Not In Use
Not in Use
Full blade
-2
-1
1
2
3
4
5
6
7
8
9
10
11
12
13
Full blade
Full blade,
2x in
parallel
Not in Use
Not in Use
In Use
In Use
In Use
In Use
In Use
Not in Use
Not in use
Not in Use
Not in Use
Location of the fans varies over time, either on the top or bottom door of the upper or lower ramp airlock.
Initial development of the mine uses auxiliary ventilation through the Eagle portal and ventilation
adits to ventilate the development headings. Shortly below the portal, the ramp splits to the north
and south. Approximately one level below the portal are ventilation connections to surface (a raise
off the Eagle ramp to the northern Eagle ventilation adit and another raise off the 1,330 m EL level
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to the southern Eagle ventilation adit). Once these connections are completed, flaps are installed
at the Eagle portal and a primary ventilation exhaust fan is installed on the northern ventilation
adit drawing 90 CMS from the southern ventilation adit. As development progresses towards New
Abra, another adit is developed off the 1,310 m EL New Abra level to surface. Once development
reaches the upper levels of Old Abra, a raise is developed into the HMW areas at the southern end
of the Deposit, and a portion of the “Windows” area of the HMW is slashed out to provide an
additional fresh air intake. At this point a pair of primary ventilation fans are installed in the New
Abra ventilation adit to draw 145 CMS of additional flow through the New and Old Abra areas,
and the flaps are removed from the Eagle portal. At this point, production mining commences and
the mine has a total airflow of approximately 230 CMS.
As mining progresses, certain raises or adits are activated, blocked, or reversed to supply airflow
where necessary. An additional primary fan is eventually installed on the southern Old Abra
ventilation raise to draw air to deeper mining areas, and another primary fan is temporarily installed
in the southern Eagle ventilation adit during transverse mining operations in the upper levels of
the Eagle Vein. Airlocks are installed in the upper and lower portions of the Old Abra access
ramps to maintain airflow in the primary haulage routes. As mining progresses deeper at Old
Abra, the Farmyard portal is opened to provide parallel air intakes and reduce fan operating
pressure. Total airflow peaks at 240 CMS in YR 4 and gradually declines from there as various
mining areas are depleted. Peak power draw occurs around YR 8 and requires approximately 400
kW for the primary ventilation system in addition to 1,400 kW of auxiliary ventilation (for
production crosscuts and sills). Maximum power consumption for the primary ventilation system
occurs in YR 13 and requires 530 kW. At this point, due to decreased number of active mining
faces, auxiliary ventilation requirements are reduced to 350 kW. Average power draw over the
LOM for the combined auxiliary and primary ventilation systems is estimated at 1.4 MW.
Figures 16.10 to 16.13 show the ventilation system at various points over LOM.
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FIGURE 16.10
LONGITUDINAL PROJECTION OF VENTILATION SYSTEM YEAR -1
Source: P&E (2025)
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FIGURE 16.11
LONGITUDINAL PROJECTION OF VENTILATION SYSTEM YEAR 4
Source: P&E (2025)
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FIGURE 16.12
LONGITUDINAL PROJECTION OF VENTILATION SYSTEM YEAR 8
Source: P&E (2025)
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FIGURE 16.13
LONGITUDINAL PROJECTION OF VENTILATION SYSTEM YEAR 12
Source: P&E (2025)
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16.1.8.2
Dewatering
The majority of water inflow into the UG mine will come from drilling operations, with a total of
approximately 7.5 L/s of inflow combined from drilling, diesel condensate, and groundwater
inflows. Small level sumps are used to capture water on-level, and will be used to transfer water
(either pumped or through gravity) to pump stations located up to 200 m vertically apart in the
mine. In the pump stations, settling sumps will be used to segregate solids and clean water, and
clean water will be returned to the service water system, with excess pumped to surface
sequentially through each pump station above. Most pump stations will be equipped with ~400
kW centrifugal pumps (6 stages at 67 kW each) operating on a 33% duty cycle with a maximum
20 L/s capacity, however, the top pump station at the Abra Vein will be slightly smaller due to
reduced head required to reach surface (~65 m), and be equipped with ~280 kW centrifugal pumps
(5 stages at 55 kW). Each pump station will also have a small electric submersible pump (~30
kW) to provide initial head to the larger centrifugal pumps. Similar-sized submersible pumps will
be located in level sumps to allow level-to-level pumping prior to the installation of the pump
stations.
Dewatering of the HMW will be completed through the use of a 165 kW borehole pump located
in a borehole intersecting the 870 m EL in the historical shaft, pumping to the 1,150 m EL. This
pump will dewater the HWM areas from 1,150 m EL to 870 m EL over a period of three years,
and functions as an emergency dewatering system in the event of a 100-year flood situation; water
can be diverted from the new levels into the HMW areas below the new levels for immediate
containment prior to eventual pumping to surface.
Due to constraints on fresh water availability at the site, excess mine water will be treated and
recycled through the process plant where possible. It is expected that approximately 2.5 L/s will
be available to the process plant at a steady state. Any drainage to the Farmyard portal will be
captured in a sump near the portal and returned to the main mine dewatering system. Normal
outflows from the mine will be from the Eagle portal, with back-up/redundant lines run through
the northern Eagle vent adit.
Figure 16.14 shows the mine dewatering system at its maximum extents.
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FIGURE 16.14
LONGITUDINAL PROJECTION OF MINE DEWATERING SYSTEM
Source: P&E (2025)
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16.1.8.3
Compressed Air
The majority of underground machinery is expected to be electric-hydraulic in nature, with
minimal compressed air requirements. Two 120 kW centrifugal screw compressors will be utilized
to provide ~1,200 cfm of flow to each of the ramp areas. It is expected that compressed air will
be used primarily for explosive loading, and for ancillary construction work (drilling with
jacklegs).
An ethyl mercaptan (“Stench Gas”) injection system will be installed at each compressor for
emergency use. In addition to Stench Gas, methyl salicate (“Wintergreen”) injection systems will
be installed for line flushing after a Stench Gas injection. A small stand-alone compressor will be
located in the underground shop.
16.1.8.4
Electrical
Electrical power will be provided to the portal of each mine, and transmitted underground at a
nominal voltage of 13.8 kV. Modular transformers will be installed every two to three levels to
step down the primary voltage to a nominal usage of 600 V. The transformers will be relocated as
necessary as levels are mined out. Maximum load for the UG is estimated at 4.5 MW in YR 5,
with an average load of 3.4 MW. Level transformers are nominally 1,500 kVa.
Figure 16.15 shows the electrical and compressed air systems for the mine at their maximum
extents.
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FIGURE 16.15
LONGITUDINAL PROJECTION OF ELECTRICAL AND COMPRESSED AIR SYSTEMS
Source: P&E (2025)
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16.1.8.5
Refuge Stations and Egress
Emergency egress from the UG mine will be via ladderway in the ventilation raises. Underground
lunchrooms configured to serve as refuge stations will be installed according to applicable laws
and regulations.
There are seven permanent refuge stations located in the underground (two in Eagle, two in New
Abra and three in Old Abra), and a modular refuge station will be utilized during development
prior to the installation of permanent ones.
Escapeways are generally installed in drop raises using timber or prefabricated metal ladders
installed after the raises are supported (support will be installed using hand-held drills working off
the muck pile). Where escapeways are installed in raisebores, prefabricated systems (SafeScape
or similar) will be used instead. Raisebored raises will not be supported, however, they may be
lined.
Figure 16.16 shows the refuge stations and escapeways in the mine.
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FIGURE 16.16
LONGITUDINAL PROJECTION OF REFUGE STATIONS AND EGRESS
Source: P&E (2025)
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16.1.8.6
Explosives Storage
Explosives will be delivered to secure storage areas underground that conform to applicable laws
and regulations. Magazines are located on the 1,330 m EL level of Eagle and also on the 1,150 m
EL ramp connecting New Abra and Old Abra. It is expected that the total capacity of the two
explosives magazines will be approximately equivalent to one week’s explosive requirements (~10
t of blasting agents and 300 kg of initiating systems per magazine area). Control of explosives
will conform to applicable laws and utilize daily magazine inventory collection. In general,
blasting for the UG will utilize emulsion for upholes and ANFO for downholes and development.
All blasting will use non-electric (Nonel) detonators initiated by an electric starter as standard,
however, some more complex blasting requirements (uphole raises or complex geometries around
HMW areas) may utilize electronic and/or remote detonators.
16.1.8.7
Materials Handling
Blasted material is hauled to a remuck bay and then re-handled into a truck. Contractor quotes for
haulage are based on distance travelled, so moving material to lower levels to centralize loading
results in both increased development costs, and increased OPEX costs, hence the choice of onlevel loading. All truck loading is done with 10 t LHDs sized to provide three-pass loading of 30
t capacity haul trucks, and to allow flexibility of the haulage fleet, since the production unit on a
level can immediately load a truck once the level remuck bay is full without waiting for a different
class of LHD to arrive.
The intersection between each level access and the ramp is slashed out laterally and vertically to
create a passing bay loadout. Prior to loading (as the truck comes down-ramp), it enters the level
access and turns around to face up-ramp. It is then side-loaded from the level access while
remaining out of the line of travel for any other equipment in the ramp. Once filled, it immediately
proceeds up-ramp. Two additional passing bays without elevated backs are also developed in the
connector ramps from New Abra to Old Abra to facilitate trucks passing in these areas.
The loadout areas will require secondary support in the form of cable bolting due to their large
spans, however, operations in Canada have shown the design is effective for limiting truck travel
on-level (minimizing damage to level services from steel in the excavated rock or from the rock
itself) while essentially eliminating interruptions in ramp travel (the loadouts are wide enough to
allow two trucks to pass side-by-side).
16.1.8.8
Maintenance Facilities
Maintenance on the underground fleet is performed at a shop on surface in the initial years. As
the underground progresses deeper, it becomes less efficient to transit equipment to surface for
preventative maintenance operations, and underground maintenance facilities are installed.
Preventative maintenance bays are provided in Eagle at the 1,230 m EL (roughly 110 m below
surface) mid-way through the pre-production period. These bays are suitable for small-scale
preventative maintenance, however, do not provide facilities for major overhauls or significant
breakdown maintenance, which will still be performed on surface or offsite.
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When the mine progresses to the Old Abra area, a full underground maintenance facility is installed
on the 1,150 m EL elevation (roughly 200 m below surface) between New Abra and Old Abra at
the start of YR 3. This facility will be equipped with cranes, welding bays, wash bays, and all
other facilities necessary to service general maintenance requirements for the underground fleet.
It is still assumed that major overhauls will be performed on surface, either in the surface shop or
offsite.
Figure 16.16, above, shows the location of the underground maintenance facilities.
16.1.8.9
Backfill Reticulation
PF is initially provided through 150 mm diameter steel pipelines run down the ramp. As the mine
progresses deeper, pipelines will be run parallel to the ramp ventilation raises to shorten the transit
distance and reduce wear on the system. Steel lines will be used for all permanent and semipermanent portions of the system, with 100 mm diameter HDPE pipelines used for temporary
installations in sills and crosscuts.
The reticulation system is entirely gravity-fed, negating the need for positive displacement pumps
within the mine. Initial head may be provided from a positive displacement pump installed at the
PF plant near the Eagle portal for filling of stopes on the top levels of the mine where minimal
elevation differences and/or significant lateral transit distances occur relative to the PF plant.
Maximum flow rates within the PF reticulation system are ~125 t/h with a steady-state of 91 t/h.
It is envisioned that up to two stope areas can be actively filled at any time.
16.1.9
Personnel Requirements
Since the mine is planning to use contract labour, actual staffing of the UG operation will be
determined by the contractor for operational mining roles. Peak labour estimates from national
mining contractors vary from 54 personnel per shift to 85 personnel per shift, utilizing a 3-shift
roster at 12 hours per shift. The Authors estimate a total staff of UG operators of approximately
180 personnel, utilizing roughly 32 operators, 16 assistants/general labourers, 2 supervisors and
10 maintenance personnel per shift.
The Owner’s technical staff, comprising the engineering and geology departments, is estimated at
a peak of 32 personnel, comprised of 10 personnel in geology and grade control, 15 personnel in
engineering and surveying, 5 personnel in underground operations management and control, one
database controller and one technical services manager. This staffing level is higher than normal
for a mine of this size, due to the need for extensive QA/QC of the contractor workforce to ensure
operations are carried out to the standards of the Company, and in the areas dictated by the mine
plan. Technical staff will work varying rosters and daily hours depending on their specific roles.
Night shift coverage of operations management roles and survey roles is expected. All other roles
in technical services are expected to be day-shift only.
After an initial ramp-up in staff, peak UG staffing will occur from YR 2 to YR 8, after which it
will decline as mining areas are depleted and mining operations become more centralized. It is
anticipated that operational improvements and streamlining of the relationship between the
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contractor and the Owner’s staff will also result in gradual reduction of the Owner’s staff over
time.
16.1.10
Underground Mining Schedule
Table 16.23 shows the UG mining schedule by year. Figures 16.17 to 16.24 show the progression
of the UG mine over time.
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FIGURE 16.17
LONGITUDINAL PROJECTION OF MINE PLAN YEAR -1
Source: P&E (2025)
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FIGURE 16.18
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 1
Source: P&E (2025)
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FIGURE 16.19
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 2
Source: P&E (2025)
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FIGURE 16.20
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 3
Source: P&E (2025)
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FIGURE 16.21
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 5
Source: P&E (2025)
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FIGURE 16.22
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 8
Source: P&E (2025)
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FIGURE 16.23
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 10
Source: P&E (2025)
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FIGURE 16.24
LONGITUDINAL PROJECTION OF MINE PLAN YEAR 15
Source: P&E (2025)
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TABLE 16.23
UNDERGROUND MINING SCHEDULE
Item
Unit
Linear m
Linear m
Linear m
Development
Metres
Linear m
Eq m / d
Lin. m / d
kt
kt
Development
Ore
kt
kt
kt
kt
Stope Ore
kt
kt
kt
AgEq g/t
Total Mined Ore Au g/t
Ag g/t
Cu %
kt
kt
Stockpiling
kt
kt
kt
AgEq g/t
Process
Plant
Au g/t
Ore
Ag g/t
Cu %
k m3
kt
Backfill
kt
kt
kt
Waste
kt
kt
Descriptor
Lateral
Lateral
Lateral
Vertical
Lateral
Lateral
Mass
Mass
Mass
Mass
Mass
Mass
Mass
Mass
Mass
Grade
Grade
Grade
Grade
Mine-to-Plant
Mine-to-Stockpile
Stockpile-to-Plant
Stockpile Close
Plant Feed
Grade
Grade
Grade
Grade
Pastefill
Pastefill
Cement in PF
Tailings in PF
IS4 Development
PF Development
Waste to surface
Year
OPEX
CAPEX
Total
CAPEX
Mine-wide
Mine-wide
Eagle
New Abra
Old Abra
Total
Eagle
New Abra
Old Abra
Total
Total
Average
Average
Average
Average
Total
Total
Total
Total
Total
Average
Average
Average
Average
Total
Total
Total
Total
Total
Total
Total
-2
218
5,342
5,560
246
18.1
15.4
3.5
1.5
1.2
6.3
0.0
0.0
0.0
0.0
6.3
109.7
0.33
68.6
0.125
0.0
6.3
0.0
6.3
0.0
0.0
0.0
0.0
0.000
0.0
0.0
0.0
0.0
276.0
0.0
276.0
-1
1,056
4,782
5,838
314
17.7
16.2
8.3
8.5
0.0
16.8
0.0
0.0
0.0
0.0
16.8
111.1
0.61
51.3
0.070
0.0
16.8
0.0
23.0
0.0
0.0
0.0
0.0
0.000
0.0
0.0
0.0
0.0
258.7
0.0
258.7
1
4,550
1,832
6,382
93
18.2
17.7
74.8
70.0
6.8
151.6
193.5
141.6
1.4
336.5
488.1
429.3
2.22
221.0
0.164
488.1
0.0
15.9
7.1
504.0
419.3
2.2
215.8
0.162
116.0
232.1
20.4
169.4
123.9
0.0
123.9
Notes:
1
Totals may not sum due to rounding.
2
Maximum stockpile size occurs in YR 8, stockpile is depleted by end of YR 10.
3
At end of production, available voids are filled with low-cement PF to minimize surface storage of mine tailings.
4
Development through in-situ waste versus drifts backfilled with PF
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2
2,857
3,435
6,292
346
18.8
17.5
44.3
14.1
15.7
74.2
523.4
82.2
84.0
689.6
763.7
491.6
2.57
252.1
0.178
720.0
43.7
0.0
50.8
720.0
515.7
2.7
264.1
0.187
309.3
618.5
45.6
451.5
217.3
1.5
218.8
3
4,003
2,356
6,359
121
18.3
17.7
64.9
12.9
31.1
109.0
276.7
88.9
209.5
575.1
684.1
322.5
1.86
145.7
0.160
684.1
0.0
35.9
15.0
720.0
311.3
1.8
141.2
0.154
256.4
512.9
46.6
374.4
163.9
4.8
168.7
4
4,186
2,249
6,435
86
18.4
17.9
48.7
49.1
33.0
130.8
290.9
97.0
232.7
620.6
751.4
341.2
1.87
161.0
0.187
720.0
31.4
0.0
46.4
720.0
351.1
1.9
165.3
0.192
273.1
546.3
48.3
398.8
141.1
6.8
147.8
5
4,845
1,569
6,414
45
18.2
17.8
40.0
19.2
63.0
122.2
193.3
145.8
310.1
649.3
771.5
303.4
1.66
145.7
0.145
720.0
51.5
0.0
97.9
720.0
316.3
1.7
151.5
0.151
329.1
658.1
52.7
480.4
142.2
8.4
150.6
6
3,475
1,214
4,689
123
13.2
13.0
35.1
0.0
54.4
89.5
184.7
138.2
355.7
678.6
768.1
294.3
1.53
151.1
0.110
720.0
48.1
0.0
145.9
720.0
306.3
1.6
157.0
0.114
329.5
659.1
56.0
481.1
101.9
7.1
109.1
7
1,175
2,282
3,457
224
10.0
9.6
0.0
0.0
31.0
31.0
116.2
149.8
441.4
707.4
738.5
288.1
1.38
156.4
0.127
720.0
18.5
0.0
164.4
720.0
292.8
1.4
158.8
0.129
314.1
628.1
56.9
458.5
117.1
5.9
123.0
8
3,446
160
3,606
21
9.9
10.0
0.0
0.0
90.8
90.8
222.5
129.3
294.7
646.5
737.3
312.5
1.58
160.4
0.156
720.0
17.3
0.0
181.7
720.0
317.4
1.6
162.6
0.159
307.7
615.4
44.3
449.2
55.5
2.6
58.0
9
785
2,530
3,315
237
9.9
9.2
0.0
0.0
17.5
17.5
126.6
120.5
350.5
597.6
615.1
276.2
1.33
151.3
0.104
615.1
0.0
104.9
76.8
720.0
252.5
1.2
138.5
0.096
278.5
557.0
43.8
406.6
134.6
2.9
137.5
10
1,630
0
1,630
0
4.5
4.5
0.0
0.0
35.8
35.8
76.8
46.0
375.3
498.1
533.9
269.2
1.24
154.5
0.078
533.9
0.0
76.8
0.0
610.7
249.7
1.1
143.0
0.075
305.6
611.2
38.2
446.2
26.3
4.8
31.0
11
349
0
349
0
1.0
1.0
0.0
0.0
7.8
7.8
17.2
0.0
262.9
280.1
287.9
271.3
1.09
171.6
0.056
287.9
0.0
0.0
0.0
287.9
271.3
1.1
171.6
0.056
211.8
423.5
25.3
309.2
2.4
3.0
5.4
12
344
0
344
0
0.9
1.0
0.0
0.0
6.1
6.1
0.0
0.0
216.3
216.3
222.4
264.2
0.76
195.4
0.030
222.4
0.0
0.0
0.0
222.4
264.2
0.8
195.4
0.030
159.2
318.4
23.1
232.4
1.7
4.5
6.2
13
239
0
239
0
0.7
0.7
0.0
0.0
5.0
5.0
0.0
0.0
122.4
122.4
127.3
263.6
0.68
202.8
0.024
127.3
0.0
0.0
0.0
127.3
263.6
0.7
202.8
0.024
293.1
586.1
16.1
427.93
0.0
3.6
3.6
Total1
33,158
27,751
60,909
1,856
15.8
15.0
319.7
175.4
399.3
894.3
2,221.8
1,139.4
3,256.8
6,618.0
7,512.3
326.4
1.65
170.2
0.134
7,278.8
233.5
233.5
0.02
7,512.3
326.4
1.65
170.2
0.134
3,483.3
6,966.7
517.4
5,085.7
1,762.6
55.8
1,818.4
Development is scheduled using equivalent metres (eq m), which assumes that development effort
is generally proportional to volume excavated, and factors the work required for faces of different
sizes to an equivalent effort to advance a standard face size. For the Project, lateral development
is scheduled utilizing a nominal 4.0 mW x 4.0 mH face as one equivalent metre. Vertical
development utilizes a nominal 3.0 mW x 3.0 mL face as one equivalent metre. Scheduling for
lateral development assumes a maximum mine-wide development rate of 19 eq m (18 linear m)
per day across all types of lateral development. Vertical development is performed by different
methods and is assumed to be independent of lateral development (outside of the need for loading
and trucking).
Ore production is scheduled using a target of 2,000 tpd mine-wide (including stopes and
development sources) for material exceeding the MCOG. Additional material from development
grading less than MCOG and more than DCOG is produced in excess of this target. Mined tonnes
in excess of the available process plant capacity are stockpiled based on grade (higher grades are
prioritized to the direct process plant feed) until the plant capacity cannot be met by the mine feed,
at which point the stockpiles are included in the process plant feed. Over LOM, the maximum
closing tonnage of the stockpile in any year is 180 kt. The stockpile is depleted by the end of YR
10. Over LOM, 95% of the stockpile is comprised of development-grade material (material grading
65-150 g/t AgEq), with 5% comprised of low-grade material (material grading 150-300 g/t AgEq).
No medium or high-grade material is ever stockpiled.
While grades vary based on location, in general the ore from Eagle mining areas is higher grade
than that from the New Abra or Old Abra mining areas. The average AgEq grade of all ore from
the Eagle area is 20% higher, while New Abra is 4% lower, and Old Abra is 13% lower, than the
overall average grade of UG ore. Overall average grade of UG ore is 326 g/t AgEq.
16.2
OPEN PIT MINING
Open pit mining will occur in two pits; the Abra Pit and the Cerro Colorado pits (North and South).
The Abra open pit will occur above an active underground mining operation, which must be taken
into account during operations. The Cerro Colorado area will have no underground mining
operations.
16.2.1
Open Pit Mining Schedule
The open pit mine production schedule consists of one year of pre-stripping (Year 10) and five
years of mine production (YR 11 to YR 15).
The Abra Pit has been sub-divided into two phases; a south and north phase (shown in Figure
16.25). When pre-stripping commences in the Abra Pit, underground mining will still be occurring.
Along the north side of the Abra Pit, there is a possibility of open pit material ravelling down upon
the operating underground portal. Hence, the Abra Pit Phase 1 has been designed to mine the
southern end of the pit and avoid any interference with the underground workings. Once the
underground mine is depleted, then stripping of upper benches of Abra Phase 2 can commence.
The Cerro Colorado area consists of two separate open pits (shown in Figure 16.26). These will
be considered as different pit phases and mine sequentially.
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The annual mining sequence is shown in Table 16.24, with mining commencing at Cerro Colorado
South and Abra Phase 1. The Abra Pit will have a significant pre-strip volume and Cerro Colorado
can provide quicker access to ore to supplement underground ore.
TABLE 16.24
OPEN PIT MINING SEQUENCE
Pit Location
Cerro Colorado North
Cerro Colorado South
Abra Pit - Phase 1
Abra Pit - Phase 2
FIGURE 16.25
PHASE 1
Years of
Operation
2
2
5
3
YR 10
YR 11
1
1
1
1
YR 12
YR 13
1
1
1
1
1
YR 14
YR 15
1
1
1
ABRA PHASE 1 AND PHASE 2 OPEN PITS
PHASE 2 (FINAL PIT)
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FIGURE 16.26
CERRO COLORADO OPEN PITS
The life-of mine open pit production schedule is shown in Table 16.25. The peak total mining rate
is 5.0 million tonnes per year occurring in the Abra Phase 1 pre-strip year. Gradually the mining
rate decreases as the pits intersect more ore and require less waste removal.
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TABLE 16.25
OPEN PIT PRODUCTION SCHEDULE
Pit Phase
Type
Total (kt)
Waste (kt)
Strip Ratio (w:o)
Ore (kt)
Total
Au (g/t)
Pit
Ag (g/t)
Cu (%)
AuEq (g/t)
AgEq (g/t)
Total (kt)
Waste (kt)
Strip Ratio (w:o)
Ore (kt)
Cerro
Colorado Au (g/t)
North
Ag (g/t)
Cu (%)
AuEq (g/t)
AgEq (g/t)
Total (kt)
Waste (kt)
Strip Ratio (w:o)
Ore (kt)
Cerro
Colorado Au (g/t)
South
Ag (g/t)
Cu (%)
AuEq (g/t)
AgEq (g/t)
Total (kt)
Waste (kt)
Strip Ratio (w:o)
Ore (kt)
Abra
Au (g/t)
Phase 1
Ag (g/t)
Cu (%)
AuEq (g/t)
AgEq (g/t)
Total
23,274
20,553
7.6
2,721
0.66
76.9
0.02
1.61
135.9
3,014
2,819
14.4
195
0.76
26.3
0.01
1.06
91.9
2,222
1,860
5.1
362
0.83
45.3
0.01
1.39
117.9
11,938
10,533
7.5
1,404
0.69
95.3
0.01
1.85
156.4
Year
10
5,000
4,891
44.7
109
0.79
50.1
0.02
1.43
119.5
-
11
4,750
4,318
10.0
432
0.70
59.1
0.01
1.42
120.2
-
1,012
903
8.3
109
0.79
50.1
0.02
1.43
119.5
3,988
3,988
-
1,210
957
3.8
253
0.85
43.3
0.01
1.38
117.1
3,540
3,360
18.7
179
0.49
81.5
0.01
1.48
124.6
12
3,500
3,002
6.0
498
0.59
82.1
0.01
1.59
134.0
1,372
1,322
26.6
50
0.59
28.4
0.01
0.92
79.1
2,128
1,681
3.8
448
0.59
88.0
0.01
1.66
140.1
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13
4,000
3,407
5.7
593
0.79
79.1
0.02
1.75
148.5
1,643
1,497
10.3
146
0.82
25.6
0.01
1.11
96.2
1,497
1,050
2.3
447
0.78
96.5
0.02
1.96
165.5
14
4,000
3,280
4.6
720
0.69
100.6
0.04
1.95
164.5
-
15
2,024
1,655
4.5
370
0.39
48.9
0.04
1.02
86.0
-
784
454
1.4
330
0.82
111.2
0.02
2.17
183.5
-
Page 368 of 577
TABLE 16.25
OPEN PIT PRODUCTION SCHEDULE
Pit Phase
Type
Total
Abra
Phase 2
Total (kt)
Waste (kt)
Strip Ratio (w:o)
Ore (kt)
Au (g/t)
Ag (g/t)
Cu (%)
AuEq (g/t)
AgEq (g/t)
6,100
5,340
7.0
759
0.49
70.9
0.05
1.40
118.0
16.2.2
Year
10
-
11
-
12
-
13
860
860
-
14
3,216
2,826
7.2
390
0.58
91.7
0.06
1.76
148.4
15
2,024
1,655
4.5
370
0.39
48.9
0.04
1.02
86.0
Open Pit Mining Practices
It is assumed that the Los Ricos South mine will be operated as a contracted conventional open pit
mining operation. While owner-operated mining may be an option, this was not considered in this
study. Numerous other mines in Mexico rely on the use of contract mining, with many experienced
contractors being available at competitive rates. Contractor budgetary quotations for mining the
Los Ricos South Deposits were obtained for this FS.
The mining contractor will undertake all drilling, loading, hauling, and mine site maintenance
activities. A separate contractor will provide blasthole loading and charging services. The Owner
will provide overall mine management and technical services, such as mine planning, grade
control, geotechnical, and surveying services.
16.2.3
Contract Mining
It is anticipated that the mining operations would be conducted 24 hours per day and 7 days per
week throughout the entire year.
The exact equipment fleet to be used by the mining contractor has not yet been defined in detail at
this stage. This will depend on the types of equipment the contractor has in its fleet once a formal
mining contact is enacted. However, based on current information, it is expected that small
hydraulic excavators and diesel-powered front-end loaders will be used, suitable for mining 5 m
high benches. The anticipated truck size is 60 t, similar to the CAT 773, although alternate truck
sizes may eventually be used.
The primary mining operation will be supported by the contractor’s fleet of equipment consisting
of dozers, road graders, watering trucks, maintenance vehicles, and service vehicles. The projected
mining equipment fleet is shown in Table 16.26.
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TABLE 16.26
CONTRACTOR OPEN PIT MINING FLEET
Contractor Fleet
Atlas DM45 Drill
CAT 395 Backhoe Excavator
CAT 988 Wheel Loader
CAT 773 Truck
CAT D8T Dozer
CAT 16M Grader
CAT 773 Water Truck
Units
2
1
1
5
2
1
1
It is assumed that most of the materials mined will require drilling and blasting to some degree.
The blasting contractor will provide the hole charging services, and it is anticipated that blasting
of the rock will be carried out using an ammonium nitrate fuel oil mixture (“ANFO”).
Portions of the deeper pits may experience groundwater seepage. No quantitative information was
available to adequately predict the expected water inflow into the open pits. The mining contractor
will be responsible to supply and operate pit dewatering pumps and pipelines to keep the open pits
dry and operable. The underground mine beneath the larger Abra Pit may also act as a drainage
gallery keeping the open pit relatively dry.
16.2.4
Owner’s Mining Team
The mine Owner will be responsible for providing contract management and overall supervision
of the mining contractor. The Owner will also provide technical services, such as mine planning
and scheduling, geotechnical engineering, grade control and surveying. Table 16.27 lists the
personnel on the Owner’s mining team.
TABLE 16.27
OWNER’S MINING TEAM
Owner Mining Team
Mine Superintendent
Foreman
Chief Engineer
Mine Clerk
Mine Engineer
Geologists
Grade Control Technicians
Surveyor
Survey Technician
Total Team
Number
1
2
1
1
2
2
2
1
1
13
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16.2.5
Waste Rock Storage Facilities
Each of the open pits will require the development of one or more waste rock storage facilities.
Some of the waste rock will be placed into hill-side facilities adjacent to the open pits. The
locations of the waste facilities are shown on the site general layout drawing in Section 18 (Figure
18.1).
Some of the waste rock will also be used to build new haul roads and widen existing roads as
necessary.
16.2.6
Mine Support Facilities
The Los Ricos South open pits will require mine offices, maintenance facilities, warehousing, and
storage areas. Some of these will already be in place for the underground mining operations.
A truck maintenance shop area will be provided for the mining contractor. A fuel and lube station
will also be provided for fuelling the mining fleet.
16.3
LIFE-OF-MINE PRODUCTION SCHEDULE
Table 16.28 summarizes the life of mine processing schedule. An ore stockpile located near the
primary crusher is utilized throughout the mine life to maximize grade and ensure constant process
plant throughput.
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TABLE 16.28
PROCESSING SCHEDULE (LIFE OF MINE)
Feed
Type
Unit
Total
UG Ore Mined
Au
Ag
Cu
OP Ore Mined
Au
Ag
Cu
Processed1
Au
Ag
Cu
AgEq
AgEq
kt
g/t
g/t
%
kt
g/t
g/t
%
kt
g/t
g/t
%
g/t
koz
7,512
1.65
170
0.13
2,721
0.66
77
0.02
10,234
1.39
145
0.10
276
79,889
1
511
2.15
214
0.16
504
2.17
216
0.16
419
6,027
2
764
2.57
252
0.18
720
2.70
264
0.19
516
10,637
3
684
1.86
146
0.16
720
1.79
141
0.15
311
6,351
4
751
1.87
161
0.19
720
1.93
165
0.19
351
7,161
5
771
1.66
146
0.14
720
1.73
152
0.15
316
6,450
6
768
1.53
151
0.11
720
1.60
157
0.11
306
6,251
Notes:
1. Includes stockpile movements.
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7
738
1.38
156
0.13
720
1.40
159
0.13
293
5,945
Year
8
737
1.58
160
0.16
720
1.61
163
0.16
317
6,445
9
615
1.33
151
0.10
720
1.21
139
0.10
253
5,114
10
534
1.24
154
0.08
109
0.79
50
0.02
720
1.09
129
0.07
230
4,672
11
288
1.09
172
0.06
432
0.70
59
0.01
720
0.86
104
0.03
181
3,686
12
222
0.76
195
0.03
498
0.59
82
0.01
720
0.65
117
0.02
174
3,525
13
127
0.68
203
0.02
593
0.79
79
0.02
720
0.77
101
0.02
169
3,437
14
720
0.69
101
0.04
720
0.69
101
0.04
165
3,323
15
370
0.39
49
0.04
370
0.39
49
0.04
86
865
17.0
RECOVERY METHODS
17.1
OVERVIEW
Ausenco has conceived a flexible and robust flowsheet, which is designed to treat a variety of
grades. The flowsheet is based on metallurgical testwork, and Ausenco’s expertise and experience
with similar plants and local operations. The flowsheet is based on well-proven unit operations in
the industry and there are no unique or novel processing methods required for gold and silver
recovery.
The Los Ricos South (“LRS”) process plant is designed to treat 2,000 tpd of ore to produce doré
as final product and a copper sulphide concentrate secondary product. The process plant will be
located at the mine site and will receive run of mine (“ROM”) ore from the underground mine.
The key project design criteria for the plant are:
•
Major equipment designed for nominal throughput of 2,000 tpd; and
•
Crushing circuit, process plant and tailings filtration area availabilities of 70%, 92%
and 85% respectively.
The process flowsheet includes the following unit operations:
•
•
•
•
•
•
•
•
•
•
•
•
17.2
Primary crushing;
Grinding (single stage SAG mill circuit);
Pre-leach thickening;
Cyanide leach;
Counter-Current Decantation (“CCD”);
Clarification filters, de-aeration and zinc precipitation (Merrill Crowe);
Mercury retorting;
Acid digestion to remove copper from Merrill Crowe (“MC”) precipitate;
Refining to produce doré;
Sulphidization, acidification, recycle and thickening (“SART” process)
Cyanide detoxification of tailings; and
Tailings thickening and pressure filtration;
PROCESS FLOWSHEET
The LRS FS overall process flow diagram is shown in Figure 17.1. Details of the flowsheet and
major process equipment are described in the following sections.
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FIGURE 17.1
LRS OVERALL PROCESS FLOWSHEET
Source: Ausenco (2025)
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17.3
KEY PROCESS DESIGN CRITERIA
The key process design criteria listed in Table 17.1 form the basis of the process flowsheet design
and selection of mechanical equipment.
TABLE 17.1
KEY PROCESS DESIGN CRITERIA
Description
Annual Throughput
Daily Throughput
Operating Availability – Primary Crushing
Operating Availability – Grinding, Leach, CCD,
Merrill Crowe, SART, Cyanide Detox
Operating Availability – Tailings Filtration
Nominal / Design Head Grade, Gold (Au)
Nominal / Design Head Grade, Silver (Ag)
Nominal / Design Head Grade, Copper (Cu)
Nominal / Design Overall Recovery, Au
Nominal / Design Overall Recovery, Ag
Nominal / Design Overall Recovery, Cu
SMC parameters, Axb
Bond Ball Mill Work Index (BWi), Design
Bond Abrasion Index (Ai), Design
ROM Maximum Feed Size, 100% Passing
ROM Feed Size, 80% Passing
Crushing Circuit Product Size, 80% Passing
Grinding Circuit Product Size, 80% Passing
Grinding Cyclone Of Solids Density
Leach Feed Solids Density
Leach Circuit Residence Time
Number of CCD Wash Stages
CCD wash efficiency Au / Ag
CCD Wash Ratio (wash water:feed solution)
Metal Recovery Method
Nominal / Design Zinc (Zn) addition rate
MC Au Recovery
MC Ag Recovery
Nominal / Max MC Cu Recovery
Acid Digestion Circuit Cu removal
Acid Digestion Circuit Zn removal
Units
Value
Mtpa
tpd
%
0.72
2,000
70
%
92
%
g/t
g/t
%
%
%
%
kWh/t
g
mm
mm
mm
µm
% w/w solids
% w/w solids
h
%
Zn:Au+Ag
%
%
%
%
%
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85
1.39 / 2.98
145 / 344
0.10 / 0.25
92.7 / 93.1
85.6 / 87. 3
72.6 / 75.4
37.8
19.0
0.69
560
250
72
74
35
55
96
5
99.3 / 99.7
3
Merrill Crowe (MC)
5 / 10
98.0
98.5
12.8 / 20.0
90.0
90.0
Page 375 of 577
TABLE 17.1
KEY PROCESS DESIGN CRITERIA
Description
Units
Value
SART Plant Au Recovery to Precipitate
SART Plant Ag Recovery to Precipitate
SART Plant Cu Recovery to Precipitate
%
%
%
0.9
90.0
90.0
SART Plant Zn Recovery to Precipitate
%
90.0
Cyanide Detoxification Method
Cyanide Detoxification Residence Time
Cyanide detoxification CNWAD target
Tailings Filter Cake Moisture
Tailings Disposal Method
mins
mg/L
% w/w solids
—
SO2/air
120
<5
16.0
Dry Stacking
The process plant layout is presented in Figure 17.2.
FIGURE 17.2
OVERALL PROCESS PLANT LAYOUT
Source: Ausenco (2025)
17.4
PLANT DESCRIPTION
17.4.1
Crushing Area
A conventional jaw crusher reduces ROM feed particle size to an 80% passing size (k80) of 72 mm,
suitable for feeding a single stage SAG mill (Figure 17.3). The nominal throughput of the crushing
circuit is approximately 119 t/h at 70% availability.
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The crushing circuit major equipment will include:
•
•
•
•
•
•
•
Static grizzly and hopper;
Apron feeder;
Jaw crusher (112 kW);
Surge bin;
Surge bin belt feeder to reclaim crushed material to feed the SAG mill;
Emergency stockpile and reclaim hopper/feeder; and
Associated material handling systems (conveyors, weightometers and tramp magnet).
FIGURE 17.3
CRUSHING AREA LAYOUT
Source: Ausenco (2025)
ROM ore will be trucked from the underground and open pit mine either to the ROM pad stockpile
or directly onto the static grizzly hopper. ROM ore from the pad stockpile will be reclaimed using
a front-end loader (“FEL”) and dumped into the static grizzly hopper. The jaw crusher will be a
Metso C106 (or equivalent) with a closed side setting (“CSS”) of 70 mm and will crush the ROM
ore from a k80 of 250 mm to k80 of 72 mm. Crusher discharge will be conveyed to the surge bin
via the primary crusher product conveyor. A tramp metal magnet will be installed at the head end
of this conveyor to remove tramp metal. The tramp metal will be manually removed from the
magnet as needed.
The surge bin will have a live capacity of 10 minutes. Crushed ore will be reclaimed via a belt
feeder beneath the surge bin and conveyed to the SAG mill feed chute by the SAG mill feed
conveyor.
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Surge bin overflow will be transferred to an emergency stockpile via the emergency stockpile
conveyor. When the primary crusher is unavailable, crushed ore will be reclaimed from the
stockpile using a FEL and dumped into the emergency reclaim hopper. The emergency reclaim
feeder will transfer crushed ore from the emergency hopper onto the SAG mill feed conveyor. The
emergency stockpile will provide 16 hours of storage.
Fresh water will be used for dust suppression in the crushing area.
17.4.2
Grinding
A conventional SAG mill, arranged in closed circuit with a cyclone cluster, was selected to reduce
the ore from a k80 of 72 mm to k80 of 74 µm (Figure 17.4). The nominal feed throughput of the
grinding circuit is approximately 91 t/hr, based on 92.0% availability.
The grinding circuit will include:
•
One SAG mill, 7.3 m (24 ft) in diameter by 4.27 m (14 ft) in length, powered by a
3,600-kW variable speed drive motor;
•
Duty/standby SAG mill cyclone feed pumps (55 kW each);
•
One cyclone cluster with ten 250 mm cyclones, eight in operation and two in standby;
and
•
Associated material handling and storage systems (sump pumps, pump boxes, bins).
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FIGURE 17.4
CRUSHING AND GRINDING LAYOUT
Source: Ausenco (2025)
Crushed ore will be reclaimed from the surge bin onto the SAG mill feed conveyor and discharged
into the feed chute of the SAG mill. The SAG mill will be a grate discharge type mill with no
recycling of pebbles.
The SAG mill product will discharge onto a trommel screen. Trommel screen undersize will
gravitate to the cyclone feed pumpbox and then get pumped to the cyclone cluster and the trommel
screen oversize reports to a scats bunker. Barren solution, high in cyanide concentration, will be
added to the SAG mill feed chute and cyclone feed pumpbox to maintain a target mill discharge
slurry solids density. The cyclone underflow will return to the SAG mill feed. The nominal
circulating load is 450%. The cyclone overflow with a particle size of k80 of 74 µm will report to
the pre-leach thickener after flowing through a trash screen.
A sump pump will service the area. Grinding media for the SAG mill will be introduced by use of
a dedicated kibble and a grinding building jib crane.
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17.4.3
Leaching
17.4.3.1
Pre-Leach Thickening
The trash screen underflow gravitates to the pre-leach thickener, where it is thickened from a
density of 35% w/w solids to 55% w/w solids in the thickener underflow. The pre-leach thickener
underflow is pumped to the leaching tanks and the pre-leach thickener overflow reports to process
water tank for distribution around the process plant. Diluted flocculant is added to the thickener
feed to assist with solids settling within the thickener.
The area will include the following major equipment and facilities:
•
•
One 16 m diameter high-rate thickener; and
Associated material handling systems (pumps, pump boxes, sump pumps).
A sump pump will be provided in the pre-leach thickener area to facilitate clean up. Spillage is
pumped to the pre-leach thickener.
17.4.3.2
Leaching
The pre-leach thickener underflow will be leached in four mechanically agitated leach tanks
operating in series (Figure 17.5).
The leaching area will include:
•
•
Four 15.4 m diameter x 16.8 m high leach tanks; and
Associated material handling and storage systems (agitators, pumps, sump pumps,
pump boxes).
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FIGURE 17.5
LEACH CIRCUIT LAYOUT
Source: Ausenco (2025)
The leaching circuit will operate continuously; leaching reagents, including sodium cyanide will
be added to facilitate gold and silver extraction. The operating pH of the leach circuit will be
maintained between 10.5 and 11.0 with additions of milk of lime, to drive leach kinetics, to limit
corrosion, and to prevent the loss of cyanide to gaseous hydrogen cyanide.
Gold and silver leaching will occur in a series of four tanks, providing over 96 hours of total
residence time. The target solids density for the leach circuit will be 55% w/w solids. The estimated
leach circuit extraction will nominally be 93% for gold and approximately 86% for silver.
Leach discharge slurry will gravitate to the CDD circuit.
Air will be supplied from the leach and detox blowers and delivered via the tank agitator shafts.
To allow for maintenance of individual leach tanks, the circuit will be configured with a provision
which allows for slurry to bypass any single leach tank.
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17.4.4
Counter-Current Decantation (CCD) Circuit
A five-stage CCD washing circuit and a clarifier will be used to recover pregnant solution from
the cyanide leached slurry (Figure 17.6).
The washing circuit will include:
•
•
•
•
Five 16 m diameter high rate thickeners;
One 19 m diameter clarifier;
Pregnant solution storage tank; and
Associated material handling and storage systems (feed boxes, pumps, sump pumps,
pump boxes).
FIGURE 17.6
CCD CIRCUIT LAYOUT
Source: Ausenco (2025)
Leached slurry will gravitate to the first CCD thickener, after mixing with the overflow of the
second CCD thickener. Underflow from the first thickener will be fed to the subsequent CCD
thickener after mixing with the overflow of the next downstream thickener. The process will repeat
until the solids flow reports to the last CCD thickener (CCD No. 5). The underflow of CCD No. 5
will be pumped to a cyanide detoxification circuit as washed tailings. Barren solution and process
water will be added to CCD No. 5 as process wash water. Overflow from the first CCD thickener
will flow to the clarifier feed box. Pregnant solution will be clarified in the clarifier prior to storage
in the pregnant solution tank which will feed the Merrill Crowe circuit. The clarifier underflow
will be pumped to the CCD circuit (CCD No. 1).
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The wash ratio, wash solution volume to feed solution volume, will be 3:1 to achieve an overall
CCD wash efficiency greater than 99%.
Settling of solids will be aided by the addition of diluted flocculant at each CCD thickener and
coagulant to the clarifier.
Antiscalant will be added to the pregnant solution tank as required to inhibit scale formation in the
Merrill Crowe circuit.
One sump pump will be provided in the CCD area to return spillage to the circuit.
17.4.5
Merril Crowe Circuit
Pregnant solution from the clarifier overflow will be stored in the pregnant solution tank. Clarified
pregnant solution will be processed through the Merrill Crowe process, zinc-dust cementation, to
recover the contained precious metals. Merrill Crowe barren solution will be directed to the SART
(sulphidization, acidification, recycling, and thickening) circuit to precipitate dissolved copper and
regenerate sodium cyanide for reuse in the leach circuit.
The Merrill Crowe circuit will be provided as a vendor package, and will include:
•
•
•
•
•
•
•
•
•
Two rotating pressure leaf clarifiers;
One de-aeration tower;
One air/water separator;
One de-aeration tower vacuum pump;
One zinc mixing cone, including a hopper and a feeder;
Two precipitation filter press units;
One pre-coat preparation tank;
One body feed preparation tank; and
Associated material handling and storage systems (pumps, sump pumps, pump boxes,
feeders).
Pregnant solution from the clarifier will be discharged to the pregnant solution tank which will
provide approximately 1.5 hours of surge capacity.
A second stage of clarification is required to reduce the suspended solids content to <5 mg/L for
efficient zinc precipitation. The clarifying filter feed pump will pump the pregnant solution from
the pregnant solution tank to rotating pressure leaf clarifiers to remove any residual solids. Two
filters will be provided in a duty/standby arrangement. A diatomaceous earth slurry pre-coat step
enhances capture of the fine solids at the start of each cycle. Diatomaceous earth slurry is also
added throughout a cycle to maintain filter cake porosity. At the end of the filtration cycle, pressure
leaf clarifier sludge will be pumped back to the CCD circuit via a sump pump, to minimize any
losses of precious metals in the entrained solution.
Filtrate from the rotating pressure leaf clarifiers will feed the de-aeration tower. Dissolved oxygen
will be removed under vacuum. De-aerated pregnant solution will be mixed with the zinc dust and
lead nitrate and pumped to the precipitate filters using a precipitate filter feed pump. Zinc dust will
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be slurried with process water in a zinc mixing cone. Cyanide can be added to the process as
required, to maintain adequate free cyanide for the precipitation reaction. Lead nitrate solution will
also be added to improve the precipitation efficiency. The precipitate filter feed pump will be a
horizontal centrifugal pump with mechanical seals such that air cannot enter the system.
The precipitate filters will be recessed plate filter presses furnished with filter cloths.
Diatomaceous earth slurry will be used as a pre-coat at the beginning of the filter cycle to prevent
cloth blinding and body feed will be added to provide acceptable filtration rates. The filters will
typically be operated in a duty/standby configuration and operated until the pressure reaches a
predetermined value. Filtrate will report to the precipitate filter filtrate tank and then will be
pumped to the SART circuit.
On a scheduled basis the operating filter press is bypassed and the spare filter put online. The filter
press will be drained and compressed air will be used to further dewater the cake. The filter cake,
containing approximately 12% w/w solids of precious metals, will be dropped onto precipitate
carts for transfer to the doré room for smelting.
The zinc mixing cone and precipitate filters will be located in a secured, closed room within the
gold room building.
17.4.6
Refinery
Zinc precipitate from the Merrill Crowe circuit will be loaded into a mercury retort for removal of
mercury, undergo acid digestion to dissolve residual levels of copper and zinc (as required) and
then be smelted into gold–silver doré. The acid digestion and smelting process will be performed
in batch mode. The circuit will be in a secure enclosed area with closed circuit television (“CCTV”)
cameras and restricted access. The gold room was designed to be able to manage the volume of
doré to be produced at design grades for gold and silver.
The smelting circuit will be a vendor package, and the main equipment will include:
•
•
•
•
•
•
•
One 216 kW (1.1 m3) electric retort and adsorption skid;
One acid digestion plant including one acid digestion tank and one precipitation filter;
One 320 kW, (0.28 m3) induction furnace;
Flux dosing and flux mixer system;
One gold–silver doré safe;
Mechanized slag handling; and
Associated material handling and other systems (molds, dryers, dust collection system).
Damp Merrill Crowe precipitate will be loaded into the mercury retort for removal of mercury.
The mercury retort will include the retort oven, condenser, mercury trap, sulphur-impregnated
carbon adsorber, and a vacuum pump with seal water separator. Mercury will be collected in a
mercury trap and decanted into a mercury flask. Mercury collected will be treated off site for
storage or disposal.
Cooled retorted precipitate will be processed as required through the acid digestion circuit to
dissolve copper and zinc in the Merrill Crowe precipitate. The acid digestion step is a batch
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process, where dried precipitate filter cake is added with process water and sulphuric acid in an
acid digestion tank to allow copper and zinc to dissolve into solution. After an adequate dissolution
time, the digested slurry is fed to the acid digestion precipitate filter to dewater the precipitate to
produce a filter cake and produce a copper and zinc rich filtrate. Sodium hydroxide is added to the
filter press feed line to neutralize the acidic solution, so the produced filter cake is safe to handle.
The acid digestion step in the gold room can be bypassed when the copper and zinc levels in the
Merrill Crowe precipitate decrease, which is anticipated in the later years of production when the
copper plant feed head grade drops.
After acid digestion, the filter cake will be dried and mixed with fluxes and loaded into the electric
furnace for smelting. The fluxes will react with base metal oxides to form a slag, whilst the gold
and silver will remain as molten metal. The molten metal will be poured into 40 kg moulds, to
form doré ingots at nominal composition of approximately 0.8-1.1% Au and 83-91% Ag, and other
impurities including copper and zinc. The doré bars will be cleaned, assayed, stamped, and stored
in a secure vault prior to shipment to refineries.
Slag generated from refining will be handled through a mechanized slag handling system. Once
solidified, slag will be tipped from the slag pot onto spikes and broken slag will be collected in a
bin underneath. The slag bin will be removed by forklift and recycled back to the SAG mill.
The gold room will operate seven days a week, for 12 hours each day.
Sufficient ventilation and off-gas handling will be provided in the gold room for a healthy work
environment. Fume and dust exposure for the melting furnace and material handling will be
controlled through a ventilation system installed in the gold room, including hoods, enclosures,
and fans to follow local regulations/guidelines.
A sump pump, complete with a precious metals trap, will be installed in the gold room to remove
mercury retort condenser return water, scrubber liquid and any hose-down or spillage, and return
it to clarifier feed box.
17.4.7
SART Plant
The gold-silver-copper-bearing solution is fed to a conventional Merrill Crowe plant for the
recovery of the gold and silver precipitate and subsequent doré. Discharge from the Merrill Crowe
plant feeds the SART (Sulphidization, Acidification, Recycling and Thickening) process to
recover a saleable Cu2S precipitate and regenerate the soluble copper associated cyanide. The high
cyanide solution from the SART is recycled for repulping prior to leaching. Process water recycled
from the overall plant including thickener and chamber filters is used for the plant’s required
process water. Make-up water is provided by a single surface dam and is pumped to the process
plant facility.
17.4.7.1
SART Circuit
The SART feed solution is pumped at a nominal rate of 278 m3/hr with a design copper and zinc
concentration of 826 mg/L and 137 mg/L. Based on three test campaigns, and internal MetSim
testwork, the free cyanide as NaCN in solution is estimated to be in the range of 2,174 mg/L. The
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SART technology is well established and commercially utilized on multiple mine sites throughout
Mexico, South America, and China. Presently, GoGold operates a SART circuit at its Parral Heap
Leach mine in Parral, CH, Mexico.
The solution feed enters the circuit where sodium hydrosulphide (“NaHS”) is injected prior to the
copper contactor. The solution next enters the copper contactor and acidified with sulphuric acid
(“H2SO4”) to a controlled pH pf 4.5. This allows for:
•
•
Effective precipitation of copper sulphide (Cu2S); and
Regeneration of cyanide associated with soluble copper.
The resultant copper precipitate (Cu: 56% design, 61% nominal, Zn: 11% design, 10% nominal)
overflows to the copper clarifier to allow for solid / liquid clarification. The underflow is pumped
to the copper filter press and dewatered to 50% moisture, impounded, and shipped as a saleable
product to a smelter as discussed in Section 19 regarding marketable saleable products.
The clarifier overflow acidified solution is pumped to the neutralization circuit where lime solution
is added to raise the pH to 10.5. During this process, some gypsum is produced and removed in
the downstream gypsum clarifier. The resultant gypsum produced is pumped to the planned inplant detoxification system for overall plant tailings dewatering and dry stacking lined
impoundment area. The overflow solution from gypsum clarifier is pumped back to the process.
The major equipment in this area includes:
•
•
•
•
•
•
•
•
•
•
•
•
•
3.6 m diameter x 4.1 m high copper 304 SS contactor, 6 kW agitator;
10 m diameter copper clarifier 316 SS and totally covered;
Copper concentrate (Cu2S) horizontal filter press, 1,250 mm (47) recessed plates;
Two 6.2 m dia. x 7.7 m high Neutralization mild steel tanks, 12 kW agitators;
10.0 m diameter gypsum clarifier, mild steel;
HCN Gas Scrubber with Caustic Addition System to Monitor and Control
HCN and H2S Concentration <2.5 ppm and <5.0 ppm;
Reagent Mixing and Storage Areas
Acid tank,
NaHS tank for Liquid NaHS and Caustic Mixing,
Other SART Reagents (lime, flocculant) fed by Main Plant Distribution System;
SART Battery Storage Tanks; and
Dedicated SART PLC Circuit Control System.
The SART facility will have its own water distribution system and will feed off the main plant
water distribution system. Air requirements for the SART plant are fed from the main
compressor/dryer system. A simplistic process flow diagram (“PFD”) is shown in Figure 17.7.
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FIGURE 17.7
SART PROCESS FLOW DIAGRAM
Source: BQE Water (2025)
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Table17.2 details the SART process design criteria.
TABLE 17.2
SART MAJOR PROCESS DESIGN CRITERIA
Criteria
Plant Operations
Availability
Plant Feed Flowrate
Design Copper Quality
Design Zinc Quality
Target pH for Copper Contactor
Hydraulic Retention Time
Copper Removal
Zinc Removal
Silver Removal
Copper Clarifier Underflow Solids
Re-seed Ratio
Copper Clarifier Flocculant Dose
Cu2S Filter Press Plate Size
Filter Press Precent Solids
Target pH for Neutralization Contactors
Hydraulic Retention Time, each
Gypsum Clarifier Flocculant Dose
Gypsum Clarifier Underflow Solids
Units
%
m3/hr
mg/L
mg/L
pH
min
%
%
%
%
%
mg/L
mm
%
pH
min
mg/L
%
Value
92
278
826
137
4.5
10
90
90
90
20
83
2.0
1,200
50
10.5
45
2.0
20
Source: BQE Water (2025) and Section 13
The SART facility for the Los Ricos South Project is an integral part of the overall process facility
as discussed and shown in this and other sections of this Technical Report. Figure 17.8 illustrates
the proposed general arrangement for the SART facility.
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FIGURE 17.8
SART PLANT GENERAL LAYOUT
Source: D.E.N.M (2025) and BQE Water (2025)
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17.4.8
Cyanide Detoxification
The washed leach residue slurry from the CCD circuit will be treated using a sulphur dioxide
SO2/air process to reduce the weak acid dissociable (CNWAD) cyanide concentration to <5 mg/L.
The cyanide detoxification circuit will include:
•
Two cyanide detoxification reaction tanks of 5.5 m in diameter x 8 m high, operating
in parallel; and
•
Associated material handling systems (pumps, pump boxes, sump pumps).
Thickened, washed tailings slurry from the final CCD thickener, with solids concentration of
approximately 55%, will be pumped to the cyanide detoxification tanks. In the SO2/air process,
sodium metabisulphite (“SMBS”), air, copper sulphate (catalyst) and milk of lime will be added
to oxidize free cyanide and CNWAD to cyanate, thereby reducing the CNWAD concentration to the
target level prior to filtration and dry stacking of the tailing’s solids. The cyanide detoxification
circuit will consist of two mechanically agitated tanks, each providing a residence time of two
hours.
Air will be provided from the blowers as required and will be added to the tanks via agitator shafts.
CNWAD levels of the cyanide detoxification discharge will be measured by an on-stream analyzer.
The detoxified tailings will be pumped to a thickener to thicken the slurry prior to filtration and
dry stacking of the final solids.
The cyanide detoxification circuit will be serviced by a dedicated sump pump. Any spillage within
this area will be returned to the cyanide detoxification feed box.
17.4.9
Tailings Dewatering
Detoxified tailings will be thickened and filtered, and the filtered solids will be impounded in an
on-site storage facility (Figure 17.9).
The tailings circuit will include:
•
•
•
•
•
•
One 19 m diameter high rate thickener;
One tailings filter feed tank providing 12 hours of storage time;
One Metso 3512 FFP plate and frame filter;
One 5.0 m diameter x 6.0 m high filter filtrate tank;
One 4.1 m diameter x 4.9 m high cloth wash water tank; and
Associated material handling systems (pumps, pump boxes, sump pumps).
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FIGURE 17.9
TAILINGS FILTRATION LAYOUT
Source: Ausenco (2025)
Thickener underflow slurry, at approximately 59% w/w solids, will be pumped into an agitated
filter feed tank, prior to being pumped to a filtration circuit for further dewatering. This tank will
provide approximately 12 hours of surge capacity between the thickener and filter. One Metso FFP
3512 68/74 plate and frame filter was selected to dewater the tailings and produce a filter cake
with a moisture of 16.0%. The filter has been sized to provide sufficient processing catch-up
capacity; however, an additional 6 plates can be retrofitted to the machine during operation to
provide an additional 9% of filtration area if required.
Tailings filter cake will discharge into a bunker where it will be loaded onto trucks for transport
and disposal at the dry stack tailings storage facility (“DSTF”). Filtrate will be recycled back to
the tailing thickener, where it will report to the tailing’s thickener overflow stream. The tailings
thickener overflow will be used for filter cloth wash water, make-up water for the cyanide and
lime reagent systems with excess water ultimately reporting to the process water tank for
distribution throughout the process facilities. Any excess process water will be bled from the plant
as required.
Any spillage within this area will be returned to the sump pump in the cyanide detoxification area,
and in turn will be pumped to a cyanide detoxification feed box.
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17.5
REAGENTS HANDLING AND STORAGE
The mixing and storage area for each reagent will be located proximate to various additional points
throughout the plant. Some reagents will be delivered in 25 kg bags and manually handled from a
pallet to the bag breaker for mixing and further storage. Reagents delivered in bulk bags will be
moved from storage to the mixing area by forklift. Electric hoists servicing in the reagent area will
lift the reagents to the respective reagent bag breaker that will be located above the reagent mixing
area.
The reagent handling systems will typically include unloading and storage facilities, mixing tanks,
stock tanks, transfer pumps, and feeding equipment. Each reagent mixing and storage system will
be located within specific containment areas to prevent incompatible reagents from mixing and to
contain any spillage. Storage tanks, where used, will be equipped with level indicators and alarms
to ensure spills do not occur during filling or normal operation. Appropriate ventilation, fire and
safety protection, eyewash stations and Safety Data Sheet (“SDS”) stations will be provided for
the reagent facilities. Sumps and sump pumps will be provided for spillage control.
Reagents used in the process plant are outlined in Table 17.3, together with their anticipated LOM
nominal annual consumptions. The preparation method, storage and delivery details of each main
chemical is outlined in the following sections.
TABLE 17.3
REAGENT LOM AVERAGE ANNUAL CONSUMPTIONS
Reagent
(minimum purity)
Lime (Hydrated) (90%)
Sodium Cyanide (98%)
Flocculant
Sodium Metabisulphate (67%)
Copper Sulphate (pentahydrate) (99.5%)
Coagulant (100%)
Lead Nitrate (99%)
Antiscalant
Diamataceous Earth
Zinc Powder
Sodium Hydroxide (50% w/w)
Sodium Hydrosulphide (42% w/w)
Sulphuric Acid (98% w/w)
Silica
Borax
Sodium Nitrate
Sodium Carbonate
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LOM Average Annual
Consumption
(tpa)
4,767
1,036
195
2,267
11
56
53
20
252
211
343
996
4,482
996
228
47
294
Page 392 of 577
17.5.1
Hydrated lime
Hydrated lime will be delivered to the plant in regular 26 t bulk shipments and received in a 33-t
storage silo, which, at design operating rates, will provide one day of storage. Tailings thickener
overflow water and hydrated lime will be metered into the agitated lime storage tank to produce a
hydrated lime slurry at approximately 20% w/w solids. The milk of lime slurry is used as a pH
modifier and will be distributed to the milling circuit, SART plant and cyanide detoxification area
via a pressurized ring main.
17.5.2
Sodium Cyanide
Sodium cyanide, supplied in solid (briquette) form, will be received in regular bulk shipments
from regional suppliers in 1 t bulk bags. The bags will be broken into a mixing tank and mixed
with tailings thickener overflow process water to produce a sodium cyanide solution of 20% w/w
concentration. The solution in the mixing tank is transferred to the stock solution storage tank,
which will provide 12 hours of cyanide supply at nominal production rates. From the stock solution
storage tank, sodium cyanide solution will be provided to the leaching tanks and Merrill Crowe
circuit as required via a double walled feed line.
The cyanide make-up and storage area will be fully contained in a bunded area and separated from
the plant site. The solid sodium cyanide will be stored in a fenced and locked area before being
used to prepare a stock solution.
17.5.3
Flocculant
Flocculant is received on site as a dry powder in 1 t bags. The powder is transferred into and stored
in a powder hopper. Flocculant is transferred from the hopper to a flocculant eductor by a screw
feeder, where fresh water is added. The flocculant and water are transferred into and mixed in the
flocculant mixing tank to a strength of 0.5% w/v. The mixed solution is transferred to the flocculant
storage tank and is pumped to the dosing points as required by individual dosing pumps.
Flocculant is fed into the pre-leach, CCD and tailings thickeners and to the SART process to assist
with solid settling. The flocculant is diluted to a concentration of 0.05% at each dosing point.
17.5.4
Coagulant
Coagulant is used to assist with solid settling in the clarifier. Coagulant is delivered to site as a
neat liquid in 1 t intermediate bulk containers (“IBC”) totes. The coagulant dosing pump will
deliver the neat solution to the clarifier as required.
17.5.5
Antiscalant
Antiscalant is used to prevent build-up of calcium and other scale deposits in the Merrill Crowe
and process water circuits. Antiscalant is delivered to site as a neat liquid in 1 t IBC totes. The
antiscalant dosing pumps will deliver the antiscalant solution to the pregnant solution and process
water tanks as required.
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17.5.6
Fluxes – Silica, Borax, Sodium Carbonate, Nitre
Various fluxes will be added to the gold sludge to smelt doré in the induction furnace in the gold
room. Borax, silica, sodium or potassium nitrite (“nitre”) and sodium carbonate will be delivered
as a granulated or powdered solid in small 25 kg or larger 1 t bags and stored and used in the gold
room.
Fluxes will be added to the gold sludge in the flux mixer prior refining in the induction furnace.
17.5.7
Sodium Metabisulphite
Sodium Metabisulphite (“SMBS”), supplied in solid powder form, will be received in regular bulk
shipments from regional suppliers in 1 t bulk bags. The bags will be broken into a mixing tank and
mixed with fresh water to produce a SMBS solution of 20% w/w concentration. The solution in
the mixing tank is transferred to the SMBS storage tank. The SMBS solution is then pumped to
the cyanide detoxification tanks, where it is used as source of sulphur dioxide (SO2), by a dedicated
dosing pump.
17.5.8
Copper Sulphate (Pentahydrate)
Copper sulphate (Pentahydrate), supplied in solid blue crystal form, will be received in regular
bulk shipments from regional suppliers in 25 kg bulk bags. The bags will be broken into the mixing
portion of the copper sulphate preparation tank and mixed with fresh water to produce a copper
sulphate solution of 20% w/w concentration. Once the batch is prepared, it will be transferred to
the solution storage portion of the copper sulphate preparation tank and dosed to the cyanide
detoxification tanks, where it is used as a catalyst.
17.5.9
Lead Nitrate
Lead nitrate, supplied in solid powder form, will be received in regular bulk shipments from
regional suppliers in 25 kg bulk bags. The bags will be broken into a heated mixing tank and mixed
with barren solution to produce a lead nitrate solution of 10% w/w concentration. Lead nitrate is
used as a precipitation reagent in the Merrill Crowe process and will be transferred to precipitate
filter feed line by a dedicated dosing pump.
17.5.10
Zinc Powder
Zinc, supplied in solid powder form, will be received in regular bulk shipments from regional
suppliers in drums. The drums will be unloaded into a zinc powder hopper and metered into a zinc
mixing cone by a powder feeder. Process water is added to the zinc mixing cone from the zinc
mixing cone head tank to form a zinc powder slurry which is dosed into the precipitate filter feed
line by the zinc cone discharge pump. Zinc is used in the Merrill Crowe process to facilitate the
gold and silver cementation process.
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17.5.11
Diatomaceous Earth
Diatomaceous Earth, supplied in solid powder form, will be received in regular bulk shipments
from regional suppliers in bulk bags. Diatomaceous earth is used in both the Merrill Crowe clarifier
leaf and precipitate filters in both a pre-coat and body coat application. Pre-coat is used at the start
of the filter cycle to prevent cloth blinding and body coat is added during filtration to provide
acceptable filtration rates. There is make-up tank for each application and bags will be broken into
each tank as required with process water added to make-up the solution. Dedicated pre-coat and
body coat dosing pumped transfer the solutions to the filters.
17.5.12
Sodium Hydroxide
Sodium hydroxide will be supplied in liquid form, delivered to site in 22 m3 bulk deliveries at a
solution concentration of 50.0% w/w. Sodium hydroxide will be transferred into a bulk storage
tank. The reagent is used as a pH modifier to neutralize acidic solutions in the SART and acid
digestion circuits. Sodium hydroxide will be dosed to the acid digestion plant at a 50%
concentration by a dedicated dosing pump, but will be diluted down and dosed at a 5%
concentration within the SART plant.
17.5.13
Sodium Hydrosulphide
Sodium hydrosulphide will be supplied in liquid form, delivered to site in 22 m3 bulk deliveries at
a solution concentration of 42.0% w/w. Sodium hydrosulphide will be transferred into a bulk
storage tank that will provide 5 days of storage. The reagent plays a key role in the SART
sulphidization process and is pumped to the SART plant by a dedicated dosing pump.
17.5.14
Sulphuric Acid
Sulphuric acid will be supplied in liquid form, delivered to site in 22 m3 bulk deliveries at a solution
concentration of 98% w/w. Sulphuric acid will be transferred into a bulk storage tank that will
provide two and a half days of storage. Sulphuric acid plays a key role as a pH modifier in the
SART process and acid digestion plant within the gold room. Sulphuric acid is pumped to both the
SART and acid digestion dosing points by dedicated dosing pumps.
17.6
17.6.1
ENERGY, WATER, AND PROCESS MATERIALS REQUIREMENTS
Electrical Power
The total installed power, average operating demand and estimated annual power consumption for
each of the process areas, including the SART plant, is provided in Table 17.4.
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TABLE 17.4
POWER REQUIREMENTS AND ENERGY CONSUMPTION
Plant Area
Installed
Power
(kW)
Crushing
Stockpile and Reclaim
Grinding
Leaching and Solution Management
Solution Recovery
Tailings Filters
Merrill Crowe and Refinery
SART
Process Plant Services and Common
Total
412
24
4,349
518
283
777
983
284
3,103
10,733
Normal
Operating
Power
(kW)
167
15
2,783
418
192
157
698
174
1456
6,060
Energy
Consumption
(MWh/a)
1,024
121
22,429
3,369
1,547
1,169
5,622
1,402
11,734
48,417
Source: Ausenco, (2025)
17.6.2
Fresh Water
Fresh water for the process facilities will be supplied from the La Labor (Martin) Dam, Los Huajes
Dam and the site contact water pond by dedicated intake pumps. Water is stored in and distributed
to the processing facilities from the fresh/fire water tank. Fresh water will be used for the following
duties:
•
•
•
•
•
Safety showers;
Paste plant make-up water;
Reagent mixing and preparation;
Dust suppression in crushing and emergency stockpile areas; and
Gland water.
Two fresh water pumps, operating in a duty/standby arrangement, are used to distribute fresh water
to the reagent mixing and preparation areas on the plant.
The anticipated processing facility fresh water consumption is included in Table 17.5.
TABLE 17.5
PROCESSING FRESH WATER ESTIMATED CONSUMPTION
Item
Water
Consumption
(m3/t plant feed)
Water
Consumption
(m3/a)
Fresh Water Requirement
0.30
220,800
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17.6.2.1
Fire Water
Fire water is supplied from a dedicated 250 m3 fire water reserve volume within the fresh/fire
water tank. A fire water pump skid provides fire water to the processing facility hydrants and hose
reels through a pressurized ring main. The distribution system pressure is continuously maintained
by a jockey pump.
The electric fire water pump will operate in a fire event and the fire water diesel pump provides
redundancy to the electric pump in the event of electrical power loss.
The firewater pump skid and control system are a vendor supplied package.
17.6.2.2
Gland Water
Dedicated duty/standby gland water pumps supply the fresh required for the slurry pump gland
water seal requirements. A duty/standby set of strainers on the discharge of the pumps remove
suspended solids to protect the integrity of the pump seals.
A gland water booster pump supplies high pressure gland seal water required for the filter feed
pump from the plant ring main.
17.6.2.3
Safety Showers
Fresh water is first passed through a sterilization water plant to ensure the water is of acceptable
quality prior to reporting to the safety shower tank. The safety shower tank provides two hours of
storage. Safety shower water is pumped by two duty/standby potable water pumps to the plant
safety showers, which are fed from a pressurized ring main system. Safety showers and eyewash
stations are located throughout the process plant at nominated hazardous locations.
17.6.2.4
Process Water
There are two dedicated process water pumping systems for the processing plant facilities:
•
•
Process water (pre-leach thickener overflow); and
Tailings thickener overflow process water.
Process water will consist of reclaimed water from the pre-leach thickener overflow, barren
solution and excess tailings thickener overflow. Process water is recycled to the CCD washing
circuit as wash water, used in the SART plant and used for flocculant dilution. Process water is
supplied from two duty/standby process water pumps. The process water supplied from the process
water tank will have high levels of residual cyanide.
Tailings thickener overflow process water will consist of reclaimed water from the tailings
thickener overflow and will have low levels of residual cyanide (<5 mg/L CNWAD) as it
downstream of the cyanide detoxification circuit. This process water is used for lime and cyanide
reagent make-up water, tailings filter cloth and manifold flush water and tailing thickener
flocculant dilution water. Excess process water will be returned to the process water tank. The
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tailings thickener overflow process water is supplied from two duty/standby tailings thickener
overflow pumps.
Any excess process water will be transferred from the tailing’s thickener overflow standpipe to the
collection pond. If impurities build-up in the process plant, a flow of process water to a collection
pond will allow for the required bleeding from the system. No water is expected to be discharged
to the environment.
17.6.2.5
Barren Solution
Barren solution will be high in residual free cyanide, as free cyanide is regenerated in the SART
process, and will be used primarily for make-up within the grinding circuit. The residual barren
solution will be used wherever possible, to minimize the fresh water consumption. In addition to
the grinding circuit, barren solution will be used for pre-leach thickener flocculant dilution, lead
nitrate solution preparation and within the Merrill Crowe circuit. Barren solution within the Merrill
Crowe circuit will be used for sluicing water for clarifier filters and pre-coat and body feed
preparation. Any excess barren solution will report to the process water circuit.
Barren solution will be stored in the SART barren solution tank and is pumped through a
pressurized header by the duty/standby SART barren solution pumps.
17.6.3
17.6.3.1
Air Services
Plant and Instrument Air
Plant and instrument air for the balance of the process plant facilities will be supplied at 700 kPa(g)
by two process plant air compressors, operating in a lead-lag configuration. The entire air stream
produced will be dried and filtered and will be used to satisfy both plant and instrument air demand.
Dried air will be distributed via the process plant air receiver, with an additional receiver for the
SAG Mill provided in the grinding area.
17.6.3.2
High Pressure Air for Tailings Area
High pressure air at 1,000 kPa(g) will be provided to the tailing’s filters by two high pressure air
compressors, operating in a lead-lag configuration. Air will be distributed locally to the tailings
filter for cake blowing and manifold blowing through two duty tailings filter air receivers.
17.6.3.3
Precipitate Filter Blow and Pressing Air
Precipitate filter blowing air will be provided to the precipitate filter by two filter blowing air
compressors, operating in a lead-lag configuration. Air will be distributed locally to the precipitate
filters from the precipitate filter blowing air receiver.
Precipitate filter pressing air will be provided to the precipitate filter by two filter pressing air
compressors, operating in a lead-lag configuration. Air will be distributed locally to the precipitate
filters from the precipitate filter pressing air receiver.
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17.6.3.4
Low Pressure Air for Leach and Detox Circuit
Low pressure air to the leaching and cyanide detoxification tanks will be supplied by two dedicated
blowers.
17.7
PROCESS MATERIALS
Major process plant wear consumables consist of crushing liner, SAG mill liner, SAG mill media
and pressure filter cloth replacements. and crushing screen deck panel replacements. The
consumption rates for crusher liners, SAG mill liners and SAG mill grinding media were based on
ore hardness properties, information provided by equipment suppliers and from Ausenco
experience with similar operations The consumption rates for the pressure filter cloths were based
on information obtained from equipment suppliers.
The annual consumption rates of the major process wear consumables are included in Table 17.6.
TABLE 17.6
LOM ANNUAL AVERAGE WEAR CONSUMABLE RATES
Wear Consumable
Primary Crusher - Lower Cheek Plate
Primary Crusher - Upper Cheek Plate
Primary Crusher - Fixed Jaw
Primary Crusher - Movable Jaw
SAG mill media
SAG Liners
Pressure Filter cloth
Unit
Item/a
Item/a
Item/a
Item/a
tonnes/a
Item/a
Item/a
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LOM Annual
Consumption
28
23
9
17
1,009
2
757
Page 399 of 577
18.0
PROJECT INFRASTRUCTURE
18.1
INTRODUCTION
The Los Ricos South Property is located in the state of Jalisco, Mexico, 75 km northwest of
Guadalajara. Access to the site is available via a 5 km paved road leading from the Magdalena
community, followed by an additional 20 km of well-maintained gravel road. The site will include
the following facilities:
•
Mining facilities include administration offices, truck shops, explosives storage, fuel
storage and distribution;
•
Process facilities include the process plant, crushing facilities, process plant workshop,
assay laboratory, fresh water infrastructure, SART plant, and paste plant;
•
Dry stack tailings facility; and
•
General facilities consist of the gatehouse, administration building, communications
centre, switchyard, and weigh station scale.
The location of site facilities was based on the following criteria:
•
Locate the facilities within the claim boundaries and permitting constraints;
•
Locate the primary crushing and run-of-mine pad between the mine and the process
plant to minimize hauling;
•
•
Locate the processing plant to take advantage of the natural topography and avoid water
courses;
Utilize the existing site access road to the greatest extent possible; and
•
Proximity to the power line.
The overall site layout is illustrated in Figure 18.1.
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FIGURE 18.1
OVERALL SITE LAYOUT
Source: Ausenco (2025)
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18.2
ROADS AND LOGISTICS
18.2.1
Site Preparation
Scrub brush clearing and topsoil removal are expected requirements to allow the construction of
the process plant and other buildings and facilities. Site civil work includes design for the following
infrastructure:
•
•
•
•
•
•
18.2.2
Light vehicles and heavy equipment roads;
Access roads;
Topsoil and overburden stockpiles area;
Mine facility platforms and processing facility platforms;
Water management ponds, ditches, and channels; and
Dry stack tailings storage facility.
Access to Site
The site is located near the City of Guadalajara, the capital of Jalisco State. From the capital, the
site can be accessed by two main roads: a four-lane tolled Highway 150 and a two-lane free-access
State Highway 15. Both roads converge in the town of Magdalena. From Magdalena, Highway 15
continues to State Road 15, which leads to the Project, while Highway 150 veers away from the
Project and leads to Hostotipaquillo (Figure 18.2).
The access road from Magdalena will be utilized for all equipment and materials during
construction and operations. The road is paved from Magdalena to the town of San Simon, then is
a dirt gravel road for a distance of 13.7 km to the Project site. The gravel road is planned to be
upgraded to improve the existing surface, clean and shape water ditches, and place a 15 cm layer
of compacted gravel on the road. Construction of 25 cm thick concrete slabs have been planned at
locations that have been identified as potential flood areas.
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FIGURE 18.2
ACCESS ROADS TO THE PROJECT SITE
Source: GoGold (2025)
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18.2.3
Plant Site Roads
The roads within the process plant area will be 7 m wide, integrated with the earthworks of the
process plant pad and designed with adequate drainage. The roads will accommodate two lanes,
each 3.0 m wide, with 0.5 m wide shoulders. These roads will provide access between the
administration building, warehouses, mill building, crushing buildings, and the mining truck shop.
The typical methods of clearing, topsoil removal, and excavation will be employed, incorporating
drains, safety bunds, and backfilling with granular material and aggregates for the road structure.
The entrance to the process and mine site will be via the gatehouse.
18.2.4
Airports
The nearest international airport is the Guadalajara International Airport (GDL) in the city of
Guadalajara, 118 km southeast of the Project site, with daily flights to the US, Mexico City, and
other Mexican destinations.
18.2.5
Security
The site will be accessible year-round via State Highway 15 through State Road 15 to
Hostotipaquillo. The main access road to the site begins at the intersection of J. Jesus Caldera
Rubio Avenue and Jose Maria Morelos Street. Access to the process plant and truck shop areas is
controlled by a security gatehouse. There is a second security gatehouse on the Southeast permitter
of the site to allow for controlled access to the explosive magazines and the toe of the tailings
facility from the town of Cinco Minas.
18.2.6
Shipping Logistics
18.2.6.1
Gold-Silver Doré
Gold-silver doré will be stored in the dedicated safe in the refinery. Doré product transportation to
third party refiners will be undertaken on a frequent basis by contractors using armoured trucks.
18.2.6.2
Copper Sulphide
The SART process will recover a saleable Cu2S precipitate. Transportation of the precipitate to
third party smelter/refiners will also be undertaken on a frequent basis by contractors.
18.3
18.3.1
POWER AND ELECTRICAL
Electrical System Demand
The maximum demand for the Los Ricos South Project is estimated at 14.6 MW. The power
demand is summarized in Table 18.1.
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TABLE 18.1
PROJECT SITE ELECTRICAL DEMAND ELECTRICAL DEMAND
Area
Underground Mining
Mine Infrastructure
Crushing
Grinding
Leaching
CCD, Detox, and Tailings
Filters
Merril Crowe and Refinery
SART
Paste Plant
Process Plant Services
Fuel Storage
Infrastructure Buildings
Total
Installed Power
(kW)
5,850
180
412
4,349
518
Maximum Demand
(kW)
4,973
126
335
4,029
425
Average Demand
(kW)
4,973
126
306
2,964
145
777
614
501
983
284
1096
3,103
23
876
18,450
815
232
895
1,521
19
613
14,595
672
174
758
1,264
17
613
12,511
Emergency power for process plant critical loads will be provided by three on-site diesel-powered
generators capable of producing 1,250 kW each. The emergency generators will be connected to
the 34.5 kV primary switchgear located at the power generation area.
18.3.2
Facility Power Supply
The Project will be fed by two circuits of medium voltage overhead power lines at a voltage of
34.5 kV, three phase and 60 Hz. Overhead distribution lines will be constructed using aluminum
conductor steel-reinforced cable (“ACSR”) and supported by concrete poles. The routing of the
lines will be from the CFE (national utility company) La Yesca Substation through Hostotipaquillo
Municipally and then to the Project site (Figure 18.3). The lines will be approximately 40.8 km in
length. Each circuit will have 7 MW of available capacity.
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FIGURE 18.3
OVERHEAD POWER LINE ROUTE TO PROJECT SITE
Source: GoGold (2025)
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18.3.3
Site Power Distribution
The overhead lines coming from the Las Yesca Servicios substation will arrive to a main
switchgear which will have a tie-in arrangement for manoeuvring in case of a utility failure.
Prefabricated sea container electrical rooms will be considered to accommodate the MV
switchgear to distribute power to all electrical substations located throughout the Project site.
The overhead powerlines will provide power from the 34.5 kV switch gear to the following
facilities:
•
Circuit 1
o Mine Truckstop
o Crushing
o Grinding
o Leaching
o CCD, detox, and tailings filters
o Merrill Crowe and refinery
o SART
o Process plant services
o Fuel storage
o Instructure buildings.
•
Circuit 2
o Paste plant
o Underground mining
o Process plant services.
A step-down transformer is used to reduce the voltage from 34.5 kV to 4.16 kV or 480 V to supply
the motor control centre (“MCC”) or variable frequency drive (“VFD”) as required by the
equipment.
Standby power will be provided by diesel generator set(s) installed in a strategic location of the
plant and will power essential loads identified in the electrical load list. The change-over from
utility power to the standby power and return to normal will be implemented with an automatic
transfer switch in a closed transition method which will be monitoring the system in case of a
failure event.
18.3.4
Plant Power Distribution
The largest electrical loads at the process plant are the SAG mill. The drive system for the SAG
mill includes motors, VFDs, and bypass switchgear to minimize voltage disturbances throughout
the power distribution system during start-up. The SAG mill drive systems will be supplied via
cable circuits from the plant’s primary 34.5 kV switchgear, circuit 1.
Power to the electrical rooms will be supplied by oil-filled distribution transformers located
adjacent to the respective electrical room. All electrical rooms will be adequately rates for the
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environment and outfitted with lighting and small power transformers, distribution boards,
uninterrupted power supply (“UPS”) systems, fire alarm and detection, and HVAC systems. To
reduce installation time, the electrical rooms will be prefabricated modular buildings installed on
structural framework above ground level for bottom entry of cables. Additionally, electrical rooms
will be located as close as practical to the electrical loads to optimize conductor sizes and minimize
cable lengths.
Pad-mounted and pole-mounted transformers will be used to step down the voltages at the truck
shop, mine dry, fuel station and tire storage, administration building, gatehouse and security,
laboratory, mill workshop, and warehouse areas. Power will terminate at the local 480 V or 220 V
distribution boards.
18.4
SUPPORT BUILDINGS
As shown in the site layout in Figure 18.1 above, the main plant site area consists of several
buildings. Most of the process plant is planned to be open air, as is the fuel station, power
generation area, water treatment area and paste plant. The On-Site buildings are listed in Table
18.2.
TABLE 18.2
DESCRIPTION OF ON-SITE BUILDINGS
Building Name
Merrill Crowe
Mine Truckshop
(Maintenance Shop)
Warehouse
Primary Crushing
Reagent Storage
Grinding Area
Hazmat Storage
Main Office
Process Plant Office
(x10)
Mess Building
#1-Gatehouse & Security
Building
#2-Gatehouse & Security
Building
#1-E-Room
#2-E-Room
#3-E-Room
#4-E-Room
#5-E-Room (MCC)
Pre-Engineered
L
(m)
28
W
(m)
25.5
H
(m)
6.5
Area
(m2)
714
Fabric
33
18
12
594
Fabric
Stick Built
Stick Built
Stick-Built
Stick Built
Modular
27
14
18
28
20
24.3
15
7.5
7.5
13.25
7.5
17.15
9.5
10
4.5
8.5
3.5
3
405
105
135
371
150
833
40-foot Trailers
12.2
2.45
2.6
299
Modular
30.35
14.1
3
428
Modular
6
4.9
3
29
Modular
6
2.45
3
15
Modular (Prefabricated Container)
Modular (Prefabricated Container)
Modular (Prefabricated Container)
Modular (Prefabricated Container)
Modular (Prefabricated Container)
14
11
21
11
23
4
4.5
4.5
4.5
9
4.5
4.5
4.5
4.5
5.4
56
50
95
50
207
Construction Method
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TABLE 18.2
DESCRIPTION OF ON-SITE BUILDINGS
Construction Method
L
(m)
W
(m)
H
(m)
Area
(m2)
Prefabricated Modular Sections
25
13.5
3.5
338
Building Name
Laboratory Sample
Preparation
18.5
ORE STOCKPILE
A run of mine (“ROM”) ore stockpile pad to hold approximately 100,000 t will be constructed
next to the primary crusher.
18.6
FUEL
Diesel fuel requirements for mining equipment, process and ancillary facilities will be supplied
from two modular above-ground horizontal diesel fuel storage tanks located adjacent to the
truckshop.
The fuel storage system shall be provided by a third party contractor. The fuel contractor will
supply a 60,000 L diesel storage tank and a 10,000 L gasoline storage tank. The fuel storage system
will include above-ground tanks include containment, and dispensing equipment to conform to all
applicable regulations.
18.7
FIRE PROTECTION
A complete fire detection and alarm system, fire water supply and distribution system, and waterbased fire protection system will be constructed and installed in accordance with applicable
regulations. Fire detectors, alarms, and extinguishers will be installed where required. The
firewater distribution network will be maintained under a constant pressure with a jockey pump
and will be sectionalized to minimize loss of fire protection during maintenance.
Yard hydrants will be limited to the fuel storage area. Wall hydrants will be used in lieu of yard
hydrants, and these will be located on the outside walls of buildings.
Fire protection within buildings will include sprinkler systems and portable fire extinguishers.
18.8
SEWAGE
Sewage collected from the ancillary buildings will be pumped to the sewage treatment module
prior for treatment to being discharged. The sewage treatment module will be a Moving-BedBiofilm-Reactor (“MBBR”) type. Treated effluent will be discharged to a holding tank before
being trucked off site.
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18.9
HAZARDOUS WASTE STORAGE FACILITY
An interim storage facility for hazardous waste will be built at site to allow for the storage of
ancillary wastes, such as used oils, before they are transported off-site for disposal by a third party.
This facility is fenced to limit entry and utilizing a corrugated plate roof to limit contact with the
natural environment.
18.10 MINING INFRASTRUCTURE
18.10.1
Haul Roads
The initial main road required for underground mining access is between the ROM pad and the
Eagle portal. It will be approximately 100 m long at a 10% gradient, and will be double lane.
General access tracks (for delivery of infrastructure such as fans, transformers, etc.) will be
required for:
•
•
•
•
•
•
18.10.2
Eagle main vent adit;
Eagle secondary vent adit;
New Abra vent adit;
Old Abra vent raise;
“The Windows” portal; and
“Farmyard” portal.
Explosives Facilities
An explosives magazine site will be constructed at the southeast corner of the permit area so that
it is at least 1 km away from the process plant.
18.10.3
Paste Plant
Plant dewatered tailings will be used as backfill for the underground mine. A paste backfill plant,
located close to the underground mine portal, will operate intermittently pumping thickened
tailings underground as needed for consolidated fill and stability.
Dewatered filter cake will be hauled from the plant filter cake bunker to the paste plant by truck
on an as needed basis. The paste plant will include a dewatered tailings filter cake reclaim and
conveying system, a cement binder silo and binder addition system, a paste mixer, a paste hopper
and a positive displacement paste pump.
The paste plant is sized to process 91 t/h and will operate between 1,448 hrs and 5,102 hours per
year, depending on the mine production schedule. The plant will require approximately 13.2 m3/h
of make-up water when operating and the annual water requirements will range between 19.1 m3/y
and 67.3 m3/y. The paste plant installed power is 1,096 kW with an average operating demand of
626 kW. The annual power consumption will range between 906 MWh/y and 3,194 MW/y.
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Process water make-up for the paste plant will be provided from contact water collected in the site
contact water pond.
Over the LOM, approximately 60 kg of cement binder will be added to each tonne of paste backfill
tailings.
18.10.4
Waste Rock Storage Facility
A storage facility for underground mine waste rock will be established near the DSTF. Waste rock
will also be used in the construction of the DSTF.
Initial waste rock from underground mining will be primarily used for construction purposes.
18.11
WATER SUPPLY AND MANAGEMENT
The Los Ricos South Project relies on two primary water sources, the Huajes and La Labor Dams,
to supply raw water to the site. Each dam has a permitted extraction rate and annual allocation,
with infrastructure in place to transport water through overland pipelines. A key component of the
water management plan involves handling both contact and non-contact water through diversion
channels, collection channels, and collection ponds, which help minimize runoff and allow for
water reuse in the process.
The Project site experiences a temperate climate with significant rainfall during the summer
months, and the water balance analysis indicates that the site will generally face a water deficit
throughout the year, requiring supplemental water from external sources. Water management
structures have been designed to manage surface runoff and ensure water availability for the plant
while maintaining a zero-discharge system, where surplus water will be stored.
Potable drinking water will be bottled and delivered to the Project site.
18.11.1
18.11.1.1
Water Supply
Off-Site Water Dams and Overland Water Pipelines
The Los Ricos site utilizes two primary water sources: Huajes Dam and La Labor, both of which
supply raw water to the site. The permitted extraction rate for each dam is 30 m³/h. Huajes Dam is
allocated a maximum annual volume of 250,000 m³, while La Labor is allocated 265,000 m³
annually.
To supply raw water to the site, each dam is equipped with two sets of floating pontoon pumps
operating in a duty/standby configuration.
The overland piping from Huajes Dam to the site extends 11.85 km and is constructed of DN50
HDPE DR17 (outer diameter approximately 60 mm, wall thickness approximately 3.5 mm) piping.
The overland piping from La Labor to the site spans 6.33 km, also using DN50 HDPE DR17
piping.
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18.11.1.2
Waste Water Pipeline
Wastewater generated on site is directed to the Sewage Treatment Plant from three primary
sources: the Office and Mess Buildings, the Warehouse, and the Maintenance Shop. The total
length of the wastewater piping system is 494 m, constructed from DN50 PVC piping. After
treatment, the Sewage Treatment Plant discharges the treated wastewater to a holding pond via a
10-m length of DN80 PVC piping.
18.11.1.3
Contact Water Pipeline
Contact water is conveyed to a UV Sterilization Skid through a DN100 PVC pipe, which is 250 m
in length.
18.11.1.4
Reclaim Water Pipeline
Reclaim water from the Dry Stack Tailings Facility (“DSTF”) is directed to the Plant Pond. From
there, the water is pumped into the Process Water Tank via two contact water pumps. The pipeline
from the DSTF to the Plant Pond is constructed of DN200 HDPE piping and spans 1.025 km. The
line from the Plant Pond to the Process Water Tank is also DN200 HDPE, with a length of 113 m.
18.11.1.5
Hydrometeorology
The Los Ricos South Project is located near the village of Cinco Minas, in the State of Jalisco,
Mexico. The site is situated at an elevation of approximately 1,184 metres above sea level (“masl”)
and experiences a temperate climate influenced by its moderate altitude. Rainfall mainly occurs as
heavy thunderstorms during the hot summer months (Ausenco, 2024).
The Hostotipaquillo climate station (ID: 14068), located in Jalisco, Mexico, was selected as the
primary data source to support the design of surface water management infrastructure and water
balance analysis. This station, situated at an elevation of 1,036.30 masl. has provided long-term
data collected over 70 years between 1949 and 2019. The data indicates that annual average
precipitation ranges between 800 mm to 1,200 mm. (Ausenco, 2024) A detailed summary of the
average monthly temperature, precipitation, and evaporation recorded at this station is presented
in Table 18.3 and Table 18.4.
TABLE 18.3
DETAILS OF CLIMATE STATIONS INCLUDING NAME, ID,
COORDINATES, AND DISTANCE TO SITE
Name
Hostotipaquillo
Distance
Record Latitude Longitude Altitude
from
ID
Location
Years
(°)
(°)
(masl)
the Site
(km)
Jalisco,
194914068
21.0581
104.051
1,300
15
Mexico
2019
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TABLE 18.4
SUMMARY OF MONTHLY TEMPERATURE, PRECIPITATION, AND
EVAPORATION DATA FOR HOSTOTIPAQULLO STATION
Parameter
Minimum Monthly
Temperature (°C)
Average Monthly
Temperature (°C)
Maximum Monthly
Temperature (°C)
Average Monthly
Precipitation (mm)
Average Monthly
Evaporation(mm)
Jan Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec
27
29
31
33
34
33
30
29
30
30
29
27
20
20
21
22
23
23
22
22
22
22
21
20
10
11
12
13
15
16
16
16
16
15
13
11
17
7
5
7
13
141 234
183
125
44
13
11
135 151
211
242
268
201 167
156
140 156 144 126
The Rainfall Depth-Duration-Frequency data from the Hostotipaquillo station provides rainfall
information for the design and assessment of surface water management infrastructure. Table 18.5
presents the corresponding rainfall values for different durations and return periods.
TABLE 18.5
INTENSITY-DURATION-FREQUENCY VALUES FOR LOS RICOS SOUTH SITE (MM)
Duration
(min)
5
10
20
40
60
100
120
240
300
600
900
1200
1500
2
12.1
18.1
25.2
33.7
39.4
47.4
50.5
54.5
56.5
62.7
66.3
68.9
71
5
15.9
23.8
33.3
44.5
52
62.6
66.7
70.5
73.1
81.1
85.8
89.2
91.9
Return Period (Years)
10
25
50
18.9
22.3
25
28.2
35.3
39.7
39.4
50.6
56.8
52.6
65.8
73.9
61.5
74.1
83.2
74
83.7
94
78.9
86.9
97.6
82.6
98.6
110.8
85.7
102.3
114.8
95
113.4
127.4
100.5
120
134.8
104.5
124.8
140.1
107.7
128.6
144.4
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100
27.8
44
63
82
92.3
104.3
108.3
122.9
127.4
141.3
149.5
155.5
160.2
200
30.5
48.4
69.2
90.1
101.4
114.5
119
135
140
155.2
164.3
170.8
175.9
Page 413 of 577
18.11.2
Water Management Structures
The key facilities for the water management plan are:
•
Diversion channels – Diversion channels are required to divert non-contact water of
existing watercourses currently flowing over the proposed mine footprint. Channels
will intercept natural drainage and divert it around the mine infrastructure to minimize
the amount of contact runoff to be collected and managed and to minimize disruption
to the downstream environment. The design criterion for the diversion channels is the
conveyance of 1/100-year; 24-hour peak flow without overflow.
•
Collection channels – Collection channels are designed to capture runoff from
disturbed areas, including the process plant area, rock storage facilities, and stockpiles.
The design criterion for these channels is to safely convey the 1/100-year, 24-hour peak
flow without overflow.
•
Collection ponds – Collection ponds are required to capture runoff from the collection
channels. The stored contact water will be re-used for process make-up water whenever
possible. The design criterion of the collection ponds is to store water from the 1/100year, 24-hour flood with a minimum freeboard of 0.5 m.
Thirteen diversion channels, with a combined length of 3.9 km are proposed. The channels are
planned with a 1 m base width, 2 H:1 V side slopes, and depth ranging from 0.14 to 4.82 m.
A series of collection channels are proposed to capture and manage contact water from the rock
storage facilities and process plant area. The collected contact water will be retained in five
collection channels, spanning a combined length of 0.8 km. The proposed configuration is
presented in Figure 18.4.
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FIGURE 18.4
WATER MANAGEMENT PROPOSED STRUCTURE FOR THE LOS RICOS SOUTH MINE SITE
Source: Ausenco (2025)
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18.11.3
Site-Wide Water Balance
A site-wide water balance analysis was conducted to evaluate the general water flow within the
site and potential water available for the process purposes. The water balance included runoff from
disturbed mine components as well as the surrounding catchments. Table 18.6 presents the overall
annual and peak monthly runoff for the Los Ricos South mine site.
Key factors considered in the water balance include surface runoff form contact areas.
Assessments for each facility was based on average climate conditions at the site using appropriate
runoff coefficients. The water components considered in the site-wide water balance calculations
include surface runoff from precipitation on process plant areas and stockpiles. Additionally, the
site-wide water balance accounts for reclaimed water from the dry stack tailings facility and
dewatering from the old underground mine throughout the mine’s life.
The results of the water balance indicate that the mine will run a water deficit for most of the year
and will require additional make-up water from outside sources. It is envisioned that the site is
zero discharge, and surplus water derived onsite during wetter months will be stored for process
use or will be lost to evaporation.
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TABLE 18.6
SITE WIDE WATER BALANCE FOR LOS RICOS SOUTH MINE SITE
Month
Item
Annual
Jan
Feb
Mar
Apr
May
Jun
Jul
Aug
Sep
Oct
Nov
Dec
19,440
20,088
20,088
19,440
20,088
19,440
20,088
236,520
4,337
7,183
5,617
3,849
1,351
387
341
24,556
470
778
608
417
146
42
37
2,660
3,907
7,814
2,208
2,038
1,189
679
3,227
3,781
7,562
2,137
1,973
1,151
658
3,123
3,907
7814
2,208
2,038
1,189
679
3,227
3,781
7,562
2,137
1,973
1,151
658
3,123
3,907
7,814
2,208
2,038
1,189
679
3,227
3,781
7,562
2,137
1,973
1,151
658
3,123
3,907
7,814
2,208
2,038
1,189
679
3,227
46,000
92,000
26,000
24,000
14,000
8,000
38,000
Los Ricos South Water Components (m3/month)
Process
Plant
Make-up
20,088
18,144
20,088
19,440
20,088
Water
Demand
Contact Water from Net Precipitation (m3/month)
Surace
Runoff
from
534
224
141
203
390
Collection
areas
(combined)
Direct Precipitation on the Collection Ponds (m3/month)
Combined
58
24
15
22
42
Ponds
Dewatering Rates of the Old Underground Mine (m3/month)
Year -2
3,907
3,529
3,907
3,781
3,907
Year -1
7,814
7,058
7,814
7,562
7,814
Year 1
2,208
1,995
2,208
2,137
2,208
Year 3
2,038
1,841
2,038
1,973
2,038
Year 5
1,189
1,074
1,189
1,151
1,189
Year 6
679
614
679
658
679
Year 7
3,227
2,915
3,227
3,123
3,227
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TABLE 18.6
SITE WIDE WATER BALANCE FOR LOS RICOS SOUTH MINE SITE
Month
Item
Year 9
Annual
Jan
Feb
Mar
Apr
May
Jun
Jul
Aug
Sep
Oct
Nov
Dec
5,605
5,063
5,605
5,425
5,605
5,605
5,425
5,605
5,425
5,605
5,245
5,065
66,000
12,882
24,813
21,793
20,246
23,322
21,933
9,994
19,396
17,032
15,813
18,256
17,168
6,860
13,339
11,717
10,869
12,560
11,811
2,692
4,927
4,351
4,025
4,584
4,315
1,098
1,683
1,511
1,388
1,504
1,420
1,030
1,537
1,384
1,269
1,364
1,289
47,881
88,072
77,663
72,012
82,036
77,203
389
363
326
363
334
293
4,878
-8,609
-4,703
2,574
14,335
10,466
8,409
14,033
15,022
-10,319
-6,412
-2,023
7,209
3,995
2,267
7,258
8,548
-11,719
-7,938
-6,503
-188
-2,632
-3,973
184
1,736
-15,048
-11,141
-14,054
-11,989
-13,414
-14,250
-11,143
-9,034
-15,565
-11,784
-16,111
-15,690
-16,684
-17,300
-14,718
-12,501
-16,097
-12,190
-16,765
-16,428
-17,430
-18,055
-15,412
-13,109
-168,182
-122,182
-140,301
-102,110
-122,519
-134,170
-94,146
-70,979
Inflows from Dry Stack (m3/month)
Year 1
1,529
876
762
853
1,198
8,107
Year 3
2,279
1,150
898
1,095
1,747
15,208
Year 5
2,045
1,047
829
1,000
1,573
13,381
Year 6
1,890
961
756
916
1,449
12,430
Year 7
2,024
985
743
931
1,542
14,221
Year 9
1,911
933
705
882
1,457
13,379
3
Pan Evaporation from Ponds (m /month)
Combined
314
350
490
564
624
467
Ponds
Deficits/Excess (-/+) Water Available after Process Demands (m3/month)
Year -2
-15,903
-14,717
-16,515
-15,998
-16,373
-11,319
Year -1
-11,996
-11,188
-12,608
-12,218
-12,466
-7,538
Year 1
-16,072
-15,375
-17,451
-16,789
-16,874
-4,856
Year 3
-15,492
-15,255
-17,485
-16,712
-16,495
2,080
Year 5
-16,576
-16,125
-18,404
-17,629
-17,518
-568
Year 6
-17,240
-16,671
-18,986
-18,206
-18,152
-2,013
Year 7
-14,558
-14,346
-16,451
-15,725
-15,511
2,244
Year 9
-12,293
-12,250
-14,111
-13,473
-13,218
3,704
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18.12 DRY STACK TAILINGS FACILITY
18.12.1
Site Description and Overview
The concept of using a DSTF was adopted based on the mine plan, limited space for tailings and
waste rock storage, a shortage of available construction materials, and the need to mitigate risks
associated with conventional tailings storage facilities that utilize tailings dams, especially given
the proximity of residential buildings in this seismic area. The tailings stored on the surface will
be filtered and transported by trucks to the DSTF. The design incorporates a fully lined
geosynthetic liner for the DSTF, which is a permitting requirement aimed at preventing any
potential cyanide leaching into the environment, even though the tailings undergo a cyanide
destruction process before being placed in this facility. The liner provides environmental
protection by creating an impermeable barrier that reduces the risk of contaminants migrating into
the surrounding soil and groundwater. The DSTF is designed to store approximately 5.3 million
tonnes of filtered tailings and is embanked by waste rock at the south end. Additionally, the facility
features two small berms to enhance stability. The proposed geometry and key features of the
DSTF are illustrated in Figure 18.5, and the DSTF has a projected maximum elevation of 1,279
masl.
Filtered tailings, because of their unsaturated state and primarily dilatant geotechnical behaviour,
eliminate the necessity for large retention embankments (dams) and significantly reduce physical
stability risks. Once placed and compacted, these tailings exhibit very low permeability. When
combined with effective surface water management, this characteristic greatly minimizes the
amount of water that can infiltrate the natural environment. Another significant advantage of
filtered tailings is the ability to recycle water recovered during the filtration process back into
metallurgical operations. Furthermore, compared to conventional slurry tailings storage facilities,
dry stack filtered tailings facilities typically require a smaller storage footprint, are more conducive
to progressive reclamation, and present lower long-term closure liabilities regarding
environmental concerns impacts. The optimal location selection for the DSTF was based on
several key criteria. These included compliance with the standards outlined in NOM-141SEMARNAT-2003 as mandated by Mexican authorities, ensuring the physical stability of the
DSTF, maximizing the use of available real estate by prioritizing land owned by the Company,
and minimizing the distance to the filter plant. Accordingly, geotechnical considerations, such as
the location and properties of the soil and rock material underlying the Project area and the
hydraulic factors associated with the catchment area, were evaluated.
The DSTF will encompass additional infrastructure, including a contact water collection system,
storage ponds and channels, underdrain collection systems, and access roads. These components
are integrated into the areas designated for reclamation under the closure plan. Noncontact water
diversion channels will be built to reduce the generation of contact water from exposed regions of
the DSTF. These channels are expected to stay in place, as they offer advantages by diverting
water during low-return-period precipitation events, provided they remain functional. Over time,
these structures are likely to be filled with sediment and gradually merge into the surrounding
landscape.
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FIGURE 18.5
DSTF LOCATION AND FOOTPRINT
Source: Ausenco (2025)
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18.12.2
Tailings Characterization
Two representative composite samples of tailings designated “ABRA” and “EAGLE” were
shipped to the BGC Engineering (“BGC”) geotechnical laboratory in Fredericton, Nova Scotia,
Canada, and the SGS laboratories in Lakefield, ON, Canada (and their subcontractors). These
composite tailings samples were subjected to geotechnical tests.
According to the Unified Soil Classification System (“USCS”), both samples were classified as
non-plastic silt with some fine-grained sand (ML). EAGLE and ABRA samples had approximately
78% and 79% by weight of their particles passing the No. 200 mesh (0.075 mm) and specific
gravity (SG) values of 2.71 and 2.68, respectively. Based on the standard Proctor compaction test,
EAGLE and ABRA samples have both a maximum dry density of 17.4 kN/m3 and optimum
geotechnical moisture contents (weight of water over the weight of solids) of 15.0% and 15.3%,
respectively.
Flexible wall falling head hydraulic conductivity tests showed that EAGLE and ABRA samples
have average hydraulic conductivity values of 3.7E-07 and 2.8E-07 m/s, respectively.
Rowe (hydraulic) consolidation tests were performed on both EAGLE and ABRA samples,
applying a maximum cell pressure of 800 kPa. The samples were tested at a maximum dry density
of 17.4 kN/m³ and optimum geotechnical moisture contents of 15.0% and 15.3%, respectively, as
determined by the standard Proctor compaction test. These tests provided data on the void ratio,
coefficient of consolidation, volume compressibility, and hydraulic conductivity of the samples.
The soil water characteristic curve (“SWCC”) was determined (ASTM D6836-16) for the ABRA
sample at 95% compaction.
Consolidated undrained triaxial tests were performed on remodeled ABRA samples under
confining stresses of 200, 400 and 800 kPa, which resulted in maximum internal friction angle φ
of 37.4º, 36.3º, 35.9º, respectively.
Cyclic triaxial tests were conducted on ABRA samples remolded to 95% of their maximum dry
density. The tests were performed with cyclic stress ratios (“CSR”) of 0.1, 0.15, and 0.2, under a
consolidation stress of 1,200 kPa. The results indicated that the maximum number of cycles to
failure was 1,500 for CSR = 0.1, 127 for CSR = 0.15, and 14 for CSR = 0.2.
18.12.3
DSTF Foundation
A firm foundation is crucial for ensuring the stability and proper function of the DSTF. Foundation
preparation begins with the thorough removal of all vegetation and roots in the native soil. This
process involves clearing away brush and debris, as well as grinding and removing stumps.
Organic-rich soils removed during this process should be carefully collected and stored in a
designated topsoil deposit for future use in revegetation and closure activities. Once clearing and
grubbing are complete, most of the soil cover within the DSTF area should be extracted and
repurposed as borrow material. The remaining soil cover must be prepared and compacted in areas
where soil removal is unnecessary—either for grading or because sufficient borrow material is
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already available. This preparation is vital to prevent excessive deformation and address any
potential weak zones in the foundation.
The foundation parameters for the DSTF were determined using historical laboratory data
collected from seven test pits and five boreholes located within the footprint of the starter DSTF.
However, the available laboratory data from these boreholes and test pits was limited, and no
geotechnical information was available for areas within the DSTF footprint but outside the starter
area. A geophysical survey was conducted in 2024 over the southern portion of the DSTF to
address these gaps. This survey included three geophysical lines, covering a total distance of
approximately 1,250 m, to provide additional insights into the foundation characteristics and
enhance the understanding of site conditions.
18.12.4
Non-Contact Surface Water Diversion Systems
The non-contact surface water diversion system will involve constructing ditches to redirect
surface water from uphill areas away from the DSTF footprint, ensuring it does not come into
contact with the tailings downstream. These diversion channels must be excavated before DSTF
construction begins to minimize contact water generation and reduce the size requirements for
contact water management structures. The channels are designed to safely handle runoff from a
1/100-year, 24-hour storm event. At the current design stage, the channels are specified with a
base width of 1 m, side slopes of 2 H:1 V (horizontal to vertical), and variable depths depending
on the flow rate.
18.12.5
Underdrain Non-Contact Water Collection System Installation
The underdrain non-contact water collection system is designed to collect and manage non-contact
groundwater beneath the liner system, preventing uplift or hydrostatic pressure from affecting the
stability of the DSTF structure. This system ensures that groundwater is effectively drained away
from beneath the facility, maintaining the integrity of the liner and reducing potential risks
associated with water pressure. The system incorporates 100 mm and 300 mm perforated and solid
dual-walled pipes installed within a gravel drainage layer to channel the collected groundwater
safely away from the DSTF footprint. This critical component of the facility’s design supports
overall structural stability while ensuring compliance with environmental and geotechnical
standards.
18.12.6
Composite Liner System
The composite liner system consists of compacted subgrade, GCL, and geomembrane placed over
these materials to ensure effective containment and environmental protection. The geomembrane
liner is made of a 2.0 mm LLDPE DST liner, providing a robust and durable barrier. In steep
terrain, a GCL is installed directly on the ground surface with the geomembrane placed on top,
creating an effective seal against seepage. In other areas, a 0.3 m thick low-permeability soil layer
serves as an alternative to the GCL. At the junction where the GCL meets the low-permeability
soil layer, the GCL extends 2 m into the low-permeability soil, ensuring seamless integration and
minimizing the risk of leaks. Detailed installation procedures and configurations are provided in
the design drawings of the DSTF.
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18.12.7
Overliner Collection System
The overliner collection system is designed to manage water from potential seepage through the
tailings where it intersects with the liner system. Numerical modelling, which simulates the natural
processes of precipitation, infiltration, and evaporation on the surface of the compacted filtered
tailings, indicates that infiltration into the tailings will be negligible. However, as an additional
measure for environmental and geotechnical risk mitigation, an overliner water collection system
has been proposed for installation above the geomembrane liner. This system aims to capture and
drain any potential seepage, directing contact water to a downstream pond for storage, monitoring,
and reuse in the process plant. The water collection system will include 100 mm and 300 mm
perforated solid dual-walled pipes and a gravel drain cover. Detailed information and drawings for
the over-liner water collection system are included in the design drawings of the DSTF.
18.12.8
Contact Water Collection Ponds
The DSTF design features a fully lined facility with a permitting requirement to prevent potential
cyanide leaching into the environment. Any contact water produced from tailings runoff will be
collected in the contact water collection pond located downstream of the DSTF embankment for
later reuse in the process plant. The contact water collection pond will include sedimentation zones
to remove solids before the water is stored, monitored, treated if necessary, or pumped back to the
plant for reuse. The contact water pond for the initial DSTF phase, which will be the first to become
operational, is designed to have a capacity of 400 m³, while the ultimate pond will have a capacity
of 21,350 m³.
Both the starter and ultimate contact water collection ponds are designed to manage a 100-year
flood event and the resulting contact water runoff from the active tailings placement areas,
ensuring effective water management and environmental compliance.
18.12.9
Non-Contact Water Collection Ponds
Non-contact water collection ponds will be built downstream of the DSTF, one designated for the
starter DSTF and the other for the ultimate DSTF. These ponds are designed to manage
groundwater collected by the underdrain system positioned beneath the DSTF liner. They function
as storage and control points for non-contact water, ensuring effective separation from contact
water. Groundwater captured by the underdrain system, which comprises 100 mm and 300 mm
perforated and solid dual-walled pipes, is directed to these ponds for secure containment and
management.
18.12.10 DSTF Design Key Elements
The DSTF was designed to accommodate an approximate capacity of 5.3 Mt of tailings. The design
incorporates bench face angles (“BFA”) of 26.6° (2.5 H:1 V), bench widths of 8 m, and a maximum
height of approximately 100 m, measured from the lowest point of the embankment to the highest
elevation of the DSTF (Figure 18.6).
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18.12.11 Geotechnical Analysis
As discussed in Section 18.12.2, geotechnical tests conducted on the two composite tailings
samples, ABRA and EAGLE, showed that the two samples had almost identical geotechnical
properties and characteristics. For geotechnical design purposes, an internal friction angle (φ) of
35° and cohesion (C) of 0 was adopted for use in limit equilibrium stability analyses.
18.12.12 Stability Analysis
Stability analysis was performed using the commercially available software SLOPE/W (GeoStudio 2021). This software facilitates limit equilibrium slope stability calculations employing
various methods and failure surface search routines. It allows for the analysis of individual slip
surfaces or the application of search algorithms to identify the critical failure surface—defined as
the surface with the lowest deterministic factor of safety—based on specified soil shear strength
properties, geometry, pore water pressure, and loading conditions.
The stability of the DSTF was evaluated under two scenarios:
•
•
Static loading conditions; and
Seismic loading conditions were analyzed using pseudo-static methods.
To determine the critical slip surface for both static and seismic scenarios, a representative (most
critical) cross-section was selected (Figure 18.6), reflecting the final configuration of the DSTF.
The proposed configuration of the DSTF meets the design criteria for stability under both static
and seismic loading conditions. This conclusion was reached after a thorough analysis of the site
conditions and the geotechnical properties of the materials that will comprise the structures.
FIGURE 18.6
MOST CRITICAL CROSS-SECTION OF THE DSTF FINAL CONFIGURATION
USED IN THE STABILITY AND SEEPAGE ANALYSES
Source: Ausenco (2025)
18.12.13 Seepage Analysis
A series of seepage modeling analyses were conducted to estimate seepage during the DSTF’s
construction and operational phases. These analyses were performed using the commercially
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available 2-D software SEEP/W (Geo-Studio 2021), which supports both steady-state and transient
flow analyses in porous media. A representative cross-section of the DSTF (Figure 18.6) was
selected for unsaturated flow modeling.
The results indicated that the native ground remained unsaturated after a 15-year simulation period
(LOM). By this time, the cover system would have been completed, and the DSTF would have
been fully vegetated and reclaimed. Given that the DSTF is lined with a 2.0 mm LLDPE doublesided texture (“DST”) geomembrane, the modeling demonstrated that, under the expected
conditions of filtered tailings placement and compaction, combined with site-specific climatic
conditions, no seepage into the native ground is anticipated.
Despite these favorable modeling results and the presence of the geomembrane liner, the DSTF
design includes a robust underdrain water collection system and ponds. These systems are
designed to capture any potential seepage that may occur during construction phases and
effectively manage contact water upstream of the starting embankment structure. This approach
ensures comprehensive water management and enhances environmental protection.
18.12.14 Starter Embankment and Stability Berms
The starter embankment is planned for construction at the toe of the starter DSTF to enhance
stability and provide erosion protection (Figure 18.7). The embankment is designed with slopes of
2 H:1 V on both the downstream and upstream sides. It will have a crest width of 15 m and a
maximum height of approximately 12 m.
An upper and lower stability berm are planned for construction within the starter DSTF area. These
upper and lower stability berms are critical structural elements designed to enhance the stability of
the facility. Constructed from non-acid generating (“NAG”) waste rock, these berms provide
foundational support to the starter facility, ensuring long-term geotechnical stability. Their
placement and design play a key role in maintaining the integrity of the DSTF, particularly during
the initial phases of operation. The stability berms, along with the starter embankment, will be
covered by a liner, GCL, and a low-permeability soil layer to ensure proper containment and
environmental protection.
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FIGURE 18.7
STATER DSTF CONFIGURATION
Source: Ausenco (2025)
18.12.15 Ultimate Embankment
The ultimate embankment is planned for construction at the toe of the ultimate DSTF to enhance
stability and provide effective erosion protection. It is designed with 2 H:1 V slopes on both the
upstream and downstream sides (Figure 18.7) and will be built using NAG waste rock or other
approved borrow materials. All materials used for the embankment will be verified to be free of
potentially acid-generating (“PAG”) and metal leaching (“ML”) risks, ensuring environmental
safety and long-term stability.
The embankment's crest will be 15 m wide and approximately 30 m high. To enhance containment,
the upstream slope will be covered with a 0.3 m thick clay layer topped with a geosynthetic liner.
Additionally, the crest will be reinforced with a 1 m thick layer of structural fill to cover the NAG
waste rock, ensuring durability and maintaining structural integrity.
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18.12.16 Filtered Tailings
After grading and foundation preparation are completed, along with the construction of the noncontact water diversion channels, underdrain collection system, liner system, overliner collection
system, contact water collection pond, non-contact water collection pond, and starter embankment,
the facility will be ready to receive filtered tailings. The starter DSTF will be commissioned first,
with the capacity to receive tailings up to three bench heights (60 m).
The design features a fully lined geosynthetic liner for the DSTF, which is a permitting requirement
aimed at preventing potential cyanide leaching. Results from triaxial tests on the filtered tailings,
along with seepage modeling analyses, indicate that surface infiltration will be minimal, with most
precipitation expected to run off the surface of the Deposit. As a result, contact water is expected
at the interface between the filtered tailings and the liner. This water will be efficiently managed
by the overliner collection system and directed to the contact water collection pond located
downstream of the DSTF.
The tailings will dewater in the filter plant to a gravimetric moisture content (ww/ws) of
approximately 14 to 18%. Based on standard Proctor compaction tests, the optimum moisture
content for the filtered tailings ranges from 15.0 to 15.3%. The tailings can reach 95% compaction
of the standard Proctor test with approximately ±2% of the optimum moisture content.
Given the site's tropical savanna climate, evaporation may occur during the loading, transport, and
placement of tailings from the filter plant to the DSTF. As a result, the tailings can typically be
placed and compacted at their initial moisture content directly after leaving the filter plant, even
during light rainy conditions. However, on days with heavier precipitation, when the tailings'
moisture content exceeds the optimum level, unprocessed (ungraded and uncompacted) tailings
can be protected using HDPE sheeting or rain covers until weather conditions permit for sufficient
moisture loss through evaporation. The geotechnical design specifies that the filtered tailings must
be compacted to 95% of the maximum Proctor density within a tolerance of ±2% of the optimum
moisture content and in lifts not exceeding 0.3 m.
18.12.17 Coarse Graded Filtered Tailings Cover
Once the filtered tailings are placed and compacted, and at least one bench has been constructed,
surface slopes and areas that have reached the proposed final grading will be covered with a coarsegraded, NAG waste rock layer approximately 0.5 m thick.
This cover serves multiple purposes: it protects against erosion, mitigates the resuspension of
tailings dust by wind, reduces the suspension of solids due to surface water runoff, and aids in
surface revegetation. This strategy is a crucial part of the reclamation process and promotes the
progressive closure of DSTF areas that have reached their target storage capacity (Figure 18.8).
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FIGURE 18.8
DSTF CLOSURE CONCEPT
Source: Ausenco (2025)
18.12.18 DSTF Construction
Some key elements of the DSTF can be constructed in parallel or the order can change slightly;
however, generally, the proposed construction sequence of the key elements of the DSTF design
is as follows:
18.12.18.1
Starter Phase
The following activities are part of the starter phase for preparing the DSTF:
•
Clearing, grubbing, and grading the area designated for the DSTF;
•
Removing topsoil from the area to be occupied by the DSTF;
•
Excavating unsuitable materials from the embankment, ponds, and upper and lower
stability berms;
•
Preparing and compacting the subgrade within the DSTF basin, embankment, and
ponds;
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•
Excavating trenches for the underdrain collection system;
•
Backfilling the upper and lower stability berms;
•
Preparing the foundation for the DSTF area, including the non-contact underdrain
collection system and structural components such as the starter embankment and water
collection ponds;
•
Constructing water diversion systems to manage non-contact surface water upstream
of the DSTF;
•
Installing a non-contact water underdrain collection system to channel groundwater
beneath the DSTF footprint;
•
Installing the liner system, GCL, and over-liner collection system to collect contact
water resulting from seepage within the DSTF;
•
Constructing contact water ponds for sedimentation, temporary storage, and
monitoring of contact water before its reuse in the process plant;
•
Building the upper and lower stability berms and the starter embankment at the toe of
the DSTF;
•
Placing, grading, and compacting filtered tailings; and
•
Progressively placing a coarse material cover over the filtered tailings as they reach the
proposed final grade at each bench. This prevents water erosion, reduces dust
resuspension, and supports progressive closure implementation.
18.12.18.2
Ultimate Phase
The activities outlined for the starter phase will be repeated during the ultimate phase to ensure the
complete construction and preparation of the DSTF.
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19.0
MARKET STUDIES AND CONTRACTS
Metal prices are presented in Table 19.1 and are based on an average of approximate December
31, 2024, three-year monthly trailing averages, and Consensus Economics Inc. long-term
projections from various financial institutions. The Mexican Peso:US Dollar exchange rate is the
approximate average of the past four years. The metal prices and exchange rate are subject to spot
market conditions. There are no metals streaming or hedging agreements in place.
TABLE 19.1
METAL PRICES AND EXCHANGE RATE
Item
Gold (US$/oz)
Silver (US$/oz)
Copper (US$/lb)
Exchange Rate (MXN:US$)
19.1
Price
2,330
26.80
4.00
20
SALEABLE PRODUCTS
The Los Ricos South Project will produce two saleable products:
•
•
Silver and gold doré bars; and
Copper precipitate (Cu2S).
The doré bars will be transported to a gold refinery. The assumed terms for gold are 99% payable
and $5/oz refining charge. The assumed terms for silver are 99.9% payable and $0.50/oz refining
charge.
Indicative quotations were received from several smelting companies and offtakers to treat the
copper precipitate. Indications were that a net 70% Cu would be payable (includes all-in charges
such as insurance, transportation, smelting and refining).
19.2
CONTRACTS
Currently there is one contract in place that is material to the Project. GoGold has entered an
agreement with CFE, the national power company. A dedicated two circuit 41 km overhead power
line will be installed from the La Yesca Dam where CFE has a power generation plant and
electrical substation. The routing of the line will be from the La Yesca Substation through
Hostotipaquillo Municipally and then to the Project site. Two overhead power lines will connect
34.5 kV, three phase and 60 Hz via a sub-station located near the process plant. The installation
contract amounts to $8.0M of which $0.74M has been paid by GoGold, with $7.26M remaining.
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20.0
ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY
IMPACT
20.1
BACKGROUND AND PROJECT LOCATION
Beginning in 2019, Consultores Interdisciplinario en Medio Ambiente S.C. (“CIMA”) conducted
environmental and social assessments in the region of the Los Ricos South Project. A primary
consideration was of the environmental baseline within the area of influence of the Project. In
order to know the local natural aspects, such as climatology, edaphology, hydrology, wild flora
and wildlife, as well as the social and economic aspects of the local villages and municipalities, an
area-relevant environmental system (“ES”) was established.
This section of the Technical Report is a condensation of a much more detailed report prepared by
CIMA titled “CIMA, 2024, Technical Report 43-101 Los Ricos South 2024 - Environmental
Section”.
Land use in the surrounding area is mostly agricultural with livestock grazing, although there exists
some small-scale artisanal mining activities (gambusinos) and forestry practices. The Project
location (yellow outline) and a suggested Environmental System (red outline) suitable for baseline
and environmental impact studies are shown in Figure 20.1. Existing electrical powerlines are
shown in blue (400 kV) and green (115 kV).
FIGURE 20.1
ENVIRONMENTAL SYSTEM OF THE LOS RICOS SOUTH PROJECT
Source: CIMA (2024)
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20.2
PROJECT LOCAL FEATURES
The Project is located near the towns of Cinco Minas and Huajacatlán, in the municipality of
Hostotipaquillo in the State of Jalisco.
For development of the Project, selected areas suggested to the regulatory authority Secretaria de
Medio Ambiente y Recursos Naturales, Secretary of the Environment and Natural Resources
(“SEMARNAT”), represent a total area of 34.2 ha, of which 19.0 ha are currently in forested zones,
while the remaining 15.2 ha are listed as agricultural and road allowances. Among the proposed
area uses, a dry-stack tailings facility (“DSTF”) and waste rock storage facility (“WRSF”) are the
largest features. There is also a collection of processing and support infrastructure, underground
mine workings and open pits, all of which require significant surface area. The list in Table 20.1
identifies the estimated surface areas of the Project components.
TABLE 20.1
SURFACE AREA OF PROJECT COMPONENTS
No.
Uses
1
Ore Storage
2
Other Storage
3
Entrance Area and Guardhouse
4
Process Plant and Industrial Areas
5
Connection and Perimeter Roads
6
Dry Stack Tailings Storage
7
Administrative Building
8
Fuel Station
9
Maneuver Yards / Mine Services
10
Water and Ventilation Plants
11
Powder Magazine / Topsoil
12
Open Pits
13
Electrical Substation and Diesel Generators
14
Maintenance Workshop
15
Waste Rock Storage
16
Plant Nursery and Topsoil Storage
Total Project Area
20.3
20.3.1
Hectares
1.2
0.6
0.02
5.9
3.9
10.0
0.7
0.2
1.4
0.05
0.3
3.2
0.2
1.2
4.7
0.7
34.2
MAJOR PROJECT COMPONENTS
Waste Rock Storage
The WRSF is the area where broken rock with no economic value will be located. This includes
rock which must be removed from underground development drifts, and waste rock, overburden
and organic soils in the open pits to expose the ore.
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The storage of the waste material is part of the operational stage, however, part of this waste can
be used in road improvement works, thereby reducing the total volume of material to be stored in
the waste rock facility. Within the WRSF there will be haul roads that are continuously modified.
20.3.2
Dry Stack Tailings Facility (DSTF)
The tailings are stacked in a so-called "dry stack" pile, which differs from conventional tailings
management methods. This storage method surpasses the common features of conventional
methods. The method minimizes surface area, allows rehabilitation activities to begin before mine
closure, favors more efficient water recycling, can be built adjacent to the process plant site. The
method significantly reduces the need for containment contingencies, since the risk of spills is
reduced.
20.3.3
Process Plant
The process plant will utilize conventional technology: crushing, grinding and cyanide leaching of
gold and silver. Los Ricos South process variations include the production of a flotation
concentrate that will be subject to intense cyanide leaching followed by conventional leaching of
the flotation tailings.
“Pregnant” leaching solution will be clarified, and treated in a Merrill Crowe process, which
utilizes the addition of zinc powder to precipitate the gold and silver. The precious metal precipitate
will be filtered and treated in a retort furnace to remove naturally occurring mercury by
volatilization and then smelted on site in an electric furnace to produce gold and silver doré bars.
The captured mercury will be collected in a gas scrubber and safely disposed off-site.
In addition to the Merrill Crowe process, a Sulphidization, Acidification, Recycling and
Thickening (“SART”) process will be carried out to recover a saleable copper sulphide (Cu2S)
precipitate and regenerate the cyanide associated with the soluble copper. The high cyanide
solution from SART will be recycled to leach feed.
The process plant design and operation will comply with an extended list of applicable codes and
standards.
20.4
PROJECT DEVELOPMENT
20.4.1
Site Preparation and Construction
After environmental authorizations are obtained, prevention and mitigation actions associated with
the implementation of the Change of Use of Soils in Forest Land (“CUSTF”) will be carried out,
in particular, the delimitation of the authorized forest polygons, the implementation of the Wild
Flora Rescue and Relocation Program, and the implementation of the Wild Fauna Repelling,
Rescue and Relocation Program. The site preparation stage also includes the following actions:
•
•
Removal of vegetation cover and organic soil (clearing and stripping);
Construction or adaptation of temporary roads for access to the works and activities;
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•
•
•
Integral waste management (hazardous, special management and urban solid waste);
Construction of stormwater diversion works; and
Construction and testing of wells for drainage of the open pit mining areas.
Construction will begin with the leveling of the land once the topsoil has been removed. This will
be followed by the construction of access roads, drainage works, water diversion and the
establishment of construction platforms for the administrative building, offices and warehouses.
At the same time, work will begin on the installation of platforms to place diesel electric generators
and fuel tankage.
20.4.2
Operations
The operation of the mine site involves a series of interrelated areas, as one depends on the other
in order to continue with the general work process, starting with the exploitation and mining
activities, followed by the milling, beneficiation plant, doré room, SART circuit, cyanide
detoxification and tailings deposition. Operational details are provided in Sections 16 and 17 of
this Technical Report.
Key environmental aspects of operations will be related to cyanide management, in particular
cyanide destruction in tailings porewater and the storage/disposal of tailings. Free cyanide and
copper-complexed cyanide will be recycled in the leaching process by solution management and
SART processes.
The washed leached tailings slurry from the milling circuit will be treated by the proven sulphur
dioxide-air process to reduce the cyanide concentration of WAD CN (weak acid dissociable
cyanide) to <5 mg/L.
The detoxified tailings will be thickened and filtered and “dry stacked” in an on-site storage
facility.
20.4.3
Closure Planning
In advance of the post-operational stage of the Project, a Closure Plan will be prepared. The main
objective of the Closure Plan is to ensure that the physical and biological environment, where the
mining activity occurred, recovers to a status that will ensure sustainability. The Closure Plan will
include these objectives:
•
Develop a Closure Plan that allows for the physical stability of the components
involved in the Project over a long term;
•
Implement closure activities that allow for an abandonment or passive care solution,
i.e., minimize active maintenance requirements after closure;
•
Develop environmentally sound procedures for demolition and/or decommissioning of
facilities;
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•
Develop a monitoring plan to corroborate whether the closure objectives have been
successfully achieved;
•
Make budgetary provisions for final closure based on cost estimates that will be
reviewed and updated periodically; and
•
Integrate closure approaches that simultaneously interact.
Costs to be incurred after mining were estimated by CIMA, and reviewed by the Authors, to be
$3.7M to close and rehabilitate the Project site. Details of the cost estimate can be found in Section
21 of this Technical Report.
20.5
20.5.1
BACKGROUND INFRASTRUCTURE AND ENVIRONMENTAL CONDITIONS
Historical Mining Activities
The Project region contains the residuals of historical mining activity: the most relevant was the
mining activity which reached an area of 273 ha in 1909. The old mines of Tepantería, Mololoa,
Albarradón, Copala and Cinco Minas were located in the region, as well as the process plant called
Santo Tomás. All were located in the municipality of Hostotipaquillo. The main exploited minerals
were native silver, argentite, miargyrite (AgSbS2), auriferous pyrite and native gold.
Anaconda Mining Company was a more recent owner of the Cinco Minas property. Between 1922
and 1928, 1,083,000 tons of ore were mined at an average grade of 3.17 g/t gold and 476 g/t silver.
This is equivalent to 2,760 kilograms of gold and 427,954 kilograms of silver. Subsequent to this
period, new companies entered and continued with the exploration and processing of minerals such
as TUMI Resources with 1970s activity.
The locations of historical mining activities near the Project location (blue in figure) are shown in
Figure 20.2.
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FIGURE 20.2
REGIONAL HISTORICAL MINE LOCATIONS
Source: CIMA (2024)
In addition to the above, the results of a survey of mining works close to the Project area, including
open pits, ventilation shafts, adits and production shafts are shown in Figure 20.3.
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FIGURE 20.3
CLOSE REGIONAL HISTORICAL MINING WORKS
Source: CIMA (2024)
20.5.2
Historical Environmental Liabilities
Environmental liabilities can be considered to be those sites contaminated by the potential release
of hazardous materials or acid/metal contaminated drainage from historical mine waste. No
specific environmental liability caused by historical mining activity and related to the Project has
been identified by on-ground surveys.
20.5.3
Current Activities within the Environmental System
Within the environmental system, there are different primary activities, the most important of
which are:
Artisanal Mining
As mentioned above, the region has historical and existing mining activities, however, these are
mostly artisanal or "gambusinos".
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Livestock
Cattle ranching is a widespread activity in the local area making it one of the main sources of
livelihood. There are dairy and beef cattle, pigs, goats, and poultry. Cattle have been identified as
causing the greatest impact on the ecosystem due to poor livestock practices, such as overgrazing,
by the owners of the ejido lands in the area.
Agriculture
Rain-fed, wetland and irrigated agriculture is practiced in the area, to assist in the production of
annual, semi-perennial or perennial crops. The activity includes animal-hitched or tractor use,
agrochemicals and use of fertilization-enhanced seeds. The crops grown include corn, sorghum,
maguey, mezcal, chickpeas, peanuts, agave and sugarcane.
20.6
20.6.1
APPLICABLE PROJECT ENVIRONMENTAL LAWS AND REGULATIONS
Studies, Procedures or Authorizations for the Project under Federal
Jurisdiction
SEMARNAT is the agency with the greatest relevance to Mexican mining projects, which are
regulated mostly by the General Law of Ecological Balance and Environmental Protection
(“LGEEPA”) and the General Law of Sustainable Forest Development (“LGDFS”), the former in
environmental matters and the latter in forestry; however, it is not the only regulatory agency; there
are other important agencies such as the National Water Commission (“CONAGUA”), CONANP,
and Instituto Nacional de Antropología e Historia (“INAH”). Regulations are applicable to the
type of activities or stages in which the Project is developed: exploration, construction, operation,
closure and post-closure. The studies, permits and procedures applicable to the Project are listed
Table 20.2.
TABLE 20.2
PERMITS AND PROCEDURES REQUIRED FOR THE APPROVAL OF THE PROJECT
Permits and Procedures
Registration of Hazardous Waste Management
Use of Federal Waterways
Annual Operating Statement
Environmental Risk Study
Technical Justification Study for Land Use Change in Forest Land
Electric Power Feasibility (Electric Power Contract).
Preventive Environmental Impact Report
Single Environmental License
Operating License for Fixed Sources of State Jurisdiction.
Environmental Impact Assessment - Individuals
Environmental Impact Assessment - Individual with risk
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Government
Agency
SEMARNAT
CONAGUA
SEMARNAT
SEMARNAT
SEMARNAT
CFE
SEMARNAT
SEMARNAT
SEMARNAT
SEMARNAT
SEMARNAT
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TABLE 20.2
PERMITS AND PROCEDURES REQUIRED FOR THE APPROVAL OF THE PROJECT
Permits and Procedures
Environmental Impact Assessment - Regional
Environmental Impact Statement - Regional with Risk Assessment
Permitting of Accesses and other Facilities on Free Federal Highways.
Wastewater Discharge Permit
Permit for Storage and Use of Explosives
Permit to Construct Hydraulic Works
Special Handling Waste Management Plan
Mining Waste Management Plan
Hazardous Waste Plan
Accident Prevention Program
Registration of the Company before the Ministry of Health (SS) and
Municipal Administrations of the Sanitary License and Sanitary Control
Cards.
Registration in the Mexican Business Information System (SIEM).
Title of Concession or Assignment of National Water Use (Surface and
Groundwater)
Title of Concession for Extraction of Materials
Unified Procedure for Change of Land Use.
Registration of the Business Registry before the IMSS
Proceedings before CNA for the Installation of a Company not Connected
to the Municipal Grid
Single Environmental License (LAU)
Annual Operation Card (COA)
20.6.2
Government
Agency
SEMARNAT
SEMARNAT
SCT
CONAGUA
SEDENA
CONAGUA
ECOLOGIA
SEMARNAT
SEMARNAT
SEMARNAT
SS
SE
CONAGUA
CONAGUA
SEMARNAT
IMSS
CNA
SEMARNAT
SEMARNAT
Studies, Formalities or Authorizations for the Project Under Municipal
Jurisdiction
The States of Mexico have various laws and regulations of state and municipal jurisdiction, which
should be considered, as well as different procedures or permits. These are listed in Table 20.3.
TABLE 20.3
REQUIREMENTS BY STATE-MUNICIPAL AUTHORITIES FOR THE PROJECT
Procedure
Zoning Certificate
Environmental Risk Study
Preventive Environmental Impact Report
Explosives Safety Inspection
Municipal Construction License
Government Level
Municipality
State
State
SEDENA
Municipality
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TABLE 20.3
REQUIREMENTS BY STATE-MUNICIPAL AUTHORITIES FOR THE PROJECT
Procedure
Government Level
Operating License
Municipality, if no procedure exists, State
Land Use License
Municipality
Environmental Impact Assessment
State
Internal Civil Protection Program
State Civil Protection / Municipality
Registration of Special Handling Waste Generator State
20.6.3
Environmental, Mining Safety and Labour/Social Security Regulations
The exploration, exploitation and beneficiation of materials are subject to the regulation of
standards on which the corresponding procedures for obtaining mining permits must be based.
There are many regulations applicable to the following aspects:
•
•
•
•
•
•
•
•
Air Emissions;
Hazardous Wastes;
Water and Water Resources;
Air Pollution;
Soil Contamination;
Wild Flora and Fauna Protection;
Environmental Impact Studies; and
Exploitation of Mineral Resources.
Labour and Social Security standards that need to be addressed include:
•
•
20.7
Safety and Health Standards; and
Organizational Standards.
ENVIRONMENTAL STUDIES
Environmental and social baselines, as well as community assessments have been performed by
CIMA in a comprehensive report (CIMA, 2024). For its preparation, it was necessary to compile
available information from bibliographic and digital sources, as well as the collection of
environmental and social data in the field to obtain a general diagnosis of the conditions of the
physical, environmental and social environment where the Los Ricos South Project is located. The
preparation of the study began in 2019.
Favourable approvals were received from SEMARNAT in 2019 and 2022 that allowed a mineral
exploration program for the Project.
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20.7.1
Definition of the Environmental System
In order to conduct a comprehensive Environmental Assessment an Environmental System (“ES”)
was defined. The ES was defined as a set of elements that interact and are interdependent, in such
a way that the interrelations can modify one or all of the other components within the region where
the Project is to be developed. As shown above in Figure 20.2 the ES represented 24,600 ha (red
outline) in which the Project (34 ha, blue figure) is located.
20.7.2
Physical Studies
The following are brief summaries of the completed studies:
•
Soil – Three different soil types were identified, with leptosol being the most abundant;
•
Erosion – More than half of the soil surfaces did not demonstrate water erosion; wind
erosion was found to be minimal with an estimated 20 t per year of soil being mobilized;
and
•
Surface hydrology – Overgrazing by livestock has reduced infiltration. Runoff occurs
only during the rainy season. A hydrology water balance is shown in Table 20.4.
TABLE 20.4
ES HYDROLOGICAL BALANCE
Hydrological Balance
Precipitated volume
Volume Evaporated
Runoff
Infiltration
20.7.3
Metres3/Year
207,500,000
162,100,000
24,400,000
20,900,000
Annual
Precipitation
(mm)
845
660
99.4
85.2
%
100.0
78.1
11.8
10.1
Water Quality
Ground water was obtained from four sites south of the Project site. Thirty-four parameters were
determined by analyzing multiple samples. Only three parameters exceeded permissible limits (the
Permissible Limits were not defined by CIMA) and these values occur naturally in the terrain. No
relation to previous mining or current exploration activities was suggested (as analyzed by CIMA).
20.7.4
Climatology
Five different types of climates were identified, with the warm climate being the most influential
in the ES as shown in Table 20.5. Average precipitation and temperature records from the nearby
Hostotipaquillo weather station are shown in Figures 20.4 and 20.5.
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TABLE 20.5
CLIMATE TYPES WITHIN ES
No.
Types of climates
1
2
3
4
5
Total
Subhumid temperate
Very warm semi-dry
Warm subhumid
Subhumid temperate
Semi-warm, subhumid
FIGURE 20.4
Surface
(ha)
760
1,980
17,600
2,410
1,860
24,700
AVERAGE
MONTHLY
HOSTOTIPAQUILLO
Surface
(%)
3.10
8.05
71.5
9.82
7.56
100.0
PRECIPITATION,
STATION
14068
Precipitation (mm)
December
November
October
September
August
July
Jun
May
April
March
February
January
0
50
100
150
200
250
300
Source: CIMA (2024)
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FIGURE 20.5
MAXIMUM, AVERAGE AND MINIMUM TEMPERATURES
Temperature (°C)
35
30
25
20
15
10
5
0
Temperature maximum (°C)
Temperature average (°C)
Temperature minimum (°C)
Source: CIMA (2024)
20.7.5
Biological Characterisation of the ES
A detailed study was completed of the biotic community of the ES and the Los Ricos South Project
area, comprising the living elements that coexist in this specific zone. This community can be
assessed as biodiverse. Furthermore, the community is important because it contains the biological
organisms that represent the dynamic part of the ecosystem, influencing interactions among living
elements and the flows of matter and energy that regulate environmental systems.
To assess the Project's potential impact regarding changes in land use and vegetation, CIMA
conducted a series of comparative analyses. The vegetation, plant associations, and flora of the ES
have been studied and precisely defined by the official Mexican entity responsible for
biogeographical aspects (“INEGI”). However, these definitions have evolved over time as terms
and concepts have been refined, which is typical in dynamic communities with changing species
and associations that suggest certain conditions that may vary over time. This is the case with the
ES under study, classified in terms of vegetation type and land use with certain variations, which
are analyzed further.
The analysis considers an accessible period where information and cartographic data could be
obtained; these are the years 1991, 2005 and 2021, with a time spanning 32 years, which is
pertinent.
20.7.6
Disturbance Index
To determine the condition of the ES region in terms of ecosystem health and vulnerability to
vegetation loss, a disturbance index was applied for a period of seven years. According to the
information generated, the only significant disturbance in the ES would be related to agricultural
activity. The potential Project-related disturbance was designated by the disturbance index as
small.
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20.7.7
Cultural and Human Status
This sub-factor generally encompasses the economic status and health of the populations that are
within or near the ES, and the health of the workers who would be employed in the different
activities of the Project. This is counted with towns with few inhabitants, where their original main
economic activity is agriculture, livestock and mining (gambusinos). Health centres such as
hospitals are located outside the ES, in the Town of Hostotipaquillo. The inhabitants of nearby
rural towns tend to migrate to urban towns in search of work or study.
Village growth is low because there is little diversity of trades or activities that generate
opportunities. In the near future, this trend is not expected to change unless new activities, such as
the Project, arrive and would allow adequate remuneration and permanent jobs.
20.7.8
20.7.9
Modification of the Environment by Project Perimeter Noise, Fugitive Dust
and Use of Dangerous Substances
•
Perimeter noise. The maximum permissible perimeter limits of 68 dB for daytime and
65 dB nighttime are expected to be met. Most of the Project activities to be carried out
in the area will be in agricultural and livestock areas, which should not be significantly
affected by noise generation from the mining site;
•
Fugitive Dust. Once the Project is under construction and/or operation, the monitoring
of suspended particles in the air will be carried out in order measure dust levels
originating from the Project and to determine compliance with ambient dust
environmental regulations; and
•
Environmental Risk Due to the Use of Chemical Substances. Of the total of 19 reagents
that will be used in the Project, the use of sodium cyanide will be the most important.
The Project is anticipated to comply with the standards of the International Cyanide
Code1, a very practical and responsible way to verify the management and containment
of sodium cyanide. This will ensure the protection of Project workers, public persons
as well as the environment.
Summary – Existing Environmental Aspects
The current trend of deterioration of the ES is slow, due in part to the private or ejido property
owners only allow access to local people, so there is little interference from the outside population.
The primary activities such as livestock, agriculture and artisanal mining are the most developed
activities.
According to the existing productive activities and the dominant land use, it is estimated that the
ES will remain stable, while showing slight variations in some of its dominant communities such
as low growth deciduous forest, induced pasture and agriculture.
1
https//cyanbidecode.org, International Cyanide Management Code for the Manufacture,
Transport and Use of Cyanide in the Production of Gold
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20.8
SOCIAL ASPECTS
20.8.1
Population
A social baseline study (Borealis, 2019) considered the economic, cultural, social, demographic
and geographic aspects of the local communities. It involved measuring the state of individuals,
households, and institutions at a specified time; describing initial conditions through appropriate
indicators, prior to the start of the Project to assess progress or make a comparison once completed.
Table 20.6 shows the information available from 2019 and 2020 data from INEGI. The
municipalities of Hostotipaquillo, Magdalena and Tequila sum to a population close to 0.90% of
the total population of the State of Jalisco. The municipality of Hostotipaquillo has 8,730
inhabitants, Magdalena has 21,780 inhabitants, while Tequila has 44,350 inhabitants.
The total population of the municipality of Hostotipaquillo is 50.02% male gender and 49.98%
female gender; the municipality of Magdalena has 48.62% male population and 51.38% female
population, while the municipality of Tequila has 49.36% male population and 50.64% female
population.
TABLE 20.6
POPULATION IN AGE-SEX RELATIONSHIP OF THE MUNICIPALITIES OF
HOSTOTIPAQUILLO, MAGDALENA AND TEQUILA, STATE OF JALISCO
Population
(2020)
Total population
Female population
Male population
Population 15 and over
Female population aged 15
years and over
Male population aged 15
years and over
Population 60 and over
Female population aged 60
and over
Male population aged 60
and over
Men-women relationship
8,730
4,360
4,370
6,210
21,800
11,200
10,600
15,300
44,400
22,500
21,900
31,600
8,350,000
4,250,000
4,100,000
6,190,000
Importance of
municipalities
with respect to
the State (%)
0.90
0.89
0.90
0.86
3,140
8,000
16,000
3,180,000
0.85
3,070
7,300
15,600
3,010,000
0.86
1,220
2,800
4,640
999,000
0.86
630
1,460
2,490
539,000
0.85
590
1,340
2,150
460,000
0.88
1.00
0.95
0.97
0.96
--
HostotipaMagdalena
quillo
Tequila
Jalisco
State
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20.8.2
Public Health Services
The municipality of Hostotipaquillo has a total population of 8,730 inhabitants, of which 7,210 are
entitled to some health service; the municipality of Magdalena has 21,800 inhabitants of which
17,800 have some health service; and the municipality of Tequila has 30,100 of the 44,400
inhabitants who are entitled to some health service.
20.8.3
Housing and Services
According to INEGI 2020, the total number of inhabited private homes in the municipality of
Hostotipaquillo was 2,230, the municipality of Magdalena had 5,530 inhabited homes, while the
municipality of Tequila had 11,110 inhabited homes.
20.8.4
Indigenous Ceremonial Centers
No presence of indigenous peoples, who tend to have a culture more rooted in religious or
ceremonial traditions, were identified. Within the locality of Cinco Minas, no persons linked to an
indigenous ethnic group were identified in surveys.
20.8.5
Local Perceptions about Mining and by Inference, the Project
A series of surveys were conducted in 2021 to obtain the variables that would allow the generation
of the main socioeconomic indicators of the community for its characterization. Surveys were
conducted in three communities (Huajacatlan, Cinco Minas and Sayulimita); the fieldwork
consisted of 54 simple questions that addressed the following topics:
•
•
•
•
•
•
•
Home;
Economy;
Supports;
Health;
Sources of employment;
General environmental; and
Community problems.
Within the 54 questions that made up the survey, specific questions of perception were asked, such
as perception and expectations of mining. Example comments by Cinco Minas residents indicated:
•
The population of the Cinco Minas community generally expressed a positive
perception of mining activity;
•
A minor sector of Cinco Minas viewed mining as a negative factor that affects the
natural environment; and
•
The citizens hoped for an economic boost and support for the community.
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In general, the local citizens cautiously supported a significant mining development, such as the
Project. In addition, there appeared to be reasonably strong support by the municipalities for
mining development, such as the Project.
20.8.6
Environmental Impacts Resulting from the Development and Operation of the
Project
CIMA conducted a wide-ranging review of the potential impacts of planning, development,
operational and closure stages of the Project. The detailed review processes and impact results
using a Modified Leopold Matrix, and results can be found in the CIMA 2024 report.
Overall, 50 impacts were classified as resulting in very low impacts, 56 as low impact, 18 as
medium impact and only one as possible high impact. The high impact item corresponds to the
impact on the availability of water in the regional aquifer. This is due to the nature of the Project’s
potential future underground mine and its geographic location.
Possible archaeological zones were also investigated, with information available in the GIS
(Geographic Information Systems). The study concluded that there are no archaeological zones
within the ES, however, it is necessary to consult the Instituto Nacional de Antropología e Historia
(“INAH”) to rule out the existence of an archaeological site.
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21.0
CAPITAL AND OPERATING COSTS
21.1
INTRODUCTION
All costs are in Q4 2024 US dollars. No provision has been included in the cost estimates to offset
future escalation. Contingencies of various percentages have been added to capital costs
(“CAPEX”) and average approximately 10%. No contingency has been added to operating costs
(“OPEX”).
The total initial CAPEX of the Los Ricos South Project is estimated at $227M including $21M in
contingency costs over an expected two-year build. Sustaining capital costs incurred during the 15
production years are estimated to total $100M. Total OPEX over the life-of-mine (“LOM”) is
estimated at $915M which averages $89.43/t processed. The following subsections provide details
of these costs.
The capital and operating costs presented in this Feasibility Study (“FS”) provide substantiated
costs that can be used to assess the economics of the Los Ricos South Project. The scope of the
estimates includes:
•
A combined underground and open pit mining operation;
•
The construction and operation of a process plant;
•
The construction and operation of a tailings storage and management facility and paste
backfill facility;
•
The construction and operation of Project infrastructure required to support the
operation; and
•
Owner’s costs and contingency.
The estimates conform to Class 3 guidelines for an FS-level estimate with a ± 15% accuracy
according to the Association for the Advancement of Cost Engineering International (“AACE”
International).
21.2
CAPITAL COSTS
21.2.1
Introduction
The Los Ricos South Project has been planned as an underground mining operation with open pit
mining in the final five years of mine life. Contract underground mining is planned in production
years one to 13 of the mine plan, with contract open pit mining starting in year 10 and ending in
year 15. Both operations are planned to produce simultaneously for four years.
The process plant is comprised of conventional crushing and grinding followed by cyanide tank
leaching. Back-end filtration is required to maximize water recycling (for dry stack tailings) as
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well as a SART (sulphidation, acidification, recycling and thickening) circuit to re-generate
cyanide back to the process and to produce a saleable Cu2S copper sulphide product. Water supply
to the process plant is provided by a nearby seasonally re-charged water dam, and high voltage
grid power is provided by the local utility.
The initial capital cost estimate addresses the engineering, procurement, construction and start-up
of the Los Ricos South Project, with a construction period to develop an underground mine, build
a process plant capable of treating 2,000 tpd, prepare a dry stack tailings management area, and
install associated ancillary surface facilities.
The total estimated cost to design, procure, construct and start-up the underground mine and
process plant facilities described in this Technical Report is $227M. Table 21.1 summarizes the
capital cost estimate. The estimated cost includes a contingency allowance of $21M.
Sustaining capital represents capital expenses for additional costs that are not included in the
normal operating costs, equipment purchases that will be necessary during the operating life of the
Project, tailings storage capacity increases, all capital costs associated with underground mining,
and start-up of open pit mining. Life-of-mine (“LOM”) sustaining capital is estimated at $100M.
Total capital costs over the LOM are estimated at $326.4M. Items not included in the capital cost
estimate are:
•
•
•
•
•
•
Sunk costs and costs prior to the start of basic engineering phase;
Escalation;
Working capital;
Interest and financing costs;
Taxes and duties; and
Reclamation and associated bonding requirements.
TABLE 21.1
CAPITAL COST SUMMARY
Item
Process Plant Directs & Tailings
Underground Mine Development
Open Pit Pre-Stripping
Infrastructure
EPCM
Project Indirects
Subtotal
Contingencies (~10%)
Total
1
Initial
($M)
83.54
51.05
0.00
38.56
17.97
14.58
205.70
20.99
226.69
Sustaining
($M)
10.22
62.76
17.66
0.00
0.00
0.00
90.65
9.07
99.71
Total
($M)1
93.77
113.82
17.66
38.56
17.97
14.58
296.35
30.05
326.40
Totals may not sum due to rounding.
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The capital cost estimate was established using a hierarchical work breakdown structure (“WBS”)
as presented in Table 21.2. The estimate includes costs associated with the following major areas:
mining, process plant, tailings facilities, on-site infrastructure, off-site infrastructure, Project
indirect costs (“indirects”), Project delivery, and contingency (“provisions”).
TABLE 21.2
WBS BREAKDOWN OF CAPITAL COSTS
Description
WBS
Capital
Cost
($M)
Mining Infrastructure
1100
9.89
0.00
9.89
Open Pit Mining
1200
0.00
17.66
17.66
Underground Mining
1300
51.06
62.75
113.81
Process Plant
2000
75.62
0.00
75.62
SART Plant
2700
4.32
0.00
4.32
Tailings Facilities
3000
3.60
5.42
9.02
On-Site Infrastructure
4000
16.01
4.11
20.12
Off-Site Infrastructure
5000
12.65
0.0
12.65
173.15
89.94
263.10
Total Direct Costs
Sustaining
Cost
($M)
Total
Cost
($M)1
Indirect Costs
6000
4.56
0.70
5.26
Project Delivery
7000
17.97
0.00
17.97
Owner’s Costs
8000
10.02
0.00
10.02
Contingency
9000
20.99
9.07
30.06
Total Indirect Costs
53.54
9.77
63.31
Total Capital Cost
226.69
99.71
326.40
Notes::1 Totals may not sum due to rounding.
Data for the estimates have been obtained from numerous sources, including:
•
•
•
•
•
•
Mine schedules;
FS-level engineering design by the consulting firms that contributed to this Technical
Report;
Budgetary equipment quotes from suppliers;
Geotechnical investigations;
Budgetary unit costs from local contractors for civil, concrete, steel, electrical, piping,
mechanical works; underground and open pit mining; and
Data from similar recently completed studies and projects.
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Major cost categories (permanent equipment, material purchase, installation, subcontracts, indirect
costs, and Owner’s costs) were identified and analyzed. A contingency was applied to the cost
estimates based on an assessment of the data source by discipline and by WBS level 3, and
applying a probabilistic method (Monte Carlo Simulation). An overall contingency amount was
derived in this fashion.
The capital cost estimates for WBS 2000 to 5000 items, excluding WBS 2700 SART Plant, were
developed based on firm and budgetary quotations for equipment and construction contracts, as
well as Ausenco’s in-house database of projects and studies including experience from similar
operations. A total of 83% of Ausenco’s scope were firm or budget quotes, 8% were from a
database, 8% were factored, and 1% was allowance.
Vendors and contractors were requested to price in native currency. The capital cost estimates are
prepared in the base currency of USD, where relevant exchange rates were used to convert to US
Dollars. Table 21.3 shows the foreign exchange rates used in converting pricing to USD.
TABLE 21.3
EXCHANGE RATES
Currency
Abbreviation
AUD
EUR
CAD
MXN
USD
21.2.2
21.2.2.1
Currency
Australian Dollar
Euro
Canadian Dollar
Mexican Peso
United States Dollar
Symbol
AU$
€
C$
MX$
US$
Exchange
Rate
0.6477
1.0966
0.7233
0.0511
1.00
Initial Capital Costs
WBS 1100 Mining Infrastructure
The mine infrastructure capital covers a maintenance truck shop, a paste backfill plant and
preparation of the run-of-mine (“ROM”) pad and ore stockpile area. The total cost of the mine
infrastructure capital is estimated at $9.9M.
Budgetary quotations for the mine truck shop were obtained for both pre-engineered metal building
(“PEMB”) and fabric building options. After evaluating the vendor quotations and considering
factors such as cost-effectiveness, construction schedule, and specific Project requirements, the
fabric building was determined to be the most appropriate solution and was subsequently
incorporated into the final design at an estimated cost of $1.0M.
A vendor was selected for the paste backfill plant based on a proposal which met all key technical
and operational criteria, including equipment quality, system integration, and compliance with
safety and performance standards. The chosen vendor ultimately provided the optimal balance of
cost, quality, and technical fit. The cost of the paste backfill plant is estimated at $8.7M.
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The initial capital cost to construct the ROM pad and ore stockpile area is estimated at $0.2M.
21.2.2.2
WBS 1300 Underground Mine Development
This subsection addresses the capital cost estimate for engineering, procurement and start-up costs
of the UG mine at the Los Ricos South Project. The UG is expected to be mined with contractors,
reducing the overall capital costs versus owner operation. Costs for development, production and
haulage were based on three contractor quotes. All costs used for UG development and production
are either sourced directly from quotes or calculated from first principles from the contractor costs.
Total CAPEX cost for the underground, including both initial and sustaining CAPEX, is estimated
at $113.8M, as shown in Table 21.4.
TABLE 21.4
CAPITAL COST SUMMARY FOR UNDERGROUND PORTION OF PROJECT
Area
Lateral Development
Vertical Development
Capitalized OPEX Development
Services
Backfill Delivery System
Other CAPEX
Other Capitalized OPEX
Total
Initial CAPEX
($M)
27.55
2.13
4.28
6.53
0.13
1.44
9.00
51.06
Sustaining CAPEX
($M)
46.78
5.92
0.00
7.21
0.31
2.54
0.00
62.75
Total CAPEX
($M)1
74.33
8.06
4.28
13.74
0.44
3.97
9.00
113.81
Notes::1 Totals may not sum due to rounding.
Major initial capital expenditures for the UG, prior to production, include: the construction of a
backfill reticulation system; 13 km of underground development; three portals (two access, one
ventilation); and the installation of supporting services (electrical power, ventilation, compressed
air and dewatering). Initial capital costs for the UG portion of the Project are estimated at a total
of $51.1M.
Development
A total of 12.5 km of lateral development and approximately 0.6 km of vertical development will
be driven during the pre-production period spanning YR -2 and YR -1. Of this, approximately
2.4 km of development would normally be classified as OPEX, however, is reclassified as CAPEX
due the costs being incurred during the pre-production period. Table 21.5 provides a summary of
development quantities, cost per heading size, and total costs incurred during the pre-production
period for development.
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TABLE 21.5
INITIAL DEVELOPMENT CAPITAL COSTS
Heading Type
Heading Profile
(m W x m H)
Vent Raise (Drop Raise)2
3.0 x 3.0
Vent Raise (Raisebore)
3.0 dia.
Subtotal – Vertical Development
Ramps, FW Drifts, and 4.0 x 4.0 (FWDs)
Level Accesses
4.5 x 4.5 (Rest)
Remucks
4.5 x 5.5
Infrastructure and Vent 4.5 x 4.5 (Adits)
Accesses
4.0 x 4.0 (Rest)
5.0 x 6.5
Maintenance Bays and Shop
Capitalized OPEX: Lateral
4.0 x 4.0
Development
Capitalized OPEX: Back
4.0 x 4.0
and Wall Waste Slashing3
(equivalent)
Subtotal – Lateral Development
Total (Pre-contingency)
Required
Development
Development
Avg Cost
(m)
($/m)
347
2,554
213
5,866
3,813
560
7,688
2,815
Total
CAPEX
($M)1
0.9
1.2
2.1
21.6
679
1,719
2,581
2,339
1.8
4.0
37
1,274
3,582
2,436
0.1
3.1
1,157
1,013
1.2
12,555
13,114
2,535
2,590
31.8
34.0
Notes::1 Totals may not sum due to rounding.
2
Drop raise profile dimensions are in mW x mL.
3
Estimated volume of slashing is 18.5 k m3, equivalent to 1,157 m of 4.0 mW x 4.0 mH development.
Total initial development CAPEX is estimated at $34.0M. Of this total, approximately $4.3M
would normally be categorized as OPEX, however, this amount is reclassified as CAPEX due to
the costs being incurred during the pre-production period.
Services
Underground services initial CAPEX includes:
•
•
•
•
•
Ventilation fans and doors;
Electrical transformers and substations;
Dewatering pumps;
Compressors and accumulators; and
Communications systems.
Total initial CAPEX for services infrastructure is estimated at $6.5M.
Backfill Distribution System
The backfill plant will be located near the Eagle Zone portal, slightly below the entry to the
underground mine, necessitating the use of a positive displacement pump to move the paste
backfill into the mine, prior to utilizing reticulation by gravity through drill holes to fill the stopes.
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The backfill plant will be operational prior to initial stoping and will utilize tailings from
processing mineralized development material for initial fills. Total initial CAPEX for the backfill
plant reticulation system is estimated at $0.1M.
Other CAPEX
Other initial CAPEX includes:
•
Fitment of explosives magazines, refuge stations, maintenance bays, pump stations,
and escapeways;
•
Construction of UG/surface interfaces (portals and ventilation raise collars);
•
Contractor mobilization; and
•
Technical services software purchases.
Total initial CAPEX for infrastructure and fitment is estimated at $1.4M.
Other Capitalized Operating Costs
Outside of capitalized OPEX development, other capitalized OPEX costs include:
•
•
Owner’s costs, including:
o Technical services and Company mine supervision staff salaries;
o Consultants;
o UG Dayworks and sundries; and
Electrical power.
Items included in this category (power, Owner’s costs) would normally be classified as OPEX,
however, are categorized as CAPEX since costs are incurred prior to the start of production. The
total cost of these items is estimated at $9.0M during the pre-production period.
21.2.2.3
WBS 2000 to 4000 Process Plant, Tailings Facility and On-Site Infrastructure
Capital Costs
A summary of the process plant, tailings facility and on-site infrastructure capital cost is presented
in Table 21.6. A percentage of the costs by major discipline is presented in Table 21.7.
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TABLE 21.6
CAPITAL COST ESTIMATE SUMMARY - PROCESS PLANT, TAILINGS MANAGEMENT
FACILITY AND ON-SITE INFRASTRUCTURE
WBS Description
WBS
Crushing
Feed Stockpile and Reclaim
Grinding
Leaching and Solution Management
CCD, Detox and Tailings Filter
Merrill Crowe and Refinery
SART Plant Preparation
Process Plant Services
Tailings Facility
Bulk Earthworks
Surface Water Management
Switchyard and Power Distribution
Fuel Storage
Waste Management
Infrastructure Buildings
Site Services
Surface Mobile Equipment
Total
2100
2200
2300
2400
2500
2600
2700
2800
3100
4100
4200
4300
4400
4500
4600
4700
4800
Initial
Capital
Cost
($M)
6.07
1.41
14.90
7.98
15.99
15.36
0.62
13.29
3.60
3.41
0.24
4.36
0.44
0.15
4.71
2.09
0.62
95.24
Sustaining
Capital
Cost
($M)
5.42
4.09
9.51
Total
Cost
($M)1
6.07
1.41
14.90
7.98
15.99
15.36
0.62
13.29
9.02
3.41
0.24
4.36
0.44
0.15
4.71
2.09
4.71
104.75
Note:1 Totals may not add due to rounding.
TABLE 21.7
COST SUMMARY BY MAJOR DISCIPLINE
Disc.
A
B
C
E
F
I
M
N
P
Commodity Description
Architectural
Earthworks
Concrete
Electrical Equipment & Bulks
Platework
Instrumentation + Bulks
Mechanical Equipment
Mobile Equipment & Ancillaries
Pipework
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% of LOM
Cost
2.5
8.5
4.7
11.4
5.2
3.1
26.6
3.1
5.6
Page 455 of 577
TABLE 21.7
COST SUMMARY BY MAJOR DISCIPLINE
S
U
V
W
Z
Structural Steel
Field Indirects
Third-Party Packages/Other
Project Delivery
Taxes & Duties
Total
1.7
1.7
1.7
11.6
10.3
100
Estimate Sources
The process equipment requirements were defined based on process flowsheets and process design
criteria, as defined in Section 17. All major equipment was sized based on the process design
criteria to derive a mechanical equipment list. Mechanical scopes of work were developed and sent
for tender and budgetary pricing to equipment suppliers. For mechanical equipment costs, 34% of
the value was firm quotes and 60% of the value was sourced from budgetary quotes. The remainder
was sourced by benchmarking against other recent Mexican and South American mining projects
and studies.
Similarly, the major electrical equipment was sized based on the Project’s equipment list. Scopes
of work were developed to receive budgetary pricing from equipment suppliers. For electrical
equipment, 28% of the value was firm quotes and 66% of the value was sourced from budgetary
quotations. The remainder was sourced by benchmarking against other recent Mexican and South
American flotation concentrator mining projects and studies.
In support of the major installation construction contracts, engineering for the process plant and
infrastructure was completed to an FS level of definition. After the derivation of all the bulk
material quantities (earthworks, concrete, steel, piping, cables, etc.), for the process plant, tailings
management facility and surface infrastructure areas, major construction contracts were formed
and sent to the market for supply and installation rates.
Specific to the tailings management facility, the design and material take-off outlined in Section
18 was applied against the contractor installation rates quotes for typical activities such as clearing,
topsoil removal, excavation, backfilling with granular material and mine waste. The costs
associated with this estimate also included liner placement, drains, ground improvement, and
associated supervision, monitoring and indirect costs.
Table 21.8 presents the basis of mechanical equipment supply and Table 21.9 provides a package
list of the mechanical equipment. Table 21.10 shows the electrical equipment supply cost basis
and Table 21.11 presents the electrical equipment basis by package.
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TABLE 21.8
MECHANICAL EQUIPMENT SUPPLY PRICE BASIS
Source
Firm Quote
Budgetary Quote
Database
Factored
Allowance
Total
Initial
($M)
10.59
18.35
1.77
0
0
30.71
Sustaining
($M)
0
0
0
0
0
0
LOM*
($M)
10.59
18.35
1.77
0
0
30.71
% of
LOM
34
60
6
0
0
100
*Note: costs exclude freight.
TABLE 21.9
MECHANICAL EQUIPMENT PACKAGE LIST
Package No.
P0001
P0002
P0003
P0004
P0005
P0006
P0007
P0008
P0009
P0010
P0011
P0012
P0013
P0014
P0015
P0016
P0017
P0018
P0019
P0021
P0022
P0101
P0102
P0103
Equipment
SAG Mill
Thickeners
Tailings Filter
Conveyors
Cyclones
Primary Crusher
Apron Feeders
Screens
Leaching Tanks
Gold Room
Cranes and Hoists
Agitators
Pumps
Mine Water Treatment Plant
Sewage Treatment Plant
Fire Water Pump Skid
Piping Supply
Manual Valves
XV and Actuated Valves
Rock Breaker
Front End Loaders
Metallurgical Laboratory
Diesel Storage
Air Compressors and Blowers
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TABLE 21.9
MECHANICAL EQUIPMENT PACKAGE LIST
P0104
P0105
P0106
Lime System
Reagent System
Paste Plant
TABLE 21.10
ELECTRICAL EQUIPMENT SUPPLY PRICE BASIS
Source
Firm Quote
Budgetary Quote
Database
Factored
Allowance
Total
Initial
($M)
11.92
27.79
2.17
0
0
41.89
Sustaining
($M)
0
0
0
0
0
0
LOM*
($M)
11.92
27.79
2.17
0
0
41.89
% of
LOM
28
66
5
0
0
100
*Note: costs exclude freight.
TABLE 21.11
ELECTRICAL EQUIPMENT SUPPLY PRICE BASIS
Package No.
P0201
P0202
P0203
P0204
P0205
P0206
Equipment
Transformers
HV Substation and Switchyard
Prefabricated Electrical Rooms, Switchgear, MCCs, VFDs, I/O Cabinets
Generators
Process Control System
Power Factor Correction and Harmonic Filter Banks
The material commodity codes used for the Project are presented in Table 21.12. Material
commodity codes are a hierarchal method of grouping different materials and equipment that
contribute to building up the cost estimate. The construction contact packages are listed in Table
21.13.
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TABLE 21.12
MATERIAL COMMODITY CODES
Commodity
Code
A
B
C
D
E
F
I
M
N
O
P
S
U
V
W
Y
Z
Commodity Description
Architectural
Earthworks
Concrete
Mining
Electrical Equipment + Bulks
Platework
Instrumentation + Bulks
Mechanical Equipment
Plant & Ancillary Equipment
Mobile Equipment
Pipework
Structural Steel
Field Indirects
Third-Party Packages/Other
Project Delivery
Owner’s Costs
Taxes & Duties
TABLE 21.13
CONSTRUCTION CONTRACT PACKAGES
Package No.
C0501
C0502
C0503
C0504
C0505
C0506
C0507
C0508
C0509
C0511
Equipment
Earthworks
Concrete Works
Structural Steel, Mechanical, Platework and Piping Works
Electrical and Instrumentation Works
Pre-Engineered Buildings
Modular Buildings
Steel Fabrication Supply Pricing
Electrical Bulks Supply Pricing
Instrument Bulks Supply Pricing
Fabric Buildings
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Freight and Logistics
Freight and logistics costs were estimated as a percentage based on Ausenco's experience with
similar projects. For inland freight from Canada or USA to site, a percentage of 12% was used.
For inland freight within Mexico to site, a percentage of 8% was used. For ocean freight, a
percentage of 14% was used. For vendor supplied freight allowance, a percentage of 13% was
used. This approach was consistent with the method used for estimating spares, where an initial
partial quote was provided and then adjusted with regional reference data. This process helped to
ensure that the final logistics costs were accurate and aligned with local conditions. Total freight
costs are estimated to be $7.2M. The freight and logistics costs are built into the costs for the other
disciplines and are within the other WBS sections.
Growth
A growth allowance has been allocated to each line item in the capital cost estimate to reflect the
level of definition of design and pricing strategy, which is a provision for additional costs that will
be recognized in future Project phases as engineering is advanced.
Estimate growth is:
•
Intended to account for items that cannot be quantified based on current engineering
status that are empirically known to appear;
•
Accuracy of quantity take-offs and engineering lists based on the level of engineering
and design undertaken at FS level; and
•
Pricing growth for the likely increase in cost due to development and refinement of
specifications as well as re-pricing after initial budget quotations and after finalization
of commercial terms and conditions to be used on the Project.
Growth has been calculated on a line-item level by evaluating the status of the engineering scope
definition and maturity and the ratio of the various pricing sources for equipment and materials
used to compile the estimate. The growth rate applied was based on guidance aligning to a Class
3 AACE estimate, and the level of definition of the Project scope. The capital cost growth
allowance is presented in Table 21.14.
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TABLE 21.14
CAPITAL GROWTH ALLOWANCE
Commodity
Code
A
B
C
D
E
F
I
L
M
N
P
21.2.2.4
Discipline
Architectural
Earthworks
Concrete
Demolition & Relocation
Electrical Equipment
Platework
Instrumentation
Electrical Bulks
Mechanical Equipment
Mobile Equipment & Ancillaries
Piping
Growth
Applied
8%
8%
8%
8%
6%
8%
8%
8%
6%
6%
10 %
WBS 2700 SART Plant Cost Estimate
The initial CAPEX for the SART facility is estimated at $4.3M including all direct equipment and
construction costs as well as associated indirect costs.
The SART capital cost estimate was prepared at a Feasibility Study level to evaluate the potential
viability of SART within the Project’s overall processing costs. Estimated capital costs for the
SART plant, including direct and indirect costs, are based on the GoGold plants in Parral, Mexico.
Table 21.15 summarizes the estimated SART capital costs.
TABLE 21.15
SART PLANT CAPITAL COST ESTIMATE
Section
Description
100
200
300
600
Subtotal
Copper Circuit
Neutralization Circuit
Reagent Circuit
Utilities Circuit
Total Direct Equipment
Construction Direct – Installation
Plant SART Capex Total (without contingency)
Total
Cost
($M)1
0.59
0.92
0.29
0.08
1.88
2.44
4.32
Source: BQE Water (2025) and D.E.N.M. (2025).
1
Totals may not sum due to rounding.
Estimated costs for major direct mechanical equipment are based on quotations provided by major
vendors as well as updated costs provided in the Project’s prior PEA (P&E, 2023). Vendors
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provided specifications from the SART process design criteria discussed in Sections 13 and 17 of
this Technical Report.
Minor direct mechanical equipment costs are based on a database and the Author’s familiarity with
costs of the equipment. Bulk material costs such as bins, tanks and structures were based on builtup rates, factored costs and allowances.
The proposed SART circuit cost is also based on a database developed from similar size systems
designed and constructed during 2019 to 2023.
Direct construction costs include materials and installation for the SART process area site,
concrete, steel work, mechanical, piping, sub-station, MCC electrical and instrumentation.
21.2.2.5
WBS 5000 Off-Site Infrastructure Capital Costs
A summary of the off-site infrastructure cost estimate is presented in Table 21.16 and is estimated
at $12.7M.
TABLE 21.16
OFF-SITE INFRASTRUCTURE CAPITAL COSTS
WBS Description
WBS
Access Road
Water Supply – Overland Water Pipeline
Water Supply – Dam Capacity Upgrades
Power Supply
Total
5100
5210
5220
5300
Initial
Capital
Cost
($M)
0.80
0.59
4.00
7.26
12.65
Sustaining
Capital
Cost
($M)
0
0
0
0
0
Total
Cost
($M)1
0.80
0.59
4.00
7.26
12.65
Note:1 Totals may not sum due to rounding.
A local Mexican contractor provided a quote to upgrade the 13.7 km gravel access road from the
town of San Simon to the Project so that there would be no concerns about transporting equipment
and material to site. The cost was estimated at $0.8M to improve the existing surface, clean and
shape water ditches, and place a 15 cm layer of compacted gravel on the road.
The water supply cost estimate includes two barge pumps at the La Labor Dam with 11.85 km of
DN50 HDPE piping to the Project site, and two barge pumps at the Los Huajes Dam with 6.33 km
of DN50 HDPE piping to the Project site. The pumps and piping have been estimated at 0.6M. A
quote for $4.0M was provided by a Mexican contractor to raise the containment walls of the two
dams to provide adequate water supply over the LOM.
Electrical power to the Project will be supplied by the national power company CFE. A dedicated
two circuit 41 km overhead power line will be installed from the La Yesca Dam where CFE has a
power generation plant and electrical substation. The routing of the line will be from the La Yesca
Substation through Hostotipaquillo Municipally and then to the Project site. Two overhead power
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lines will connect 34.5 kV, three phase and 60 Hz via a sub-station located near the process plant.
The installation contract amounts to $8.0M of which $0.74M has been paid by GoGold, with
$7.26M remaining.
21.2.2.6
WBS 6000 Indirect Costs
Project indirect initial capital costs are presented in Table 21.17 and are estimated at $4.6M.
TABLE 21.17
INDIRECT CAPITAL COSTS
WBS Description
Temporary Construction Facilities and
Services
Commissioning Reps and Assistance
Spares
First Fills & Initial Charges
Modifications Team
Miscellaneous Preliminary Costs
Total
WBS
Initial
Capital
Cost
($M)
Sustaining
Capital
Cost ($M)
Total
Cost
($M)1
6100
2.7
0.0
2.7
6400
6500
6600
6800
6900
0.7
0.6
0.5
0.1
0.7
4.6
0.0
0.7
0.0
0.0
0.0
0.7
0.7
1.3
0.5
0.1
0.7
5.3
Note:1 Totals may not sum due to rounding.
Temporary Construction Facilities and Services
Contractor indirect costs are related to the contractor’s direct costs, and include the following:
•
•
•
•
•
•
•
Mobilization and demobilization;
Site offices and utilities;
Construction equipment including mobile equipment, scaffolding, safety supplies, etc.;
Head office costs/contribution;
Financing charges;
Insurances; and
Profit.
Contractors provided indirect costs as part of their pricing schedules. The total estimated cost for
temporary construction facilities and construction services is $2.7M.
Construction Camp
No costs have been allocated for a permanent camp. It has been deemed that the permanent
workforce will be using local city accommodation and bussed daily to site.
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No costs have been allocated for a dedicated construction camp, as each of the construction
contractors has included provisions for providing temporary accommodation and messing costs
for their own workforce.
Vendor Representatives
Vendor representative costs during commissioning and construction include vendor representative
support during the installation of the purchased equipment and during commissioning of the
equipment.
Vendor representative costs have been estimated based on previously completed Mexico and South
American leach mining projects. The total vendor representatives and assistance cost is estimated
to be $0.7M.
Spares
Commissioning spares were recommended and priced by equipment suppliers. Where equipment
pricing was not solicited from vendors, factors were applied based on standard estimating
practices.
Capital spares prices for mechanical equipment are based on the prices provided by equipment
vendors during the enquiry process. If vendors did not provide a cost for capital spares, a factored
allowance was included based on the supply price and benchmarked against Ausenco’s in-house
database of projects. Allowance factors were based on a six-month period of capital spares.
The total cost of spares is estimated to be $1.3M.
First Fills and Initial Charges
The first fills include the costs for the initial construction, first fills for installed equipment and
commissioning first fills which consist of chemicals, fuels and lubricants etc. and is calculated cost
from OEM suppliers and reviewed with the OPEX estimate.
Commissioning first fills costs were developed from the OPEX costs, and the cost of the initial fill
is included in the estimate.
The total first fills cost is estimated to be to be $0.5M.
Modifications Team
The modifications team cost and is an allowance for the Subcontractor Management Plan and
electrical and instrumentation contractors to be on site during plant commissioning to perform
construction work. The total modifications team cost is estimated to be $0.1M.
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Other Temporary Construction Costs
The temporary construction facilities were developed from budget quotations, historical data and
allowances and cover temporary signage, office consumables and janitorial services, medical and
personal protective equipment supplies, third-party surveying, security, light plants, rigging, rental
equipment, additional scaffolding, washrooms, temporary generators, fuel bulks, heavy lift cranes,
etc.
Costs were developed by each consultant to cover their scope of work and summarized in the
estimate to cover the construction duration. Temporary construction costs are estimated to be
$0.7M.
21.2.2.7
WBS 7000 Project Delivery
Engineering, Procurement and Construction Management (“EPCM”) services cover such items as
engineering and procurement services (home office based), construction management services
(site based), field inspection and expediting, commissioning, corporate overhead, fees and profit.
The EPCM estimate has been developed from a deliverables list and by identification of resources
over a defined schedule.
EPCM costs for the SART Plant and underground mine development are included within their
direct costs.
The total estimated EPCM cost is $18.0M.
21.2.2.8
WBS 8000 Owner’s Costs
Owner’s costs are estimated at $10.0M during the two-year pre-production period, as summarized
in Table 21.18.
TABLE 21.18
OWNER’S COSTS
Item
Construction Management
Site G&A
Process Plant (3 to 6 months)
Underground Mining
Site General Expenses
Mineral Tenure
Insurance
Total
No. Staff
4
47
142
23
Lump sum
Lump sum
Lump sum
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Pre-Production
Cost
($M)
0.49
1.54
0.75
1.70
3.10
1.90
0.55
10.02
Page 465 of 577
21.2.2.9
WBS 9000 Contingency
An overall contingency equivalent to approximately 10% was applied to the capital cost estimate.
The total contingency on initial capital costs is $21.0M.
Contingency accounts for the difference in costs between the estimated and actual costs of
materials and equipment. The level of contingency varies depending on the nature of the contract
and the client's requirements. Due to uncertainties at the time the capital cost estimate was
developed (in terms of the level of engineering definition, basis of the estimate, schedule
development, etc.), it is essential that the estimate includes a provision to cover the risk from these
uncertainties.
For Ausenco scope items such as the process plant and site surface infrastructure, a contingency
rate of 10% to 15% has been used based on a Class 3 AACE estimate and the level of definition
of the Project scope. To develop the contingency value, a Probabilistic Contingency analysis was
performed which consisted of a contingency ranging workshop taking place internally and
evaluated the major cost components in terms of confidence of pricing and quantity basis and
provided input ranges for potential underrun/overrun. The ranging inputs were applied as
percentages to the base estimate and then ran in a Monte Carlo model using the @Risk program.
No contingency has been included for “project specific risks” such as items noted in Exclusions
below. The @Risk simulation completed for Ausenco’s scope determined the contingency to be
12.4% of base accumulative costs at P50 confidence level.
The following contingency percentages were applied:
•
•
•
•
Mining – 10%;
Process Plant and Site Surface Infrastructure – 12.4% (P50 confidence level based
@Risk simulation);
SART Plant – 10%; and
Off-site contracts with a known fixed cost such as electrical power installation, water
dam capacity increases and main access road upgrade – 0%.
Exclusions
The following costs and scope are excluded from the capital cost estimate:
•
•
•
•
•
•
•
•
•
Operating costs (except capitalized mining costs and training of new operations
employees during capital period);
Taxes and duties;
Future exploration costs;
Environmental approvals;
Special incentives (schedule, safety or others);
No allowance has been made for loss of productivity and/or disruption due to religious,
union, social and/or cultural activities;
Escalation beyond the base date Q4 2024;
Environmental impact assessment;
Future scope changes;
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•
•
•
•
•
21.2.3
Demolition and salvage of any existing on-site structures;
Lost time due to weather, labour availability and disruption or force majeure events;
Training of operations personnel;
Management reserve; and
Financing costs.
Sustaining Capital Costs
Sustaining capital costs are estimated to total $99.7M including contingency as presented in Table
21.1 above. A 10% contingency on sustaining CAPEX is estimated at $9.1M.
Sustaining items include incremental capacity increases of the dry stack tailings area in years 2, 6
and 10 to store tailings material over the LOM and are estimated at $5.4M before contingency.
The total LOM sustaining costs for the process plant are estimated at $4.8M before contingency.
Underground mine sustaining costs are estimated at $62.8M over the UG LOM, production years
1 to 13. Development costs account for $52.7M of the total sustaining cost.
Contractor mobilization and pre-stripping of the open pits are required in order that open pit
production can commence. Open pit infrastructure also includes items such as haul road
construction, Owner’s technical team equipment, and preparation of areas for explosives storage
and contractor maintenance facilities. Open pit mining contractor mobilization is estimated at
$0.5M. Pre-stripping requirements are estimated at 5.0 Mt rock at a unit cost of $2.56/t for a total
of $12.8M. An additional $4.4M is estimated for haul road construction, explosives storage,
Owner’s costs and contractor demobilization. Total sustaining CAPEX for open pit items is
estimated at $17.7M as presented in Table 21.19.
TABLE 21.19
OPEN PIT SUSTAINING CAPITAL COSTS
Open Pit Item
Haul Road Construction
Bench Access Trails
Explosive & Emulsion Pad
Explosive Security, Fence
Contractor Mob / Demob
Open Pit Pre-Stripping
Dewatering Systems
Maintenance Shop Laydown
Survey Equipment
Office Equipment & Software
Communications Equipment
Years
9 and 10
($M)
2.44
0.15
0.05
0.05
0.50
12.80
0.00
0.10
0.10
0.10
0.05
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Years
11 to 15
($M)
0.00
0.15
0.00
0.00
0.70
0.00
0.10
0.00
0.00
0.00
0.00
Total
($M)
2.44
0.30
0.05
0.05
1.20
12.80
0.10
0.10
0.10
0.10
0.05
Page 467 of 577
TABLE 21.19
OPEN PIT SUSTAINING CAPITAL COSTS
Open Pit Item
Pickup Trucks
Total Open Pit CAPEX
21.3
OPERATING COSTS
21.3.1
Introduction
Years
9 and 10
($M)
0.36
16.71
Years
11 to 15
($M)
0.00
0.95
Total
($M)
0.36
17.66
The operating costs estimate includes the cost of mining, processing, tailings storage and General
and Administration (“G&A”) services. The LOM average operating costs for the Project during
production years is summarized in Table 21.20.
TABLE 21.20
OPERATING COST SUMMARY
Area
Open Pit Mining
Underground Mining
Paste Backfill Plant
Subtotal Average LOM Mining
Processing
SART Plant
Dry Stack Tailings Facility
G&A
Total cost per tonne processed1
Cost
($/t processed)
18.46
44.04
5.89
42.92
33.36
4.13
2.13
6.88
89.43
Cost
($/t mined)
2.64
Note: 1 Totals may not sum due to rounding.
21.3.2
Underground Mining Operating Costs
The operating cost estimate addresses the costs of all: development (excluding sustaining CAPEX
development); production; backfilling; services and power usage; technical services auxiliary staff
roles; and other dayworks and sundry costs associated with operating the underground mine. The
underground portion of the Los Ricos South Project is expected to be mined with contractors: costs
for development, production and haulage were based on quotes from three contractors experienced
in Mexico. All costs used for underground development and production are sourced directly, or
are based on these costs.
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Total OPEX for the underground over the LOM is estimated at $330.8M, as shown in Table 21.21.
The average OPEX per UG tonne processed is $44.04/t excluding operation of the paste backfill
plant. No contingency has been applied to underground operating costs.
TABLE 21.21
UNDERGROUND MINING OPERATING COSTS
Area
Operating Development
Production
Power Consumption
Diesel Consumption
Other Operating Costs
Total
Cost
($M)
79.7
148.3
29.4
40.4
33.0
330.8
Cost per Tonne
($/t)1
10.61
19.75
3.91
5.38
4.39
44.04
Note: 1 Totals may not sum due to rounding.
21.3.2.1
Development
A total of 33.6 km of operating development is planned during the production years. Total OPEX
development cost over the LOM is estimated at $79.7M. Table 21.22 provides a summary of
development quantities, cost per heading size, and total costs incurred for the portion developed
after the start of production.
TABLE 21.22
UNDERGROUND OPERATING DEVELOPMENT COSTS
Heading Type
Heading
Profile
(m W x m H)
Ore Development
Waste Development
Development Through PF
4.0 x 4.0
Back and Wall Waste Slashing2
4.0 x 4.0
(equivalent)
Total (Pre-contingency)
Required
Development Total OPEX
Development
Cost
Cost
(m)
($/m)
($M)1
21,204
2,436
51.7
8,630
2,436
21.0
2,050
2,586
5.3
1,683
1,012
1.7
33,567
2,374
79.7
1
Note: Totals may not sum due to rounding.
2
Estimated volume of slashing is 26.9 k m3, equivalent to 1,683 m of 4.0 mW x 4.0 mH development.
21.3.2.2
Production
Production costs include all costs associated with the delineation, drilling, blasting, excavating,
and material handling of stopes in the underground mine. Total OPEX production cost over LOM
is estimated at $148.3M. Table 21.23 provides a summary of these costs.
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TABLE 21.23
UNDERGROUND OPERATING PRODUCTION COSTS
Area
Drill and Blast
Excavation
Haulage
Long Support
Delineation Drilling
Total
Cost
($M)1
67.7
20.5
51.0
2.2
7.0
148.3
Note: 1 Totals may not sum due to rounding.
21.3.2.3
Power and Diesel
Electrical power consumption during the operating life of the UG mine is estimated at $29.4M,
with ~60% of power consumption derived from ventilation fans (both primary and auxiliary).
Consumption could likely be reduced with a ventilation-on-demand (“VOD”) system, through the
use of larger-diameter ventilation raises, or through the reduction in the use of diesel equipment.
Diesel fuel consumption cost for the UG is estimated at $40.4M.
21.3.2.4
Other Operating Costs
Other operating costs include indirect salaries associated with the Owner’s technical services team,
G&A salaries associated with direct UG mine supervision and management, equipment rental,
general UG construction works, rehab and intersection support, and costs for dayworks and
sundries. Over the operating LOM, these OPEX costs are estimated at $33.0M.
21.3.3
Paste Backfill Plant Operating Costs
The paste plant operating cost estimate was based on a 91 t/h paste plant facility. Based on the FS
mine plan, the paste plant will operate for the first 14 years of production with annual run times
ranging between 1,448 and 5,102 hours. Table 21.24 summarizes the LOM operating costs
expected for the paste plant.
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TABLE 21.24
PASTE PLANT LOM OPERATING COST SUMMARY
Cost Area
Labour
General Maintenance
Mobile Equipment
Power
Reagents & Operating Consumables
Total Paste Plant Cost
LOM Total
$M
3.2
6.2
2.4
4.4
44.0
60.2
$/t
Processed
0.31
0.61
0.24
0.43
4.30
5.89
%
Total
5.3
10.3
4.0
7.3
73.1
100.0
The LOM paste plant operating costs were calculated based on the following assumptions and
inputs:
•
Annual paste plant throughput and run time hours were based on the annual FS mine
plan;
•
Process plant labour will be sourced locally;
•
Underground paste operating consumables, maintenance costs and labour costs are
excluded from the operating costs;
•
Reagent consumption rates have been estimated based on tailings paste testwork. A
quoted unit cement supply cost of $152/t, which includes delivery to site, was provided
by GoGold and used in the estimate;
•
Process mobile equipment costs include fuel and maintenance for the equipment.
Leasing costs are excluded from the operating costs (included as sustaining capital);
•
A unit power supply cost of $0.129/kWh, confirmed by CFE; and
•
A unit diesel supply cost of $1.11/L, confirmed by GoGold.
Paste plant power costs were calculated from the average power utilization in the electrical load
list for the equipment used together with the associated annual run time hours. The LOM average
electrical power consumption is estimated to be 2,614 MWh per year.
The LOM total and unit power cost is estimated at $4.4M and $0.43/t processed, respectively. The
power contributes 7.3% of the overall paste plant operating cost.
The paste concrete addition rates were estimated from paste backfill testwork with annual cement
tonnages nominated in the mine plan. The estimated LOM total and unit reagent cost was $44.0M,
and $4.30/t processed respectively. The paste plant reagent costs are the largest contributor to the
paste operating costs and account for 73.1% of the overall cost.
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The estimated LOM total and unit labour cost was $3.2M, and $0.31/t processed. A total of 12
personnel are required to support the paste plant operation, with each shift including one supervisor
and two operators.
Maintenance costs were derived by applying an average maintenance factor to the installed
mechanical equipment costs. The LOM total and unit operating maintenance costs were estimated
at $6.2M and $0.61/t processed, respectively. Overall, maintenance represents 10.3% of the total
paste plant operating cost.
Mobile equipment operating costs include the fuel and maintenance costs for vehicles and mobile
equipment required to support the paste operation. Fuel and maintenance costs have been
calculated based on the number of and estimated annual run hours of each type of equipment.
Leasing costs and downpayments are not included in the operating costs as these are covered in
the capital cost build-up.
The mobile equipment includes light vehicles, a skid steer for local maintenance and a FEL used
for reclaiming stockpiled tailings material to feed the paste plant. LOM costs for mobile equipment
are estimated to be $2.4M, at a LOM rate of $0.24/t processed.
21.3.4
Open Pit Mining
The Los Ricos South open pit mine operating cost consists of a contract mining cost, a stockpile
rehandling cost, and Owner’s costs for supervision and technical services.
The average LOM unit mining operating cost is estimated at $2.64/t, as shown in Table 21.25.
Open pit OPEX is the average over the pit mining production years 11 to 15 and does not include
costs from the pre-stripping period in year 10.
TABLE 21.25
OPEN PIT MINING OPERATING COST SUMMARY
21.3.4.1
Operating
Cost
LOM
($M)
Contract Mining
Technical Services
Other Owner’s
Total Mining Cost
45.2
2.3
0.7
48.2
Years 11 to 15
Unit Cost
Unit Cost
($/t material) ($/t processed)
2.47
17.31
0.12
0.87
0.04
0.28
2.64
18.46
Contract Mining
Several Mexican mining contractors were contacted to provide budgetary quotations. These
included all open pit mining costs, including equipment, personnel, diesel fuel, and maintenance.
The lowest quotation was selected and was based on a mining unit rate of $2.02/t, with an overhaul
haulage cost of $0.25/t.km. A mining cost model applies the haulage rate to the varying haul
distances appropriate for each year to determine annual costs. The unit mining rate does not include
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explosive or blasting services, which are an extra cost. Table 21.26 is a breakdown of the mining
contractor unit rates.
The explosive supply cost, also shown in Table 21.26, includes three components. The explosive
contractor will supply ammonium nitrate (“AN”), the Owner will provide diesel fuel to produce
the ANFO, and the contractor will provide the blasting services at a fixed cost of $28,400 per
month. This monthly fixed cost includes personnel and equipment necessary to mix blasting
agents, charge holes, and blast. The powder factor estimated by the blasting contractor is 0.21
kg/tonne for both ore and waste.
TABLE 21.26
OPEN PIT CONTRACTOR UNIT RATES
Item
Drilling
Loading
Hauling 1 km
Services
Subtotal
Overhaul $/km
Ammonia Nitrate
Diesel Fuel
Blasting Service $/month
21.3.4.2
$/tonne Material
0.39
0.38
0.71
0.54
2.02
0.25
0.231
0.013
28,400
Supervision and Technical Services
The mine supervision and technical services includes salaries for the 13-person Owner’s team, and
annual costs for office equipment, vehicle maintenance, outside consultants, and other mine
management costs. This annual fixed Owner’s cost is estimated to be $674,000 per year shown by
“Technical Services” and “Other Owner’s” in Table 21.25 above.
21.3.5
Process Plant Operating Costs
21.3.5.1
Basis of Estimate
Common to all process plant operating cost estimates are the following assumptions:
•
Cost estimates are based on Q4 2024 pricing without allowances for inflation;
•
Costs are expressed in United States dollars (USD or US$);
•
For material sources in Canadian dollars, an exchange rate of 1.28 Canadian dollar per
US dollar was assumed;
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•
For material sourced in Mexican pesos, an exchange rate of 20.28 pesos per US dollar
was assumed;
•
Most (non-expatriate) of the labour requirement is assumed to come from neighbouring
municipalities;
•
Processing unit operations were developed from first principles and benchmarked
against similar or comparable processing plants to ensure their relative accuracy;
•
Equipment and materials will be purchased as new;
•
Grinding media consumption rates have been estimated based on the measured ore
characteristics, benchmark data and vendor supplied information;
•
Reagent consumption rates have been estimated based on metallurgical testwork and
standard operating practices, METSIM simulations developed for the Los Ricos
flowsheet and Ausenco in-house data and experience;
•
The process mobile equipment costs include fuel, maintenance for the equipment.
Leasing costs are excluded from the operating costs (included as sustaining capital);
•
Doré transportation, treatment, refining, and other related costs are not included in this
estimate;
•
SART plant operating costs or SART concentrate transportation, treatment, refining,
and other related costs are not included in this estimate;
•
Annual milled capacities and feed grades were based on the annual FS mine plan;
•
A crushing plant, milling and filter plant availability of 70%, 92% and 85%,
respectively;
•
A unit power supply cost of $0.129/kWh, confirmed by Comisión Federal de
Electricidad (“CFE”), Mexico’s state-owned, electrical utility company;
•
A unit diesel supply cost of $1.11/L, provided by GoGold as a quote from a Mexican
fuel distribution firm; and
•
Surface water licensing cost of $1.33/m3, provided by GoGold.
21.3.5.2
Process Operating Costs
The process operating cost estimate was based on a 2,000 tpd process plant comprising crushing,
grinding, cyanide leaching, countercurrent decantation (“CCD”) washing, Merrill Crowe precious
metals recovery, and smelting unit operations to produce gold–silver doré bars. Sulphidization,
acidification, recycling and thickening (“SART”) and tailings cyanide detoxification and
dewatering costs were also included.
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The LOM total and unit process operating cost was estimated at $341.4M and $33.36/t processed,
respectively. Table 21.27 summarizes the LOM operating costs expected for the process plant.
TABLE 21.27
LOM PROCESSING OPERATING COST SUMMARY
Cost Area
Labour
General Maintenance
Mobile Equipment
Lab
General
Power
Reagents & Operating Consumables
Total Processing Cost
Total
($M)
33.1
18.6
12.1
9.1
1.5
84.3
182.7
341.4
$/t
Processed
3.24
1.82
1.18
0.89
0.15
8.24
17.85
33.36
% of
Total
9.7
5.4
3.5
2.7
0.4
24.7
53.5
100.0
Reagents and operating consumables and power are the two highest cost areas of the processing
operating cost, with each area respectively contributing 53.5% and 24.7% of the total processing
cost.
Reagents and Operating Consumables
Individual reagent and consumable consumption rates were estimated from metallurgical testwork
results, METSIM simulations and Ausenco’s in-house database and experience. Reagent unit
costs, delivered to Guadalajara, were provided by GoGold based off existing supplier agreements
or were directly quoted from suppliers. Delivery costs from Guadalajara to site were quoted and
added to the reagent unit costs.
Surface water licensing fees required for water supplied from the Los Huajes and La Labor Dams
were included, at a LOM nominal unit cost of $0.26/t processed.
The LOM total and unit cost of reagents was estimated at $133.7M and $13.06/t processed, which
accounts for 39.1% of the overall process operating costs. Sodium cyanide, SMBS and zinc powder
contribute 69% of the reagent costs.
The operating consumables contribute a total of $49.0M to the LOM operating costs, at $4.79/t
processed. The SAG Mill liner replacements and grinding media contribute 82% of the total
operating consumables cost. The operating consumable wear and consumption rates were
estimated based on the average measured ore characteristics, benchmark data and vendor supplied
information. Operating consumable unit supply rates were quoted by selected vendors and
benchmarked against recent projects.
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Page 475 of 577
Power
The power costs of the processing facilities were calculated from the average power utilization in
the electrical load list for the equipment used and the associated annual run time hours. The
electrical power consumption is estimated to be 47,015 MWh per year or approximately 65.3
kWh/t processed.
Electricity will be provided to the site by CFE at a unit cost of USD$ 0.129/kWh. Maintenance of
the incoming high voltage (“HV”) power line will be the responsibility of CFE.
The LOM total and unit power cost is estimated at $84.3M and $8.24/t processed, respectively.
Power costs contribute 24.9% of the overall operating cost. Grinding represents approximately
48% of the total power demand.
Labour
The estimated LOM total and unit labour cost was $33.1M, and $3.24/t processed, respectively. It
is based on fully loaded Q3 2024 labour rates provided by GoGold, which reflect their existing
regional operations. A total of 134 people is required for the mill operation, maintenance,
laboratory and management, as presented in Table 21.28.
Staff employees will work on site for five days a week, from Monday to Friday on a 5/2 schedule.
A management staff member will work on site through the weekend. The operations and
maintenance crew will work on a shift that is four days on-site, four days off site as labour is local.
The sum of 134 employees can be broken down into the following labour roles:
•
•
•
69 process plant operations employees;
50 maintenance employees including mechanics, inspectors, welders and electricians;
and
15 laboratory employees.
TABLE 21.28
HOURLY LABOUR REQUIREMENTS AND
ANNUAL SALARIES
Position
Employees
Annual
Salary
($/a)
Operations - Processing
Process Manager
Production Superintendent
Senior Metallurgist
Metallurgist
Process Technician
1
1
1
1
1
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132,704
68,831
68,831
35,091
35,091
Page 476 of 577
TABLE 21.28
HOURLY LABOUR REQUIREMENTS AND
ANNUAL SALARIES
Position
Employees
Annual
Salary
($/a)
Shift Supervisor
ROM FEL Operator
Crusher Operator
Control Room Operator
Grinding Operator
Leach/CCD Operator
Merrill Crowe Operator
Refinery/Gold Room Operator
Refinery/Gold Room Helper
Mobile Equipment Operator
Tailings Filtration Operator
Tailings
Management
FEL
Operator
General Labourers/Helper
4
4
4
4
4
4
4
4
8
4
4
29,522
13,951
13,951
13,951
13,951
13,951
13,951
13,951
9,848
13,951
13,951
4
13,951
12
9,848
Operations – Processing Total
69
1,213,636
Operations – Maintenance
Maintenance Superintendent
Maintenance Supervisor
Maintenance Planer
Specialized Mechanic
Lube Inspector
Welders
Crane Operator
Helper Mechanic
Tool Crib Helper
General Electricians
Instrument Technicians
1
4
1
8
4
4
4
8
4
8
4
68,831
29,522
28,136
14,614
14,614
14,614
14,614
12,340
9,848
14,614
14,614
Operations – Maintenance Total
50
820,815
Operations – Lab
Lab Supervisor
Lab Analyst
Lab Technician
1
2
12
29,522
27,901
9,641
Operations – Lab Total
15
201,016
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Page 477 of 577
General Maintenance
The process plant maintenance costs were derived by applying an average maintenance factor to
the installed mechanical equipment costs. The factors were determined by benchmark maintenance
costs derived from recently completed projects.
The LOM total and unit operating maintenance costs were estimated at $18.6M and $1.82/t
processed, respectively. Overall, maintenance represents 5.4% of the total processing operating
cost.
Mobile Equipment
Processing mobile equipment operating costs include the fuel and maintenance costs for vehicles
and mobile equipment required to support the plant operation and maintenance. Mobile equipment
includes light passenger vehicles, maintenance equipment and heavy equipment used in and
around the process plant areas.
Fuel and maintenance costs have been calculated based on the number of and estimated annual run
hours of each type of equipment. Leasing costs and downpayments are not included in the
operating costs as these are covered in the capital cost build-up.
The mobile equipment costs include the front-end loader costs required to:
•
Reclaim stockpiled ROM ore to feed the ROM bin;
•
Reclaim stockpiled crushed ore to feed the emergency hopper and maintain feed to the
SAG mill when the primary crusher is down; and
•
Reclaim filtered tailings from the tailings bunker and transfer it to haul trucks which
transport the tailings to either the dry stack or paste plant facility.
The LOM cost for mobile equipment is estimated to be $12.1M, at a LOM unit rate of $1.18/t
processed.
Metallurgical Laboratory
The operating cost estimate for the laboratory was benchmarked against existing operations of
similar capacity and flowsheet complexity. The cost estimate covers the required process analysis
of routine plant samples. Labour costs for the laboratory operation are captured in the labour cost
estimate.
The LOM total and average unit operating cost of the assay laboratory was estimated at $9.1M and
$0.89/t processed, respectively.
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General
Costs for mercury disposal and treatment were estimated at $1.5M over the LOM and average
$0.15/t processed.
21.3.6
SART Plant
Annual SART plant OPEX for reagents, power and labour is estimated at $2.97M or $4.13/t
processed. SART operating costs to produce a copper SART (Cu2S) saleable concentrate are
summarized in Table 21.29.
TABLE 21.29
SART PLANT OPERATING COSTS
Item
Labour
Power & Fuel
Operating Consumables – Reagent
Operating Allowance
Total
$M/Year
0.32
0.13
2.43
0.09
2.97
$/t Processed
0.44
0.18
3.38
0.13
4.13
% of Total
11
5
81
3
100
Source: BQE Water (2025) and DENM (2025)
Labour positions and rates for the SART plant are based on manpower rates at the GoGold Parral
operation in Chuhuahua, Mexico. Total SART work force is 15 including plant supervisors and
plant operators. A 37% burden rate is applied to each position based on information supplied by
Human Resource at GoGold’s wholly owed Mexican subsidiary, Grupo Coanzamex.
Reagent cost estimates are based on quotations provided by Grupo Coanzamex operations and
include sulphuric acid (H2SO4), sodium hydrosulphide (NaHS), anionic flocculant, caustic soda,
and hydrogen peroxide. Consumption calculations are based on the recent Project testwork
completed on mineralization from the Project’s Abra and Eagle Zones. Consumption rates are
calculated on an annual basis at a 2,000 tpd process plant feed rate.
Based on the feasibility level flowsheet and equipment list, electricity consumption for the SART
plant is estimated to be 1,566 MWh per year.
21.3.7
Dry Stack Tailings Facility Operating Costs
The DSTF operating costs were calculated on an annual basis for the LOM and considered the use
of the following mobile equipment:
•
One CAT D7 Dozer or equivalent;
•
One CAT CP56 Padfoot Vibratory Soil Compactor; and
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•
One 28 t Articulated Haul Truck for year one (Yr1) and an additional 28 t Articulated
Haul Truck for the remainder of the LOM.
The total LOM and unit DSTF operating costs were estimated to be $21.8M and $2.13/t processed,
respectively.
21.3.8
General and Administrative Costs
General and Administration (“G&A”) costs are estimated at $4.9M annually, and average $6.88/t
processed over the LOM, as summarized in Table 21.30. The costs are based on GoGold costs
incurred at its Parrall operation and actual local payments such as mineral tenure fees.
TABLE 21.30
GENERAL AND ADMINISTRATION COSTS
Item
Number
General Manager
Administrator
Finance
Environmental & Community
Health & Safety
Security
Human Resources
Purchasing & Logistics
Surface & Infrastructure
Health & Safety, Medical
Surface Support & Maintenance
Environmental & Bonding
HR & Community Relations
Insurance
Legal & Mineral Tenure
Office, Consultants, IT & Other
Total
1
1
3
8
10
12
3
10
9
Lump sum
Lump sum
Lump sum
Lump sum
Lump sum
Lump sum
Lump sum
Annual Cost
($M)1
0.26
0.06
0.14
0.18
0.38
0.15
0.09
0.23
0.15
0.16
0.68
0.41
0.44
0.35
1.07
0.17
4.92
Note: 1 Totals may not sum due to rounding.
21.4
MINING TAX
The Project is subject to a 0.5% NSR mining tax payable to the Mexican government. Total costs
associated with this NSR royalty tax over the LOM are estimated at $10.5M.
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21.5
CLOSURE COSTS
Costs to be incurred after the LOM plan were estimated by CIMA and reviewed by the Authors to
be $3.7M to close and rehabilitate the Project site, as presented in Table 21.31.
TABLE 21.31
CLOSURE COSTS
Activity
Road Rehabilitation
Buildings
Offices
Waste Rock & Tailings Storage
Open Pit Rehabilitation
Industrial Areas
Ecological Areas
Total
21.6
Closure Cost
($M)
0.49
0.05
0.03
0.83
1.10
1.12
0.08
3.70
CASH COSTS AND ALL-IN SUSTAINING COSTS
Cash costs over the LOM, including Mexican mining taxes, are estimated to average US$11.59/oz
AgEq. All-In Sustaining Costs (“AISC”) over the LOM are estimated to average US$12.78/oz
AgEq and include closure costs.
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22.0
ECONOMIC ANALYSIS
22.1
SUMMARY
The results of the economic analyses discussed in this section represent forward-looking
information as defined under Canadian securities law. The results depend on inputs that are subject
to a number of known and unknown risks, uncertainties and other factors that may cause actual
results to differ materially from those presented here. Information that is forward-looking includes:
•
•
•
•
•
•
•
•
Mineral Reserve Estimates;
Commodity prices and exchange rates;
Mine production plans;
Mining and process recovery rates;
Mining dilution and mining recoveries;
Capital, sustaining costs and operating costs;
Closure costs and closure requirements; and
Environmental, permitting and social risks.
Additional risks to the forward-looking information include:
•
Changes to costs of production differing from what is assumed;
•
Unexpected variations in quantity of mineralized material, grade or recovery rates;
•
Geotechnical or hydrogeological considerations during mining being different from
what was assumed;
•
Failure of mining methods to operate as anticipated;
•
Failure of plant, equipment or processes to operate as anticipated;
•
Changes to assumptions as to the availability of electrical power, and the power rates
used;
•
in the operating cost estimates and financial analysis;
•
Ability to obtain permits to operate;
•
Unrecognized environmental risks;
•
Unanticipated reclamation expenses;
•
Accidents, labour disputes and other risks of the mining industry; and
•
Changes to tax rates.
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A discounted cash flow model was prepared using the production schedule described in Section
16 and the cost estimates described in Section 21 of this Technical Report. The cash flow model
was developed on a pre-tax and after-tax basis. The cash flow model is assumed to commence
from the time a production decision is made and does not include time or costs for further detailed
engineering studies.
The Los Ricos South Project economic evaluation conclusions are summarized in Table 22.1. At
base case metal prices of US$2,330/oz Au, US$26.80/oz Ag and US$4.00/lb Cu, the Project has
an estimated US$355M after-tax net present value (“NPV”) at a 5% discount rate (“NPV5%”),
and an after-tax internal rate of return (“IRR”) of 28%. The payback period is estimated to be 2.6
years from start of production from underground mining.
TABLE 22.1
ECONOMIC EVALUATION SUMMARY
Item
NPV0% ($M)
NPV5% ($M)
NPV7% ($M)
IRR (%)
Payback period (years)
22.2
Pre-Tax
851
553
468
39
1.8
After-Tax
575
355
293
28
2.6
BASIC ASSUMPTIONS
A discounted cash flow analysis of the Los Ricos South was prepared based on technical and cost
inputs developed by the Authors.
The discounted cash flow analysis was performed on a stand-alone project basis with annual cash
flows discounted. The financial evaluation uses a discount rate of 5%, discounting back to the
commencement of construction (Year -2) of the Project.
All currency values are expressed in US dollars unless otherwise noted.
Metal Price Assumptions
All metal prices remain constant throughout the LOM of the Project.
Ag: US$26.80/oz. The sensitivity of the Project return to variations in the actual silver price
received was also examined.
Au: US$2,330/oz. The sensitivity of the Project return to variations in the actual gold price
received was also examined.
Cu: US$4.00/lb.
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Metallurgical Recoveries
The Los Ricos South Project’s process plant metallurgical recovery assumptions are:
Ag:
Au:
Cu:
86% average over LOM;
93% average over LOM; and
68% average combined leach and SART over LOM.
Capital Costs
Total capital costs during the LOM are estimated to be $326.4M as outlined in the Capital and
Operating Cost Section 21. The initial capital costs are incurred over a two-year construction
period and are estimated to be $226.7M. Sustaining capital costs including open pit mining are
estimated at $99.7M.
Previous Expenses Provision
An amount of $7.0M was considered as a prior expense pool and these monies were deducted from
income in Year 1 when determining taxable income.
Income Tax Rate
The Mexican income tax is levied at a rate of 35% on the net taxable income.
Additional Mining Tax Rate
A Mexican government 0.5% gross revenue tax was applied in the cash flow economics and is
deductible when determining taxable income.
22.3
CASH FLOW SUMMARY
Based on the assumed metal prices the Project has an after-tax IRR of 28% and a 2.6 year payback
of initial pre-production capital costs for underground mining. The Project is estimated to realize
an after-tax NPV of US$355M at a discount rate of 5%.
The estimated annual production and LOM cash flows for the Los Ricos South Project are
summarized in Table 22.2 and annual cash flows are presented in Figure 22.1. An annual summary
of the cash flow model is provided in Table 22.5.
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TABLE 22.2
PROJECT CASH FLOW SUMMARY
Item
UG Ore Mined
OP Ore Mined
Unit
kt
kt
Value
7,512
2,721
Process Plant Ore
kt
10,233
Silver Grade
g/t
145
Gold Grade
g/t
1.39
Copper Grade
AgEq Grade
%
g/t
0.10
276
Item
Net Revenue
Initial Capital Cost
Sustaining Capital
Cost
Mining Operating
Cost
Process (Incl. SART
& DSTF) Operating
Cost
G&A Operating Cost
LOM Operating Cost
Silver Recovery
%
86
Operating Cash Cost
Gold Recovery
%
93
Copper Recovery
%
68
Silver Price
US$/oz
26.80
Gold Price
Copper Price
Payable Silver Metal
Payable Gold Metal
US$/oz
US$/lb
Moz
koz
2,330
4.00
41.1
423.6
Payable Copper
Mlb
11.2
Payable AgEq
Moz
79.9
All-in Sustaining
Cost
Mine Life
Average Process
Plant Rate
Pre-Tax NPV (5%)
After-Tax NPV (5%)
Pre-Tax IRR
After-Tax IRR
Pre-Tax Payback
Period
After-Tax Payback
Period
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Unit
US$M
US$M
Value
2,099
227
US$M
100
$/t ore
42.92
$/t ore
39.63
$/t ore
$/t ore
US$/oz
AgEq
US$/oz
AgEq
Yrs
6.88
89.43
tpd
2,000
US$M
US$M
%
%
553
355
39
28
Years
1.8
Years
2.6
11.59
12.78
15
Page 485 of 577
FIGURE 22.1
22.4
ANNUAL CASH FLOW PROFILE
SENSITIVITIES
The Los Ricos South Project economics were examined with a sensitivity analysis for several key
variables. The results of the sensitivity analyses on the after-tax NPV with a 5% discount rate are
shown in Tables 22.3 and 22.4.
TABLE 22.3
SILVER AND GOLD PRICE SENSITIVITY NPV, IRR AND PAYBACK
Sensitivity
-25%
-18%
-10%
Silver Price (US$/oz)
Gold Price (US$/oz)
After-Tax NPV (5%) (US$M)
After-Tax IRR (%)
After-Tax Payback (years)
20
1,739
110
13.6
4.7
22
1,913
184
18.3
3.8
24
2,087
255
22.5
3.2
Base
Case
26.8
2,330
355
28.0
2.6
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+12%
+23%
+34%
30
2,608
469
33.7
2.0
33
2,869
575
38.8
1.8
36
3,130
681
43.7
1.7
Page 486 of 577
TABLE 22.4
CAPITAL AND OPERATING COST SENSITIVITY OF NPV AND IRR
Sensitivity
-20%
-10%
Operating Costs – NPV (5%) (US$M)
Operating Costs – IRR (%)
Capital Costs – NPV (5%) (US$M)
Capital Costs – IRR (%)
440
32.6
394
35.1
399
30.6
373
30.9
Base
Case
355
28.0
355
28.0
+10%
+20%
317
26.3
331
24.5
275
24.1
310
22.0
The after-tax base case NPV and IRR are most sensitive to metal prices followed by operating
costs and capital costs.
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TABLE 22.5
CASH FLOW MODEL SUMMARY
Item
Payable Ag
Payable Au
Payable Cu
Payable AgEq
Revenue
(-) Operating
Cost
(-) Mining Tax
(-) Capital
Spending
(+) Salvage
Value
(-) Closure
Pre-Tax Cash
Flow
(-) Income Tax
After-Tax
Cash Flow
Cumul AfterTax Cash Flow
Disc After-Tax
Cash Flow
(5%)
Disc Cumul
After-Tax Cash
Flow
Units Totals
-1
-
1
3,027
33
0.9
6,027
158.3
2
3
5,315 2,794
58
38
1.5
1.2
10,637 6,351
279.5 166.9
4
3,289
41
1.6
7,161
188.1
5
3,006
37
1.2
6,450
169.5
Year
6
7
3,118 3,156
34
30
0.9
1.0
6,251 5,944
164.2 156.2
koz
koz
Mlb
koz
$M
41,078
424
11.2
79,889
2,098.9
-2
-
8
3,232
34
1.3
6,445
169.3
9
2,739
26
0.7
5,114
134.4
10
2,556
23
0.4
4,672
122.7
11
2,065
18
0.1
3,686
96.9
12
13
14
2,306 1,989 2,001
14
16
15
0.1
0.1
0.2
3,525 3,437 3,322
92.6 90.3 87.3
15
485
4
0.1
865
22.7
$M
(915.2)
-
-
(57.4)
(73.4)
(73.6) (76.4) (78.6) (71.4) (67.7) (72.1) (63.3) (58.4) (57.2) (51.8) (49.3) (43.2) (21.6)
$M
(10.5)
-
-
(0.8)
(1.4)
(0.8)
(0.9)
(0.8)
(0.8)
(0.8)
(0.8)
(0.7)
(0.6)
(0.5)
(0.5)
(0.5)
(0.4)
(0.1)
$M
(326.4) (89.7) (137.0)
(11.2)
(19.1)
(9.9)
(9.3)
(6.1)
(5.5)
(8.0)
(0.6)
(13.5) (15.5)
(0.1)
(0.2)
-
-
(0.8)
$M
8.0
-
-
-
-
-
-
-
-
-
-
-
-
-
-
-
-
8.0
$M
(3.7)
-
-
-
-
-
-
-
-
-
-
-
-
-
-
-
-
(3.7)
$M
851.1
89.0
185.6
82.6
101.6
83.9
86.5
79.7
95.8
56.8
48.2
39.0
40.2
40.6
43.7
14.2
$M
(276.3)
(19.1)
(55.1)
(19.3) (25.9) (19.7) (20.7) (18.3) (23.9) (10.3)
(7.3)
(13.5) (13.9) (14.0) (15.1)
(0.1)
$M
574.9
69.9
130.5
63.3
75.7
64.2
65.9
61.4
71.9
46.5
40.9
25.5
26.3
14.1
$M
-
(89.7) (236.2) (166.3) (35.8)
27.5
103.2
167.3
233.2
294.6
366.5
413.0
453.9
479.4
505.7 532.2 560.8 574.9
$M
355.1
(89.7) (139.6)
112.7
52.1
59.3
47.9
46.8
41.5
46.3
28.5
23.9
14.2
13.9
$M
-
(89.7) (229.3) (165.8) (53.1)
(1.0)
58.3
106.2
153.0
194.5
240.9
269.4
293.3
307.5
321.5 334.9 348.6 355.1
(89.7) (146.6)
-
-
(89.7) (146.6)
63.4
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26.5
13.4
28.5
13.7
6.5
23.0
ADJACENT PROPERTIES
The Hostotipaquillo Region hosts many low-sulphidation epithermal precious metal prospects
and deposits.
The Santo Domingo Silver-Gold Deposit of Stroud Resources Ltd. is located approximately 10
km to the north of Los Ricos South. Stroud has owned and explored the Santo Domingo
Property since 1989 and completed five drilling programs between 1999 and 2012 that totalled
45 diamond drill holes (McBride, 2017). The Santo Domingo Deposit was exploited in the
early seventeenth century as part of the San Pedro Analco mining area. McBride (2017)
estimates a Measured and Indicated Mineral Resource to be 6.01 Mt averaging 0.47 g/t gold
and 101 g/t silver containing 25.7 Moz of AgEq and an Inferred Mineral Resource of 3.48 Mt
containing 13.4 Moz of AgEq.
The Authors note that mineralization located on adjacent properties is not indicative of
mineralization present on the Los Ricos South Property.
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24.0
OTHER RELEVANT DATA AND INFORMATION
24.1
PROJECT EXECUTION PLAN
The Project Execution Plan (“PEP”) addresses the overall Project objectives, scope, strategies
and roles and responsibilities, and provides a comprehensive plan for its development and
implementation. The PEP covers the plan for engineering, procurement, construction, start-up,
and commissioning of the Project. The implementation strategy assumes an engineering,
procurement, construction management (“EPCM”) implementation with construction packages
and several smaller EPC (engineer, procure, construct) packages where either a local contractor
or specialist technology supplier has demonstrated cost benefits to the Project. The execution
strategy is to complete the Project construction and commissioning within 24 months of
receiving the environmental permits, securing funding for construction and making the final
investment decision (“FID”). A preliminary EPCM schedule has been developed and will be
updated during the next Project phase.
24.2
PROJECT ORGANIZATION AND ALIGNMENT STRATEGY
The Project organization is based on an integrated team approach, minimizing the duplication
of roles and activities between EPCM firm and the Owner’s team (Figure 24.1). Full-service
delivery of the Project will be enabled from the Project’s site office.
The government liaison activities, external affairs, and environmental services, and overall site
security will be provided by GoGold.
The alignment strategy aims to create shared understanding of the Project vision and strategy
to enable the EPCM firm, GoGold, and other internal/external stakeholders to achieve the
Project objectives. The intent is for the overall Project delivery team to operate as one team
with defined responsibilities, accountabilities, and authorities. The team is established and
supported to deliver “Best for Project” outcomes in line with GoGold’s expectations and
critical success factors.
Facilitated alignment sessions will be held to establish working relationships within the Project
delivery team and determine acceptable outcomes. The alignment effort will be concentrated
at the start of the Project, although ongoing activities will be carried out throughout Project
execution to increase overall effectiveness and commitment, and to ensure the Project team
functions cohesively.
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FIGURE 24.1
PROJECT ORGANIZATION CHART
Source: Ausenco (2025)
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24.3
HEALTH, SAFETY, ENVIRONMENT AND COMMUNITY MANAGEMENT
PLAN
The Health, Safety, Environment, Community (“HSEC”) Management Plan will be developed to
detail the strategy, systems, process, procedures, and standards that are used for recording,
analyzing, and reporting performance. HSEC standards, systems and safe operating procedures
developed by GoGold will be supplemented by the EPCM firm’s procedures and forms. Project
“on-boarding” requirements are substantial and will be upheld in accordance with established
GoGold programs. The EPCM firm will support and assist GoGold in developing safe operating
procedures for Los Ricos South.
All legislative and contractual requirements will be identified during the planning phase, including
the identification of specific Acts, Regulations, Standards, Codes of Practice and any contractual
requirements that must be complied with in relation to the work activities or risk profile. The
relevant legislative and contractual requirements will be integrated into this management plan and
the supporting procedures.
The project manager is accountable for overall delivery of the Project and for ensuring that
management structures and systems are embedded within the Project. The construction manager
will be supported by a site-based HSEC team who will monitor and advise on HSEC performance
to ensure alignment between Project team members and GoGold’s requirements.
24.4
ENGINEERING MANAGEMENT PLAN
The engineering team will deliver basic and detailed designs for the scope of facilities on the Los
Ricos South Project and provide deliverables and technical support to ensure the effective
procurement, construction, commissioning, and operation of the facilities. An Engineering
Management Plan (“EMP”) will be developed in the next phase to detail the strategy, processes,
and standards for delivering the engineering requirements of the Project.
The EMP is the guiding document defining the engineering and design objectives and strategies
and how the Project team will fulfil the engineering requirements for the design activities on the
Project. The EMP addresses the following key items:
•
Design criteria for the overall Project and applicable Engineering disciplines;
•
Systems and procedures to manage the Engineering and Drafting design process;
•
Baseline schedule for the Engineering phase;
•
Baseline Project scope – physical quantities and technical attributes; and
•
For items like the engineering team roles and responsibilities, engineering deliverables,
project scope of work, engineering scope of services and battery limits, and work
breakdown structure (“WBS”) refer to the PEP.
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When required, the engineering team will provide technical advice to the procurement,
construction management, commissioning, and quality departments throughout the Project.
Resources for ongoing field support during construction and commissioning phases will be drawn
from the engineering teams.
The EMP is intended to be reviewed and revised at regular intervals to ensure it is consistent with
Project phase objectives and strategies.
24.5
PROCUREMENT AND CONTRACTS STRATEGY AND MANAGEMENT PLAN
The Procurement Strategy and Management Plan will be developed in the next phase of work and
detail the strategy for procuring equipment and materials for the Project, in addition to the
processes and systems that will be used for procuring, expediting, shipping, receiving, storing, and
issuing the equipment and materials.
The Contracting Strategy and Management Plan will be developed in the next phase of work and
detail the strategy for identifying and engaging construction contractors for the Project and the
processes and systems that will be used to deliver the Project. It shall also explain the management
basis of the contracts formation and contracts administration procedure that will be applied along
with an explanation of tools, operational controls and earned value performance monitoring
techniques.
The procurement and contracting team will invite tenders and award purchase orders and contracts
according to packages defined by the Contract Strategy and Management Plan supported by
engineering and by GoGold’s approved bidders list.
24.6
PROJECT CONTROLS AND REPORTING PLAN
A Project Controls and Reporting Plan well be developed in the next phase of work to define the
project controls procedures and systems in support of the PEP. It serves to describe the strategies,
systems, processes, and responsibilities to achieve the following objectives: effective and efficient
project controls support for project execution; early identification of change; provision of accurate
and timely cost, schedule, and progress data together with analysis of trends and deviations;
provision of data and analysis necessary to facilitate informed decision making by Stakeholders;
and auditable record keeping and data storage for future estimating and benchmarking.
The scope of this plan covers following areas of the Project Controls function:
•
Cost and Change Management;
•
Planning and Scheduling;
•
Progress and Performance Monitoring;
•
Project Reporting; and
•
System Interfaces.
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This plan aims to satisfy the project analysis and reporting needs of the Project Management team,
EPCM Firm Corporate teams, and GoGold, including the overall Project consolidated reporting as
required by GoGold.
24.7
CONTRACTOR TEMPORARY FACILITIES
Laydown yards will be identified and nominated ahead of contractor mobilization to site.
Contractors will be required to establish their own temporary facilities under supervision from the
EPCM firm. Contractors will establish their own offices for the duration of their contract only;
these facilities will be removed on completion of the contractor’s works. Upon completion of their
specific construction activities, each contractor will demobilize all temporary buildings and
facilities and clean up the area allocated to them.
24.8
COMMISSIONING MANAGEMENT PLAN
Commissioning involves the formal handover and acceptance of process equipment and
commissioning modules between the various commissioning stages, from the completion of
installation by contractors and suppliers through verification of plant and equipment dry or precommissioning by field engineers, vendor representatives and design engineers, to final
commissioning by the commissioning team. A Commissioning Management Plan will be
developed and will detail the methodology and processes that will be used to plan, prepare, and
commission the works, including the involvement of the construction and operations (Owner’s)
teams.
The Commissioning Management Plan will also detail the Start-up/ Ramp up period. The Owner’s
operational team will drive the successful start-up with ore and ramp up period to bring the Project
to the nameplate throughput. The EPCM firm will act in a supporting role to assist during this
period.
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25.0
INTERPRETATION AND CONCLUSIONS
25.1
RISKS
25.1.1
Risk Workshop
A risk assessment workshop was carried out on January 17, 2025, with personnel from P&E,
Ausenco and GoGold in attendance. Overall Project risks are perceived by the Authors as
moderate. Eight high risks items were identified, and these are summarized in Table 25.1.
TABLE 25.1
PROJECT RISKS
Area
Project Constructability
Project Constructability
Open Pit Mining
Dry Stack Tailings Storage
Dry Stack Tailings Contact
Water
Ultimate Size of Dry Stack
Tailings Facility
Historical Underground
Workings
Risk
Fabricators may misrepresent
their capacity and capability
to deliver on major
construction contracts.
Comment/Mitigation Plan
Thorough financial
prequalification should be
conducted prior to award.
Monitor project development
of adjacent projects,
contractor engagement ahead
of a financing decision.
Permitting delay and
Federal law under review for
regulations risk.
open pit mines.
Update in next phase of
engineering. Can place gravel
Stability failure of facility.
on truck routes across stack.
More likely to slump instead
Instrumentation installed to
of debris flow.
monitor movement of stack
for proactive remedial action.
Ability to request one-off
permits to treat/release water.
Include mechanism to
Discharge from DSTF contact
evaporate water from ponds.
water is not permitted, should
Storage for contact water
be zero. In a storm event
during storm event is required
discharge may be exceeded.
in the case the permit is not
granted or granted in an
untimely manner.
Visible from nearby
Land acquisition, permit
communities, final size
amendment for size during
exceeds permitted area.
operations.
Personnel entry into old
Physical barriers, strong
workings without adequate
Health and Safety planning
risk analysis and
and culture.
rehabilitation.
Limited availability of
construction workforce,
skilled trades.
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TABLE 25.1
PROJECT RISKS
Area
Underground Open Stopes
Risk
Fall of rock hazard.
Comment/Mitigation Plan
Physical barriers, strong
Health and Safety planning
and culture, mining
sequencing where possible to
avoid people movement
across open areas, use remote
equipment.
Other Project risks noted by the Authors are set out below.
25.1.2
Metallurgical Testing
Metallurgical testing risks identified through the FS and metallurgical testwork program include:
•
Testwork grades for Eagle and Abra Deposit gold, silver and copper included those
reported in the mine plan, but the testwork gold and silver grades for the CC Deposit
were higher than those in the mine plan. It is assumed that the metallurgical
performance from the testwork is representative of the lower values in the mine plan;
•
There is an inherent risk that the samples tested do not suitably represent the Deposit,
however, the quantities of samples and spatial coverage suggest that this risk is low.
The samples were selected at the start of the FS on a preliminary mine plan which has
subsequently been revised. Variability testing on material representative of year 1 to 5
of mill feed should be repeated in an FS/detailed design bridging phase;
•
The CC open pit Deposit is introduced into the mill feed in year 10 and represents
approximately 20% of the open pit and 5% of the overall LOM mill feed. There has
been limited testwork completed on the Deposit to date with further testwork required
to characterize the Deposit;
•
Eagle variability sample CN-50 had gold and silver head grades within the expected
mine plan range but exhibited poor gold (76% to 79%) and minimal silver extraction
(2% to 3%). Elemental analysis of the sample indicated total Cu to be 2.25x and 3.24x
the average of the Eagle samples. The geographic areas within the Eagle Deposit which
the sample represents will not be mined until year 6, however, further investigation is
required to determine how much of the Mineral Resource this sample potentially
represents;
•
Elevated copper levels in the plant feed have the potential to affect precious metals zinc
precipitation, which has not been seen the testwork to date. Excess copper can also be
precipitated to the Merrill Crowe precipitate affecting doré quality and payabilities. To
mitigate this risk, the plant design incorporates an acid digestion circuit in the gold
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room to dissolve copper and zinc in the precipitate and a SART plant will remove
copper to produce a sulphide saleable copper concentrate;
•
No acid digestion circuit metallurgical testwork has been performed on the Merrill
Crowe precipitate to confirm the optimal conditions required to meet the copper and
zinc removal target of 90%;
•
The 2024 leaching testwork tests were performed at a solids density of 40% w/w solids,
which is lower than the FS solids design of 55% w/w solids. Based on experience, the
leaching results at 40% w/w solids should be representative of those at 55% w/w solids,
comparative testwork in the next Project phase will reduce this risk;
•
Leaching tests with and without pre-aeration were performed on the Eagle, Abra and
CC variability samples to determine a pre-aeration steps’ impact on leach tailings
residual grade, leach extraction and reagent consumption. These comparative tests were
not completed on the master composites due to sample unavailability. Results indicated
that the Abra and CC sample gold and silver extractions are better without pre-aeration
and Eagle sample gold extractions are better with pre-aeration. Pre-aeration prior to
leaching has not been carried in the plant design and future comparative testwork is
recommended on additional variability samples from each Deposit to confirm the
current flowsheet and provide further support for the recovery models;
•
No rheology testwork has been completed on Eagle or Abra samples at varying slurry
densities. This testwork is recommended for the next phase of the Project to finalize
pump, tank and agitator design;
•
Poor thickener overflow clarities were achieved despite high flocculant addition rates,
on a post-leach tailings Abra sample during the dynamic settling testwork. Poor
overflow clarities have the potential to affect Merrill Crowe pregnant solution
clarification. Limited testing was completed due to sample availability. Further
dynamic thickener testwork is recommended on New and Old Abra Deposits, to assess
variability and to optimize thickener and clarifier design criteria; and
•
The dewatering performance of the Old Abra sample was the poorest out of the samples
tested. Future filtration testwork is recommended to assess variability within the new
and Old Abra Zones as well as confirm the filter sizing criteria.
25.1.3
Recovery Methods
Processing plant and recovery method risks include:
•
The annual mill feed copper head grade ranges between 0.10% and 0.19% over the first
9 years of production with short term peaks higher in the first few years of production
coinciding with higher Eagle Deposit tonnes in the process plant feed. High levels of
residual copper in the PLS can impact MC recoveries and carry over to the MC
precipitate/doré reducing payabilities. These risks have been managed by including the
SART plant to lower the residual copper grades in the barren solution and including an
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acid digestion plant in the refinery to lower copper levels in the MC precipitate prior to
smelting. In addition to these, tight copper grade control will be required on the ROM
pad to ensure a smooth copper plant feed head grade;
•
When treating very high-grade gold–silver–copper ore that requires high cyanide and
zinc reagent additions, there is potential for impurities to build up in the recirculating
process water and increasing zinc consumptions or impacting precipitation efficiency.
To mitigate this, the cyanide detoxification circuit was designed to treat an additional
barren bleed stream, if needed, to purge impurities from the process water. Water that
is bled from the circuit will be directed to contact water ponds;
•
Future acid digestion testwork is required in a detailed design bridging phase to confirm
final acid digestion circuit design criteria. During the study, supplier quotes and
expected copper and zinc solubility data for the acid digestion circuit were received
based on their operating experience and benchmarked data from other operations;
•
Rheology testwork will need to be completed in the future detailed design bridging
phase for final underflow pump, leach tank and agitator design criteria and sizing;
•
High CCD thickener overflow clarity (>400 ppm) was observed in the Abra post-leach
thickening testwork results and if experienced in the plant, have the potential to impact
process water and PLS solution quality. The mitigate this risk a clarifier downstream
of the CCD No.1 thickener is included in the flowsheet to provide buffer between the
CCD and MC circuits and provide a cleaner PLS solution to the MC;
•
The plant hydrated lime storage silo is sized for one day of storage. In detailed design,
further discussions with hydrated lime suppliers are needed to confirm the minimum
days of site storage required to satisfy their logistics and delivery requirements;
•
The tailings pressure filter has been sized to provide 25% processing capacity upside
to allow the filter feed tank to be emptied after a filter goes down for cloth washing or
maintenance. There is space on the proposed filter frame to install another six plates to
increase the catch-up capacity from 25% to 35% should this be needed. Provision has
been included in the layout to allow a second filter to installed should this be an option
the operations team want to explore; and
•
From testing to date, the tailings sample from the Old Abra Deposit has displayed the
poorest dewatering characteristics. The tailings filter has been sized on this sample,
however, additional variability filtration testing within the Abra Deposit would provide
further confidence in the selected filter model.
25.1.4
Mining Methods
Underground mining risks include:
•
The historical mine workings may be more extensive at Eagle/New Abra than
surveyed;
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25.2
•
The existing fill in Old Abra stopes may not run freely when undercut;
•
Mining around the Old Abra voids may be more complicated than expected;
•
The HW geotechnical areas (breccias/faults) may be more extensive/impactful than
expected;
•
The bat guano in the historical mine workings may cause health/safety issue/delay/cost;
•
A new underground mining contract could be more expensive than expected (given
inflation and any impacts of global politics or new trade barriers over the last year);
and
•
There may be mine dewatering issues if the historical Old Abra shaft is unusable.
OPPORTUNITIES
25.2.1
Mineral Resources
The Deposits remain open along strike to the northwest and down dip, and there is potential to
increase the current Mineral Resource Estimate.
25.2.2
25.2.3
Mining Methods
•
The paste backfill may require less cement than engineering estimates indicate, which
could reduce operating costs. This would result from having more than the minimum
28-day cure time in some mine locations; and
•
Geotechnical analysis indicates that greater level spacings at areas of the Eagle Vein
are possible, up to 30 m from the designed 20 m.
Dry Stack Tailings Facility
The DSTF is designed to store approximately 5.3 Mt of tailings, and the strategic phasing of lifts
enables progressive reclamation as early as year 2 of mining. A security system for light and heavy
traffic will be installed at the entry to the DSTF and site access to provide safety for the cross
traffic between mining trucks and light vehicles. Access to the DSTF at this location allows for
the separation of light and heavy vehicles, enhancing site operations safety.
25.2.4
Recovery Methods
The processing plant and recovery method opportunities to investigate in a detailed design bridging
phase include:
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25.3
25.3.1
•
Consolidating the utility services, i.e. process water, fresh water, compressed air, etc.,
between the process and SART plants as dedicated utility systems were included within
the SART plant FS design;
•
A review of the mine plan copper head grades and production schedule should be
completed to assess if the SART plant remains economical over the life of the mine;
•
Utilize the water from underground mine dewatering for the process plant fresh water
during the ramp-up of operations and to defer the requirement of the Martin Dam fresh
water intake pumps and plant supply pipeline;
•
Based on installation of the pressure filters reducing the plant fresh water requirement,
there may be opportunity to defer the installation of the Las Huajes Dam fresh water
pumps and plant supply pipeline;
•
Utilize the site contact water from the plant collection pond as paste plant make-up
water to reduce the requirement of fresh water; and
•
There is opportunity to visit a Merrill Crowe operating site that includes an acid
digestion circuit to manage high copper precipitation by leveraging existing site and
supplier relationships. This will allow the design and operations team to further
understand circuit performance, operational challenges and how the circuit is integrated
into the process plant design.
CONCLUSIONS
Location, Mineral Tenure and Geology
GoGold Resources Inc.’s 100% owned Los Ricos South Silver-Gold Property, in northern Jalisco
State, Mexico, is situated in the Hostotipaquillo mining district. The Property contains several
historical underground gold and silver mines that produced 33.3 Moz Ag and 233.5 koz Au
between 1914 and 1930 from approximately 2.45 million tons of ore.
The Property comprises 16 concessions covering 11,994 ha around the historical underground
Cinco Minas silver and gold mine. The Property is not subject to royalties. The Project is situated
75 km northwest of the City of Guadalajara, Mexico at latitude 21° 02’ N and longitude 103° 56’
W.
The Los Ricos South mining district contains seven known gold-silver deposits located along a
4 km portion of the Cinco Minas epithermal quartz vein system. The Cinco Minas quartz vein
strikes northwest–southeast and dips approximately 65° to the west. The vein varies in width from
5 to 30 m, outcrops on surface and has been mined down-dip for 850 m. Late cross-cutting faults
have brecciated, cut and displaced the vein laterally and vertically. The Ag-Au mineralization is
classified as a low-sulphidation epithermal deposit.
GoGold acquired the Property in 2019 and carried out an exploration program to evaluate the
potential for near surface mineralization amenable to bulk mining. In August, 2020, the Company
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signed an agreement with the Ejido of Cinco Minas, which owns the surface rights over all the
concessions included in this Technical Report. The agreement allows GoGold to mine and explore
the 1,280 ha of land that is owned by the local Ejido for a period of 12 years with an option to
renew for a further 12 years.
25.3.2
Drilling
As of the effective date of this Technical Report, GoGold completed 502 HQ size drill holes on
the Los Ricos South Property over a strike length of 1,050 m, from the El Abra workings at the
south end to the Eagle outcrops at the northern end of the vein. An additional 22 drill holes were
completed on the Cerro Colorado area.
In addition to the exploration work completed, the Authors established that the Los Ricos South
Abra and Eagle Deposits contain an additional Exploration Target as follows: 1.8 to 2.2 Mt at 375
to 425 g/t AgEq for 22 to 30 Moz AgEq. The Exploration Target is based on the estimated strike
length, depth and thickness of the known mineralization.
25.3.3
Data Verification
The Authors have evaluated drilling procedures, sample preparation, analyses and security and are
of the opinion that the drill core logging procedures employed, and the sampling methods used
were thorough and have provided assay and geological data of good quality and satisfactory for
use in the current Mineral Resource Estimate.
25.3.4
Metallurgical Testing
Four metallurgical test programs were conducted on samples from the Los Ricos South Deposit
between 2021 and 2024. The FS test program completed by SGS in 2024 included Eagle, Abra
and Cerro Colorado (“CC”) Zone master composite, comminution and variability samples
representing a range of grade, depth and zone parameters from the target mining areas. The 2024
testwork samples provide representative spatial coverage of the Deposits and cover the range of
feed grades expected in the mill feed according to the mine production schedule.
Testwork indicated that the preferred flowsheet for the LRS Deposits should include WOL cyanide
leaching, CCD washing and precious metal recovery by the Merrill Crowe process. The Merrill
Crowe barren solution will be treated through the SART process to liberate the cyanide associated
with the copper cyanide complex as free cyanide, for recycling to leach, and to produce a highgrade copper sulphide product (Cu2S) as a by-product for sale.
The Eagle, Abra and CC samples responded well to cyanidation. Gold and silver extractions at a
k80 grind size of 75 µm and a 96-hour residence time ranged from 77% to 99% for gold and 74%
to 95% for silver. Copper extractions ranged from 44% to 88% for Eagle and Abra and 29% to
52% for CC. High NaCN consumptions were observed during the tests due to the high copper in
the feed. All 2024 variability samples contained <20 g/t antimony, which did not affect leaching
or recovery. Arsenic levels were below detection levels in the 2021, 2022 and 2023 testwork but
reached up to 112 g/t and 182 g/t on a few Abra samples in 2024. The arsenic levels showed no
adverse effects on leaching or recovery. Mercury levels ranged from <0.3 g/t to 2.9 g/t, with no
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adverse recovery effects. A mercury retort is included in the refinery area for mercury removal.
To manage copper concentrations in doré, the SART process and Merrill Crowe precipitate leach
circuit are included in the flowsheet. No other deleterious elements were found at levels impacting
gold-silver payability or process performance.
Robust metallurgical projection models have been derived from bottle roll and bulk leach tests for
gold, silver and copper. Based on the selected flowsheet for the plant operation, over the range of
grades (variability samples) tested, the nominal Eagle and Abra recoveries for gold, silver and
copper were estimated to be 92.6%, 85.5% and 65.4%, respectively. For the CC Zone, which is
introduced to the mill in year 10 of production, a fixed CC gold, silver and copper recovery of
92.5%, 87.6% and 42.6% has been estimated.
25.3.5
Mineral Resources
The Authors completed an Updated Mineral Resource Estimate for the Los Ricos South Project.
Measured plus Indicated Mineral Resources total 11.0 Mt at 1.56 g/t Au, 162 g/t Ag, and
0.12% Cu, or 3.64 g/t AuEq, or 308 g/t Ag Eq, for 549 koz Au, 57,150 koz Ag and 28.1 Mlb Cu,
or 1,284 koz AuEq, or 108,562 koz AgEq. Inferred Mineral Resources total 2.2 Mt at 1.11 g/t Au,
106 g/t Ag, 0.27% Cu, or 2.70 g/t AuEq, or 229 g/t AgEq, for 79 koz Au, 7,473 koz Ag and 13.0
Mlb Cu, or 191 koz AuEq, or 16,191 koz AgEq. The classification of Measured, Indicated, and
Inferred Mineral Resources conforms to CIM (2014) Definition Standards and Best Practices
Guidelines (2019).
The Updated Mineral Resource Estimate reported in this Technical Report is based on drilling and
assay data provided by GoGold and compiled, verified and validated by the Authors. The drilling
database consists of 615 drill holes of which 554 drill holes totalling 90,095 m were used to create
the constraining wireframes employed for the Mineral Resource Estimate. The Authors consider
that the current drill hole database, methodologies, and analytical procedures are appropriate for
the estimation of a Mineral Resource.
The Eagle and Cerro Colorado Veins remain open along strike to the northwest and down dip, and
the Abra Veins remain open down dip. Further drilling may provide additional Mineral Resources.
25.3.6
Mineral Reserve Estimate
The Los Ricos South Project will consist of both underground and open pit mining operations.
Underground mining will commence at the start of production. The Proven and Probable Mineral
Reserve is estimated at 10.2 Mt at 145 g/t Ag, 1.39 g/t Au and 0.10% Cu, or 276 g/t AgEq that
contains 90.7 Moz AgEq. Open pit mining will be initiated in Year 10 and there will be an
underground / open pit overlap period of four years. The duration of all mining activity will be 15
years after the start of production. The life-of-mine ore processing rate is 2,000 tpd or 720,000
tonnes per year.
25.3.7
Mining Methods
The underground mining method is longhole sublevel stoping with cemented paste backfill. The
majority (~70%) of tonnes are extracted from previously unmined areas on the Eagle and Abra
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Veins, with a minority of tonnes (~30%) mined from areas adjacent to Historical Mine Workings.
The Cerro Colorado Vein was not evaluated for underground mining for the purposes of this
Technical Report as there was insufficient tonnes of Measured and Indicated Mineral Resource
above cut-off grade to justify underground mining. All underground mining will be completed by
contractor. Total dilution in the underground mine plan is estimated at 32.1%. Overall average
underground mining loss for the Deposit is estimated at 8.8% by mass. Levels are spaced at a
nominal sublevel interval of 20 m. Longitudinal stopes are a maximum of 15 m along strike, with
a width limited by vein thickness. Stope widths at Eagle average 13.6 m and at Abra average 9.3
m. Haulage trucks will typically be 30 t, with 7 t and 10 t LHD units. Steady-state production of
2,000 tpd will be reached at the end of Year 1 after a ramp-up period, and full production will be
maintained until midway through Year 9, after which underground production declines as open pit
production begins. Over the LOM, a total of 7.51 Mt of ore will be recovered from the underground
mine at average grades of 1.65 g/t Au, 170 g/t Ag and 0.13% Cu, equivalent to 326 g/t AgEq and
a total metal content of 78.8 Moz AgEq.
The Los Ricos South Deposits are near surface and will be mined by conventional open pit mining
methods. A contractor will be engaged to mine the Abra, Cerro Colorado North and Cerro
Colorado South open pits. The contractor will undertake all drill and blast, loading, hauling, and
mine site maintenance activities. The Owner will provide overall mine management and technical
services. Mining will typically be on 5 m high benches, using conventional equipment such as
hydraulic excavators, front-end loaders and 60 t haulage trucks. For open pit dilution, the Mineral
Resource block model was regularized from a percent model to a whole block SMU model.
Regularizing resulted in waste and ore portions of a block being combined into a single block
value. Combined dilution and ore loss during mining was estimated at 10% on tonnes and 9% on
grade at the Abra Deposit. At the narrower vein Cerro Colorado Deposit, combined dilution and
ore loss was estimated at 30% on tonnes and 20% on grade. The overall strip ratio for open pit
mining is 7.6:1. Over the LOM a total of 2.7 Mt of ore will be mined from open pits, at average
grades of 0.66 g/t Au, 77 g/t Ag, 0.02% Cu equivalent to 136 g/t AgEq or 11.9 Moz AgEq. Waste
rock will be transported to nearby storage facilities, and ore will be hauled to the primary crusher
or placed on temporary stockpiles
25.3.8
Recovery Methods
The process plant will be located at the mine site and will receive run of mine ore from the
underground mine. The plant is designed to produce doré as the final product and a copper sulphide
concentrate as the secondary product.
The planned flowsheet is flexible and robust and is based on well-proven unit operations in the
industry, and there are no unique or novel processing methods required for gold and silver
recovery. The flowsheet includes semi-autogenous grinding, pre-leach thickening, whole ore
cyanide leaching, a five-stage counter-current decantation circuit, a Merrill Crowe circuit, acid
digestion, smelting, a sulphidization, acidification, recycle and thickening (SART process) circuit,
tailing cyanide detoxification, thickening and filtration.
The process plant and major equipment was designed for a nominal throughput of 2,000 tpd at a
plant availability of 92%.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 503 of 577
25.3.9
Project Infrastructure
Electrical power to the Project site will be supplied by constructing two sets of 34.5 kVA overhead
lines along a 41 km route from the CFE substation at the La Yesca Dam power station.
The Company will construct infrastructure for administration offices, warehousing, maintenance
facilities, cafeteria, fuel storage and distribution, explosives storage, and water and sewage
treatment. The contractors will establish infrastructure for offices and warehousing. The Company
will not be supplying on-site housing. Employees and contractors will commute from nearby
towns.
Process facilities include the process plant, crushing facilities, process plant workshop, assay
laboratory, fresh water infrastructure, SART Plant, and paste backfill plant.
Raw water will be supplied from two dams located approximately 6 km and 12 km from the
process plant.
A site will be prepared for approximately 5.3 Mt of dry stack tailings with a water management
system. The tailings will also be used for cemented paste backfill in the underground mine.
25.3.10
Environmental
Exploration activities are currently being conducted under a permit from the Mexican government
Secretary of the Environment and Natural Resources (“SEMARNAT”). An extensive list of
Federal and State permits will be required before mining can commence, along with environmental
impact studies. GoGold has engaged the Mexican firm CIMA to conduct an environmental
baseline study. CIMA has completed a socio-economic baseline study that considers economic,
cultural, social, demographic, and geographical aspects of the local communities. The mining
Project is expected to represent minimal risk to the processes and structure of the local ecosystem,
and to the roads, livestock and agricultural activities.
25.3.11
Capital and Operating Costs
Initial capital costs are estimated at $227M and include a 10% contingency of $21M. Sustaining
capital costs over the production years are estimated at $100M mainly for underground mine
development, increases to the dry stack tailings storage capacity, and open pit pre-stripping.
Underground OPEX mining costs have been estimated to average $44.04/t processed over the
LOM, and paste backfill plant costs have been estimated at $5.89/t processed. Open pit mining
costs have been estimated to average $2.64/t material mined or $18.46/t processed during the open
pit production years. The average LOM mining cost (including backfill) is estimated at $42.92/t
processed. Processing costs ($39.63/t processed, consisting of $33.36/t for processing, $2.13/t for
tailings handling and $4.13/t for SART) and site G&A ($6.88/t processed) contribute to a total
LOM average cost estimated at $89.43/t processed. The average operating cash cost over the
production years is estimated at $11.59/oz AgEq, and the average all-in sustaining cost is estimated
at $12.78/oz AgEq.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 504 of 577
25.3.12
Economic Analysis
Using metal prices based on approximate December 31, 2024 three-year monthly trailing averages,
and Consensus Economics Inc. long-term projections from various financial institutions, of
US$2,330/oz Au, US$26.80/oz Ag and US$4.00/lb Cu, the Project has an estimated pre-tax NPV
at a 5% discount rate of $851M and an IRR of 39%. After-tax NPV and IRR are estimated at
$355M and 28%, respectively. Simple after-tax payback of initial capital for underground mining
is 2.6 years. Mexican corporate income tax is levied at a rate of 35% on the net taxable income,
and the Project is subject to a 0.5% NSR mining tax payable to the Mexican government. Project
economics are most sensitive to metal prices. Project economics are more sensitive to overall
operating costs than capital costs.
25.3.13
Risks and Opportunities
Overall Project risks are perceived by the Authors as moderate. Eight high risks items were
identified in a risk workshop, consisting of issues on Project constructability, open pit mining
permitting, dry stack tailings storage, and historical underground mine workings. The main
opportunity is that the Deposits remain open along strike to the northwest and down dip, and there
is potential to increase the current Mineral Resource Estimate.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 505 of 577
26.0
RECOMMENDATIONS
Based on the results of GoGold’s exploration work and the positive results of this FS, the Authors
recommend that GoGold continue with Project development and permitting activities on the Los
Ricos South Property and proceed with detailed design studies.
26.1
METALLURGICAL TESTING
Further testwork is recommended to optimize the current flowsheet and design criteria in a detailed
design bridging phase. A testwork program, using industry standard testwork procedures, focusing
on the major Deposits, is recommended. The cost of this testwork is estimated at $0.25M.
Before selecting any further variability samples, a gap analysis examining the FS samples tested
against the final FS stope shapes and mine production plan should be completed.
The following work is recommended:
•
Conduct further additional leaching bottle roll tests on Eagle and Abra variability
samples representative of year 1 to 5 of the FS mine plan to confirm the final flowsheet
and support the recovery models. Comparative tests on individual samples should
compare gold and silver recovery at:
o 40% w/w solids versus 55% w/w solids
o With pre-aeration versus no pre-aeration
o Oxygen versus air for sparging;
•
Conduct acid digestion testwork to confirm circuit design criteria and leverage supplier
relationships to complete site visits to benchmark the expected performance against
operational data;
•
Conduct cyanide detox optimisation testwork on an Abra sample representative of year
1 to 3 of the mine plan;
•
Conduct oxygen uptake testing on Eagle and Abra samples to determine leaching
air/oxygen requirements to confirm sparging system design;
•
Perform ore materials handling testwork;
•
Identify species in solution that may inhibit Merrill Crowe precipitation, for variability
samples across the Eagle and Abra Deposits, such as sulphides, arsenic, antimony,
copper cyanide, nickel and cobalt. Tests should be run at the expected concentrations
and conditions should be optimized to maximize precipitation efficiency;
•
Conduct pressure filtration testwork on New and Old Abra variability samples to
confirm filter design;
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 506 of 577
26.2
•
Conduct rheology testwork at varying percent solids for composite Eagle and Abra
samples to confirm tank and agitator design;
•
Perform dynamic settling on Eagle and Abra detoxified tailings samples to confirm
tailings thickener design; and
•
Perform further Abra leached tailings dynamic and settling testwork, aiming to
optimize flocculant consumption, underflow density and reduce overflow clarity.
Dynamic settling testwork on the overflow is recommended to confirm a 1 m/s rising
velocity, with the assistance of coagulant, is acceptable for achieving a final MC circuit
feed clarity of less than 100 ppm.
DSTF GEOTECHNICAL RECOMMENDATIONS
For the next phase work plan, it is strongly recommended that vehicular road access be constructed
to reach the circumference of the DSTF area to facilitate the preparation of drilling platforms. This
would allow for the establishment of a satellite camp and helicopter staging site from which
unitized drill rigs could be transported and placed at steep platform sites using long-line methods.
The proposed FS geotechnical drilling program assumes that the specified collar locations will be
accessible. However, it is likely that this will not always be the case, and alternative platform
locations may need to be developed to access the intended subsurface targets.
It is recommended to pursue further data collection of the subsurface in the DSTF area. The
objective of this program is to investigate variability in soil horizons and confirm foundation
loading, especially for the unexplored areas of the DSTF, to characterize and verify geotechnicalgeological features, distribution, and preliminary characterization of the geotechnical units,
foundation materials, and groundwater level. Data collected will investigate variability in soil
horizons and confirm foundation loading. A geotechnical investigation program is recommended,
including seven drill holes and seven test pits at the DSTF area. An estimated cost for the program
is $1.05M for field investigations and $0.07M for laboratory testing, totalling $1.12M.
26.3
UNDERGROUND GEOTECHNICAL RECOMMENDATIONS
The completed geotechnical and hydrogeological program has known data gaps due to the
increased underground mine footprint during the advancement of the Project. To mitigate data
uncertainty and confirm geotechnical design assumptions further geotechnical drilling is required
on the order of a minimum of 6,000 m. The program should include the following:
•
•
•
•
•
Geotechnical logging of drill holes;
Laboratory strength and elastic modulus testing;
Televiewers surveys in available drill holes;
VWP installation in the deepest area of the mine; and
Additional hydraulic conductivity testing in areas of blocky ground and deeper mining
areas.
An estimated cost for the program is $1.5M for field investigations and engineering study.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 507 of 577
26.4
OPEN PIT GEOTECHNICAL RECOMMENDATIONS
There is limited geotechnical information available regarding the Cerro Colorado Pits. The current
recommendations are based on extrapolating the data which have introduced uncertainties to the
study. It is recommended to collect more data in these regions by drilling a minimum of four
oriented boreholes (two per pit) with full geotechnical logging and laboratory strength testing. An
estimated cost for the program is $0.2M for field investigations and engineering study.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 508 of 577
27.0
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Cox, W. Roland. 1911. Report on the Treatment of Ore from Cinco Minas, Jalisco, Mexico with
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Company Report prepared for the Cinco Minas Company.
Middelkamp, L. L. 1928. Annual Report for the Year 1927, Cinco Minas Company.
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28.0
CERTIFICATES
CERTIFICATE OF QUALIFIED PERSON
ANDREW BRADFIELD, P. ENG.
I, Andrew Bradfield, P. Eng., residing at 5 Patrick Drive, Erin, Ontario, Canada, N0B 1T0, do hereby certify that:
1.
I am an independent mining engineer contracted by P&E Mining Consultants.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate of the
Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14, 2025.
3.
I am a graduate of Queen’s University, with an honours B.Sc. degree in Mining Engineering in 1982. I have practiced
my profession continuously since 1982. I am a Professional Engineer of Ontario (License No.4894507). I am also a
member of the National CIM.
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that,
by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work
experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
My summarized career experience is as follows:
• Various Engineering Positions – Palabora Mining Company, South Africa
• Mines Project Engineer – Falconbridge Limited, Sudbury
• Senior Mining Engineer – William Hill Mining Consultants Limited, Toronto
• Independent Mining Engineer,
• GM Toronto – Bharti Engineering Associates Inc,
• VP Technical Services, GM of Australian Operations – William Resources Inc,
• Independent Mining Engineer,
• Principal Mining Engineer – SRK Consulting, Toronto
• COO – China Diamond Corp, China
• VP Operations – TVI Pacific Inc, Philippines
• COO – Avion Gold Corporation, Mali
• Independent Mining Engineer,
1982-1986
1986-1987
1987-1990
1990-1991
1991-1996
1996-1999
1999-2001
2001-2003
2003-2006
2006-2008
2008-2012
2012-Present
4.
I have not visited the Property that is the subject of this Technical Report.
5.
I am responsible for authoring Sections 1.1, 1.4, 1.5, 1.8, 1.20-1.24, 2, 3, 15.1, 15.2, 15.4.1, 15.4.2, 15.4.3-15.4.5, 16.2,
16.3, 18.3.2, 18.10.2, 18.10.4, 19, 21.1, 21.2.1, 21.2.2.5, 21.2.2.8, 21.2.2.9, 21.2.3, 21.3.1, 21.3.4, 21.3.8, 21.4-21.6, 22,
25.1.1, 25.3.6, 25.3.7, 25.3.11-25.3.13 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and the
Property.
7.
I have had prior involvement with the Project that is the subject of this Technical Report. I was a “Qualified Person”
for a Technical Report titled “Preliminary Economic Assessment of the Los Ricos South Project, Jalisco, Mexico”, with
an effective date of January 20, 2021, and for a Technical Report titled “Updated Mineral Resource Estimate and
Preliminary Economic Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective date of September 12,
2023.
8.
I have read NI 43-101 and Form 43-101F1. This Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the Technical
Report contains all scientific and technical information that is required to be disclosed to make the Technical Report
not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[Andrew Bradfield]
____________________________
Andrew Bradfield, P.Eng.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 513 of 577
CERTIFICATE OF QUALIFIED PERSON
JARITA BARRY, P.GEO.
I, Jarita Barry, P.Geo., residing at 9052 Mortlake-Ararat Road, Ararat, Victoria, Australia, 3377, do hereby certify
that:
1.
I am an independent geological consultant contracted by P&E Mining Consultants Inc.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate
of the Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14,
2025.
3.
I am a graduate of RMIT University of Melbourne, Victoria, Australia, with a B.Sc. in Applied Geology. I have
worked as a geologist for over 19 years since obtaining my B.Sc. degree. I am a geological consultant currently
licensed by Engineers and Geoscientists British Columbia (License No. 40875) and Professional Engineers and
Geoscientists Newfoundland & Labrador (License No. 08399), and I am registered as a Temporary Registrant
with Professional Geoscientists Ontario (Registration No. 3888). I am also a member of the Australasian Institute
of Mining and Metallurgy of Australia (Member No. 305397).
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify
that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past
relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
My relevant experience for the purpose of the Technical Report is:
• Geologist, Foran Mining Corp.
• Geologist, Aurelian Resources Inc.
• Geologist, Linear Gold Corp.
• Geologist, Búscore Consulting
• Consulting Geologist (AusIMM)
• Consulting Geologist, P.Geo. (APEGBC/AusIMM)
2004
2004
2005-2006
2006-2007
2008-2014
2014-Present
4.
I have not visited the Property that is the subject of this Technical Report.
5.
I am responsible for authoring Sections 1.10, 1.11, 11, 12, 25.3.3 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and
the Property.
7.
I have had prior involvement with the Project that is the subject of this Technical Report. I was a “Qualified
Person” for a Technical Report titled “Technical Report and Initial Mineral Resource Estimate of the Los Ricos
South Project, Jalisco, Mexico”, with an effective date of July 28, 2020, and a Technical Report titled “Preliminary
Economic Assessment of the Los Ricos South Project, Jalisco, Mexico”, with an effective date of January 20,
2021, and for a Technical Report titled “Updated Mineral Resource Estimate and Preliminary Economic
Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[Jarita Barry]
________________________________
Jarita Barry, P.Geo.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 514 of 577
CERTIFICATE OF QUALIFIED PERSON
FRED BROWN, P.GEO.
I, Fred Brown, of PO Box 332, Lynden, WA, USA, do hereby certify that:
1.
I am an independent geological consultant and have worked as a geologist continuously since my graduation from
university in 1987.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate
of the Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14,
2025.
3.
I graduated with a Bachelor of Science degree in Geology from New Mexico State University in 1987.
I obtained a Graduate Diploma in Engineering (Mining) in 1997 from the University of the Witwatersrand and a
Master of Science in Engineering (Civil) from the University of the Witwatersrand in 2005. I am registered with
the Association of Professional Engineers and Geoscientists of British Columbia as a Professional Geoscientist
(171602) and the Society for Mining, Metallurgy and Exploration as a Registered Member (#4152172).
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify
that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past
relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
My relevant experience for the purpose of the Technical Report is:
• Underground Mine Geologist, Freegold Mine, AAC
• Mineral Resource Manager, Vaal Reefs Mine, Anglogold
• Resident Geologist, Venetia Mine, De Beers
• Chief Geologist, De Beers Consolidated Mines
• Consulting Geologist
• P&E Mining Consultants Inc. – Sr. Associate Geologist
1987-1995
1995-1997
1997-2000
2000-2004
2004-2008
2008-Present
4.
I have visited the Property that is the subject of this Technical Report on August 15-16, 2019.
5.
I am responsible for authoring Sections 1.13, 14.1, 14.3-14.11, 14.12.1, 14.12.3-14.12.11, 25.1.4, 25.3.5 of this
Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101.
7.
I have had prior involvement with the Project that is the subject of this Technical Report. I was a “Qualified
Person” for a Technical Report titled “Technical Report and Initial Mineral Resource Estimate of the Los Ricos
South Project, Jalisco, Mexico”, with an effective date of July 28, 2020, and a Technical Report titled “Preliminary
Economic Assessment of the Los Ricos South Project, Jalisco, Mexico”, with an effective date of January 20,
2021, and for a Technical Report titled “Updated Mineral Resource Estimate and Preliminary Economic
Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[Fred Brown]
__________________________
Fred Brown, P.Geo.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 515 of 577
CERTIFICATE OF QUALIFIED PERSON
DAVID BURGA, P.GEO.
I, David Burga, P. Geo., residing at 3884 Freeman Terrace, Mississauga, Ontario, Canada, L5M 6P6 do hereby certify
that:
1.
I am an independent geological consultant contracted by P & E Mining Consultants Inc.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate
of the Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14,
2025.
3.
I am a graduate of the University of Toronto with a Bachelor of Science degree in Geological Sciences (1997).
I have worked as a geologist for over 20 years since obtaining my B.Sc. degree. I am a geological consultant
currently licensed by the Association of Professional Geoscientists of Ontario (License No 1836).
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify
that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past
relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
My relevant experience for the purpose of the Technical Report is:
• Exploration Geologist, Cameco Gold
• Field Geophysicist, Quantec Geoscience
• Geological Consultant, Andeburg Consulting Ltd.
• Geologist, Aeon Egmond Ltd.
• Project Manager, Jacques Whitford
• Exploration Manager – Chile, Red Metal Resources
• Consulting Geologist
1997-1998
1998-1999
1999-2003
2003-2005
2005-2008
2008-2009
2009-Present
4.
I have visited the Property that is the subject of this Technical Report on May 16-17, 2023.
5.
I am responsible for authoring Sections 1.9, 10, 25.3.2 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101.
7.
I have had prior involvement with the Property that is the subject of this Technical Report. I was a “Qualified
Person” for a Technical Report titled “Technical Report and Initial Mineral Resource Estimate of the Los Ricos
South Project, Jalisco, Mexico”, with an effective date of July 28, 2020, and a Technical Report titled “Preliminary
Economic Assessment of the Los Ricos South Project, Jalisco, Mexico”, with an effective date of January 20,
2021, and for a Technical Report titled “Updated Mineral Resource Estimate and Preliminary Economic
Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[David Burga]
____________________________
David Burga, P.Geo.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 516 of 577
CERTIFICATE OF QUALIFIED PERSON
D. GRANT FEASBY, P.ENG.
I, D. Grant Feasby, P. Eng., residing at 12,209 Hwy 38, Tichborne, Ontario, Canada, K0H 2V0, do hereby certify that:
1.
I am currently the Owner and President of:
FEAS - Feasby Environmental Advantage Services
38 Gwynne Ave, Ottawa, Ontario, Canada, K1Y1W9
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate
of the Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14,
2025.
3.
I graduated from Queens University in Kingston Ontario, in 1964 with a Bachelor of Applied Science in
Metallurgical Engineering, and a Master of Applied Science in Metallurgical Engineering in 1966. I am a
Professional Engineer registered with Professional Engineers Ontario. I have worked as a metallurgical engineer
for over 50 years since my graduation from university.
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify
that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past
relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
My relevant experience for the purpose of the Technical Report has been acquired by the following activities:
• Metallurgist, Base Metal Processing Plant.
• Research Engineer and Lab Manager, Industrial Minerals Laboratories in USA and Canada.
• Research Engineer, Metallurgist and Plant Manager in the Canadian Uranium Industry.
• Manager of Canadian National Programs on Uranium and Acid Generating Mine Tailings.
• Director, Environment, Canadian Mineral Research Laboratory.
• Senior Technical Manager, for large gold and bauxite mining operations in South America.
• Expert Independent Consultant associated with several companies, including P&E Mining Consultants,
on mineral processing, environmental management, and mineral-based radiation assessment.
4.
I have not visited the Property that is the subject of this Technical Report.
5.
I am responsible for authoring Sections 1.19, 20, 25.3.10 of this Technical Report.
6.
I am independent of the issuer applying the test in Section 1.5 of NI 43-101.
7.
I have had prior involvement with the Project that is the subject of this Technical Report. I was a “Qualified
Person” for a Technical Report titled “Technical Report and Initial Mineral Resource Estimate of the Los Ricos
South Project, Jalisco, Mexico”, with an effective date of July 28, 2020, and a Technical Report titled “Preliminary
Economic Assessment of the Los Ricos South Project, Jalisco, Mexico”, with an effective date of January 20,
2021, and for a Technical Report titled “Updated Mineral Resource Estimate and Preliminary Economic
Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[D. Grant Feasby]
________________________________
D. Grant Feasby, P.Eng.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 517 of 577
CERTIFICATE OF QUALIFIED PERSON
EUGENE PURITCH, P.ENG., FEC, CET
I, Eugene Puritch, P. Eng., FEC, CET, residing at 44 Turtlecreek Blvd., Brampton, Ontario, Canada, L6W 3X7, do hereby
certify that:
1.
I am an independent mining consultant and President of P&E Mining Consultants Inc.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate of the
Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14, 2025.
3.
I am a graduate of The Haileybury School of Mines, with a Technologist Diploma in Mining, as well as obtaining an
additional year of undergraduate education in Mine Engineering at Queen’s
University.
In addition, I have also met the Professional Engineers of Ontario Academic Requirement Committee’s Examination
requirement for Bachelor’s Degree in Engineering Equivalency. I am a mining consultant currently licensed by the:
Professional Engineers and Geoscientists New Brunswick (License No. 4778); Professional Engineers, Geoscientists
Newfoundland and Labrador (License No. 5998); Association of Professional Engineers and Geoscientists
Saskatchewan (License No. 16216); Ontario Association of Certified Engineering Technicians and Technologists
(License No. 45252); Professional Engineers of Ontario (License No. 100014010); Association of Professional
Engineers and Geoscientists of British Columbia (License No. 42912); and Northwest Territories and Nunavut
Association of Professional Engineers and Geoscientists (No. L3877). I am also a member of the National Canadian
Institute of Mining and Metallurgy.
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that,
by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work
experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
I have practiced my profession continuously since 1978. My summarized career experience is as follows:
• Mining Technologist - H.B.M.& S. and Inco Ltd.,
1978-1980
• Open Pit Mine Engineer – Cassiar Asbestos/Brinco Ltd.,
1981-1983
• Pit Engineer/Drill & Blast Supervisor – Detour Lake Mine,
1984-1986
• Self-Employed Mining Consultant – Timmins Area,
1987-1988
• Mine Designer/Resource Estimator – Dynatec/CMD/Bharti,
1989-1995
• Self-Employed Mining Consultant/Resource-Reserve Estimator,
1995-2004
• President – P&E Mining Consultants Inc,
2004-Present
4.
I have not visited the Property that is the subject of this Technical Report.
5.
I am responsible for authoring Sections14.2 and 14.12.2 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101.
7.
I have had prior involvement with the Project that is the subject of this Technical Report. I was a “Qualified Person”
for a Technical Report titled “Technical Report and Initial Mineral Resource Estimate of the Los Ricos South Project,
Jalisco, Mexico”, with an effective date of July 28, 2020, and a Technical Report titled “Preliminary Economic
Assessment of the Los Ricos South Project, Jalisco, Mexico”, with an effective date of January 20, 2021, and for a
Technical Report titled “Updated Mineral Resource Estimate and Preliminary Economic Assessment of the Los Ricos
Project, Jalisco, Mexico”, with an effective date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1. This Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief,
the Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[Eugene Puritch]
____________________________
Eugene Puritch, P.Eng., FEC, CET
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 518 of 577
CERTIFICATE OF QUALIFIED PERSON
D. GREGORY ROBINSON, P. ENG.
I, David Gregory (Greg) Robinson, P. Eng. (ON), residing at 14 Moorcrest Drive, Aurora, Ontario, Canada, do hereby
certify that:
1.
I am an independent engineering consultant contracted by P&E Mining Consultants Inc.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate
of the Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14,
2025.
3.
I am a graduate of Dalhousie University, Queens University and Cornell University, and Professional Engineer
of Ontario (License No. 100216726).
I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify
that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past
relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.
I have practiced my profession continuously since 2008. My summarized career experience is as follows:
• Associate Engineer, P&E Mining Consultants
Aug 2017 - Present
• Mine Engineer, Lac des Iles Mine, North American Palladium
May 2016 – Jun 2017
• Senior Underground Engineer, Phoenix Gold, Rubicon Minerals
Sep 2014 – Jan 2016
• Mine Engineer, Diavik Diamond Mine, Rio Tinto Diamonds
Sep 2011 – Sep 2014
• Mine Engineer, Bengalla Mine, Rio Tinto Coal and Allied
Dec 2008 – Sep 2011
• EIT, Creighton Mine, Vale-Inco
May 2008 – Dec 2008
4.
I have visited the Property that is the subject of this Technical Report on November 28-29, 2023.
5.
I am responsible for authoring Sections 1.14, 1.15, 15.3, 16.1, 16.1.1, 16.1.2.2-16.1.2.8, 16.1.3-16.1.10, 18.10.1,
21.2.2.2, 21.3.2, 25.1.4, 25.2.2 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101. I am independent of the Vendor and
the Property.
7.
I have had prior involvement with the Property that is the subject of this Technical Report. I was a “Qualified
Person” for a Technical Report titled “Preliminary Economic Assessment of the Los Ricos South Project, Jalisco,
Mexico”, with an effective date of January 20, 2021, and for a Technical Report titled “Updated Mineral Resource
Estimate and Preliminary Economic Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective
date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1. This Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[D. Gregory Robinson]
____________________________
D. Gregory Robinson, P.Eng.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 519 of 577
CERTIFICATE OF QUALIFIED PERSON
WILLIAM STONE, PH.D., P.GEO.
I, William Stone, Ph.D., P.Geo, residing at 4361 Latimer Crescent, Burlington, Ontario, Canada, do hereby certify that:
1.
I am an independent geological consultant contracted by P & E Mining Consultants Inc.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate of the
Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14, 2025.
3.
I am a graduate of Dalhousie University with a Bachelor of Science (Honours) degree in Geology (1983).
In addition, I have a Master of Science in Geology (1985) and a Ph.D. in Geology (1988) from the University of Western
Ontario. I have worked as a geologist for a total of 35 years since obtaining my M.Sc. degree. I am a geological
consultant currently licensed by the Professional Geoscientists of Ontario (License No 1569).
I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that, by
reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work
experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.
My relevant experience for the purpose of the Technical Report is:
• Contract Senior Geologist, LAC Minerals Exploration Ltd.
• Post-Doctoral Fellow, McMaster University
• Contract Senior Geologist, Outokumpu Mines and Metals Ltd.
• Senior Research Geologist, WMC Resources Ltd.
• Senior Lecturer, University of Western Australia
• Principal Geologist, Geoinformatics Exploration Ltd.
• Vice President Exploration, Nevada Star Resources Inc.
• Vice President Exploration, Goldbrook Ventures Inc.
• Vice President Exploration, North American Palladium Ltd.
• Vice President Exploration, Magma Metals Ltd.
• President & COO, Pacific North West Capital Corp.
• Consulting Geologist
• Senior Project Geologist, Anglo American
• Consulting Geoscientist
1985-1988
1988-1992
1993-1996
1996-2001
2001-2003
2003-2004
2005-2006
2006-2008
2008-2009
2010-2011
2011-2014
2013-2017
2017-2019
2020-Present
4.
I have not visited the Property that is the subject of this Technical Report.
5.
I am responsible for authoring Sections 1.2-1.8, 4, 5, 6, 7, 8, 9, 23, 25.3.1 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101.
7.
I have had prior involvement with the Property that is the subject of this Technical Report. I was a “Qualified Person”
for a Technical Report titled “Technical Report and Initial Mineral Resource Estimate of the Los Ricos South Project,
Jalisco State, Mexico” with an effective date of July 28, 2020, and a Technical Report titled “Preliminary Economic
Assessment of the Los Ricos South Project, Jalisco, Mexico”, with an effective date of January 20, 2021, and for a
Technical Report titled “Updated Mineral Resource Estimate and Preliminary Economic Assessment of the Los Ricos
Project, Jalisco, Mexico”, with an effective date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the Technical
Report contains all scientific and technical information that is required to be disclosed to make the Technical Report
not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[William Stone]
____________________________
Dr. William Stone, P.Geo.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 520 of 577
CERTIFICATE OF QUALIFIED PERSON
DAVID SALARI, P.ENG.
I, David Salari, P.Eng., of 125 Bronte Road, Unit 503, Oakville, Ontario, Canada, L6L 0H1, do hereby certify that:
1.
I am an independent metallurgical engineer with an office at Suite 300-10, 1100 Burloak Drive, Burlington,
Ontario, Canada, L6L 2Y8.
2.
This certificate applies to the Technical Report titled “Feasibility Study and Updated Mineral Resource Estimate
of the Los Ricos South Project, Jalisco, Mexico”, (The “Technical Report”) with an effective date of January 14,
2025.
3.
I am a graduate University of Toronto with a Bachelor's of Applied Science (BASc) – Metallurgy and Material
Science. I have been actively involved in mining and mineral processing since 1980 with extensive experience in
metallurgical and mill testing and design, mill capital and operating costs, construction, commissioning, and mill
operations.
I am a member in good standing of the Professional Engineers Ontario - #40416505 and I am the designated
P.Eng. for D.E.N.M. Engineering Ltd. – Certificate of Authorization – Professional Engineers Ontario #100102038 and Designation as a Consulting Engineer – Professional Engineers Ontario - # 4012.
I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify
that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past
relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.
4.
I have visited the Property that is the subject of this Technical Report on May 15-16, 2021.
5.
I am responsible for authoring Sections 1.12.2, 13.2.1.8, 17.4.7, 21.2.2.4, 21.3.6 of this Technical Report.
6.
I am independent of the Issuer applying the test in Section 1.5 of NI 43-101.
7.
I have had prior involvement with the Property that is the subject of this Technical Report. I was a “Qualified
Person” for a Technical Report titled “Preliminary Economic Assessment of the Los Ricos South Project, Jalisco,
Mexico”, with an effective date of January 20, 2021, and for a Technical Report titled “Updated Mineral Resource
Estimate and Preliminary Economic Assessment of the Los Ricos Project, Jalisco, Mexico”, with an effective
date of September 12, 2023.
8.
I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.
9.
As of the effective date of this Technical Report, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be disclosed to make the
Technical Report not misleading.
Effective Date: January 14, 2025
Signing Date: February 28, 2025
{SIGNED AND SEALED}
[David Salari]
____________________________
David Salari, P.Eng.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 521 of 577
CERTIFICATE OF QUALIFIED PERSON
TOMMASO ROBERTO RAPONI, P.ENG.
I, Tommaso Roberto Raponi, P.Eng. certify that I am employed as a Senior Mineral Processing Specialist
with Ausenco Engineering Canada ULC (“Ausenco”), with an office address of 15th Floor, 11 King Street
West Toronto Ontario M5H 4C7 Canada.
This certificate applies to the technical report titled FEASIBILITY STUDY AND UPDATED MINERAL
RESOURCE ESTIMATE OF THE LOS RICOS SOUTH PROJECT, JALISCO, MEXICO that has an effective date of
January 14th, 2025, and a report date of February 28th, 2025 (the “Technical Report”).
1. I graduated from the University of Toronto with a Bachelor of Applied Science degree in Geological
Engineering with a specialization in Mineral Processing in 1984.
2. I am a Professional Engineer registered with the Professional Engineers Ontario (No. 90225970),
Engineers and Geoscientists British Columbia (No. 23536) and NWT and Nunavut Association of
Professional Engineers and Geoscientists (No. L4508).
3. I have practiced my profession continuously for over 39 years with experience in the
development, design, operation and commissioning of mineral processing plants, focusing on gold
projects, both domestic and internationally.
4. I have read the definition of “Qualified Person” set out in the National Instrument 43-101
Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my
education, affiliation to a professional association and past relevant work experience, I fulfill the
requirements to be a “Qualified Person” for those sections of the Technical Report that I am
responsible for preparing.
5. I have not visited the Los Ricos South project.
6. I am responsible for Sections 1.12.1, 1.16, 1.17.2, 1.24.1, 13.1, 13.2.1.1, 13.2.1.2, 13.2.1.3,
13.2.1.4, 13.2.1.5, 13.2.1.6, 13.2.1.7, 13.2.1.9, 13.2.1.10, 13.3, 13.4, 13.5, 17.1, 17.2, 17.3, 17.4.1,
17.4.2, 17.4.3, 17.4.4, 17.4.5, 17.4.6, 17.4.8, 17.4.9, 17.5, 17.6, 17.7, 18.1, 18.2, 18.3.1, 18.3.3,
18.3.4, 18.4, 18.5, 18.6, 18.7, 18.8, 18.9, 18.10.3, 21.2.2.1, 21.2.2.3, 21.2.2.6, 21.2.2.7, 21.3.3,
21.3.5, 24, 25.1.2, 25.1.3, 25.2.4, 25.3.4, 25.3.8, 25.3.9, and 26.1 of the Technical Report.
7. I am independent of GoGold Resources Inc. as independence is defined in Section 1.5 of NI 43101.
8. I have had no previous involvement with the Los Ricos South project.
9. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have
been prepared in compliance with that Instrument. As of the effective date of the Technical
Report, to the best of my knowledge, information and belief, the sections of the Technical Report
for which I am responsible contain all scientific and technical information that is required to be
disclosed to make those sections of the Technical Report not misleading.
Dated: February 28th, 2025
“Signed and sealed”
Tommaso Roberto Raponi, P.Eng.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 522 of 577
CERTIFICATE OF QUALIFIED PERSON
SCOTT CAMERON ELFEN, P.E.
I, Scott Cameron Elfen, P.E. certify that I am employed as Global Lead Geotechnical Services with Ausenco
Engineering Canada ULC (“Ausenco”), with an office address of 1050 West Pender, Suite 1200 Vancouver
British Columbia V6E 3S7 Canada.
This certificate applies to the technical report titled FEASIBILITY STUDY AND UPDATED MINERAL
RESOURCE ESTIMATE OF THE LOS RICOS SOUTH PROJECT, JALISCO, MEXICO that has an effective date of
January 14th, 2025, and a report date of February 28th, 2025 (the “Technical Report”).
1. I graduated from the University of California, Davis, California, in 1991 with Bachelor of Science
degree in Civil Engineering (Geotechnical).
2. am a Registered Civil Engineer in the State of California (license no. C56527) by exam since 1996
and I am also a member in good standing of the American Society of Civil Engineers (ASCE), and
the Society for Mining, Metallurgy & Exploration (SME).
3. have practiced my profession continuously for 26 years with experience in the development,
design, construction and operations of mine waste storage facilities, such as waste rock storage
facilities and tailings storage facilities ranging from slurry to dry stack facilities, focusing on
precious and base metals, both domestic and international. In addition, I have developed
geotechnical design parameters for pit slope design, plant foundation design, and other
supporting infrastructure. Examples of projects I have worked on include Skeena’s Eskay Creek
Project PEA, PFS and FS, O3 Mining’s Marban Project PEA and PFS, First Mining Gold’s Springpole
PEA and PFS. SSR Mining’s Puna Silver In-Pit Tailings Disposal PFS, and Detailing Engineering, and
the Company’s Cangrejos Project PEA
4. I have read the definition of “Qualified Person” set out in the National Instrument 43-101
Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my
education, affiliation to a professional association and past relevant work experience, I fulfill the
requirements to be a “Qualified Person” for those sections of the Technical Report that I am
responsible for preparing.
5. I visited the Los Ricos South project on August 13th, 2024.
6. I am responsible for Sections 1.17.1, 1.24.2, 18.12, 21.3.7, 25.2.3, and 26.2 of the Technical
Report.
7. I am independent of GoGold Resources Inc. as independence is defined in Section 1.5 of NI 43101.
8. I have had no previous involvement with the Los Ricos South project.
9. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have
been prepared in compliance with that Instrument. As of the effective date of the Technical
Report, to the best of my knowledge, information and belief, the sections of the Technical Report
for which I am responsible contain all scientific and technical information that is required to be
disclosed to make those sections of the Technical Report not misleading.
Dated: February 28th, 2025
“Signed and sealed”
Scott Cameron Elfen, P.E.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 523 of 577
CERTIFICATE OF QUALIFIED PERSON
JONATHAN COOPER, P.ENG.
I, Jonathan Cooper, P.Eng. certify that I am employed as Team Lead – Water Resources with Ausenco
Engineering Canada ULC (Ausenco), with an office address of 11 King Street West, Suite 1500, Toronto,
Ontario M5H 4C7, Canada.
This certificate applies to the technical report titled FEASIBILITY STUDY AND UPDATED MINERAL
RESOURCE ESTIMATE OF THE LOS RICOS SOUTH PROJECT, JALISCO, MEXICO that has an effective date of
January 14th, 2025, and a report date of February 28th, 2025 (the “Technical Report”).
1. I graduated from the University of Western Ontario with a Civil Engineering in 2008, and the
University of Edinburgh with a Master of Environmental Management in 2010.
2. I am a Professional Engineered registered with the Engineers and Geoscientists British Columbia
(No. 37864) and Professional Engineers Ontario (No. 100191626).
3. I have practiced my profession continuously for over 15 years with experience in the
development, design, operation, and commissioning of surface water infrastructure
4. I have read the definition of “Qualified Person” set out in the National Instrument 43-101
Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my
education, affiliation to a professional association and past relevant work experience, I fulfill the
requirements to be a “Qualified Person” for those sections of the Technical Report that I am
responsible for preparing.
5. I have not visited the Los Ricos South project.
6. I am responsible for Sections 1.17.3 and 18.11 of the Technical Report.
7. I am independent of GoGold Resources Inc. as independence is defined in Section 1.5 of NI 43101.
8. I have had no previous involvement with the Los Ricos South project.
9. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have
been prepared in compliance with that Instrument. As of the effective date of the Technical
Report, to the best of my knowledge, information and belief, the sections of the Technical Report
for which I am responsible contain all scientific and technical information that is required to be
disclosed to make those sections of the Technical Report not misleading.
Dated: February 28th, 2025
“Signed and sealed”
Jonathan Cooper, M.SC., P.Eng
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 524 of 577
CERTIFICATE OF QUALIFIED PERSON
JAMES SMITH, P.ENG.
I, James Smith, state that:
(a) I am a Senior Geotechnical Engineer at:
WSP Canada Inc.
6925 Century Ave, Suite 600
Mississauga, Ontario
L5N 7K2 Canada
(b) This certificate applies to the technical report titled “Feasibility Study and Updated Mineral Resource
Estimate of the Los Ricos South Project, Jalisco, Mexico”; with an effective date of: January 14, 2025 (the
“Technical Report”).
(c) I am a “qualified person” for the purposes of National Instrument 43-101 (“NI 43-101”). My
qualifications as a qualified person are as follows. I am a graduate of University of Waterloo with a
B.ASc. in Geological Engineering from 2012, I am a member in good standing of the Professional
Engineers of Ontario (#100523579). My relevant experience after graduation, for the purpose of the
Technical Report, includes over 12 years of experience in geotechnical engineering in the areas of mine
stability, influences of hydrogeological conditions on rock mass stability, and open pit geotechnical
engineering of mineral projects nationally and internationally in a variety of commodities through 12
years of consulting experience with a strong focus on gold and base metals related projects.
(d) I did not complete a personal inspection of the property described in the Technical Report.
(e) I am responsible for portions of Items 1.24.3, 15.4.2.1, 16.1.2.1, 26.3, and 26.4 of the Technical Report.
(f) I am independent of the issuer as described in section 1.5 of NI 43-101.
(g) I have had no prior involvement with the Property that is the subject of this Technical Report.
(h) I have read NI 43-101 and the parts of the Technical Report for which I am responsible have been
prepared in compliance with NI 43-101; and
(i) At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the
parts of the Technical Report for which I am responsible, contain(s) all scientific and technical
information that is required to be disclosed to make the Technical Report not misleading.
Dated at Mississauga, Ontario this February 28, 2025.
Signed by James Smith
James Smith; P.Eng.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 525 of 577
APPENDIX A
SURFACE DRILL HOLE PLAN
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APPENDIX B
3-D DOMAINS
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APPENDIX C
AGEQ BLOCK MODEL CROSS-SECTIONS AND PLANS
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APPENDIX D
CLASSIFICATION BLOCK MODEL CROSS-SECTIONS AND PLANS
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APPENDIX E
OPTIMIZED PIT SHELLS
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APPENDIX F
DRILL HOLE LOCATION DATA
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGAG-20-001
LRGAG-20-002
LRGAG-20-003
LRGAG-20-004
LRGAG-20-005
LRGAG-20-006
LRGAG-21-007
LRGAG-21-008
LRGAG-21-009
LRGAG-21-010
LRGAG-21-011
LRGAG-21-012
LRGAG-21-013
LRGAG-21-014
LRGAG-21-015
LRGAG-21-016
LRGAG-21-017
LRGAG-21-018
LRGAG-22-019
LRGAG-22-020
LRGAG-22-021
LRGAG-22-022
LRGAG-22-023
LRGAG-22-024
LRGAG-22-025
LRGAG-22-026
LRGAG-22-027
LRGAG-22-028
LRGAG-22-029
LRGAG-22-030
LRGAG-22-031
610,102
610,102
610,176
610,175
610,165
610,164
610,058
610,172
610,170
610,142
610,141
610,056
610,055
610,055
610,059
610,058
610,058
610,058
610,143
609,704
610,026
610,096
610,142
610,096
609,703
609,806
610,026
610,062
610,058
610,061
610,061
2,328,248
2,328,219
2,328,321
2,328,321
2,328,340
2,328,339
2,328,297
2,328,402
2,328,401
2,328,443
2,328,442
2,328,370
2,328,369
2,328,369
2,328,297
2,328,296
2,328,345
2,328,344
2,328,421
2,328,800
2,328,548
2,328,608
2,328,420
2,328,607
2,328,799
2,328,557
2,328,548
2,328,514
2,328,443
2,328,513
2,328,513
Elevation
(m)
1,295
1,306
1,306
1,306
1,304
1,303
1,284
1,313
1,312
1,311
1,311
1,282
1,282
1,282
1,284
1,284
1,283
1,283
1,308
1,235
1,271
1,301
1,307
1,300
1,234
1,234
1,271
1,272
1,278
1,272
1,272
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
0
50
0
50
50
0
50
0
50
50
0
50
0
50
50
50
50
50
50
50
50
50
50
50
50
50
50
0
Dip
(°)
-65
-65
-45
-90
-45
-90
-75
-50
-90
-50
-90
-75
-50
-90
-50
-90
-50
-75
-50
-45
-50
-45
-85
-70
-90
-45
-70
-53
-50
-70
-90
Length
(m)
237
218
76
143
70
140
292
89
140
103
132
231
191
321
235
257
227
254
109
114
216
157
150
173
141
306
218
210
167
266
322
Page 554 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGAG-22-032
LRGAG-22-033
LRGAG-22-034
LRGAG-22-035
LRGAG-22-036
LRGAG-22-037
LRGAG-22-038
LRGAG-22-039
LRGAG-22-040
LRGAG-22-041
LRGAG-22-042
LRGAG-22-043
LRGAG-22-044
LRGAG-22-045
LRGAG-22-046
LRGAG-22-047
LRGAG-22-048
LRGAG-22-049
LRGAG-22-050
LRGAG-22-051
LRGAG-22-052
LRGAG-22-053
LRGAG-22-054
LRGAG-22-055
LRGAG-22-056
LRGAG-22-057
LRGAG-22-058
LRGAG-22-059
LRGAG-22-060
LRGAG-22-061
LRGAG-22-062
LRGAG-22-063
LRGAG-22-064
LRGAG-22-065
LRGAG-22-066
LRGAG-22-067
LRGAG-22-068
LRGAG-22-069
610,059
610,057
609,857
610,058
610,026
610,090
609,977
610,087
610,086
609,977
610,179
610,100
609,977
610,177
610,065
609,706
610,064
609,970
610,196
610,195
609,705
610,200
610,165
610,237
610,256
609,705
610,200
610,000
610,216
609,655
610,205
610,177
609,653
610,176
609,632
609,999
610,118
609,630
2,328,444
2,328,344
2,328,604
2,328,443
2,328,548
2,328,247
2,328,575
2,328,271
2,328,270
2,328,575
2,328,608
2,328,212
2,328,575
2,328,669
2,328,643
2,328,729
2,328,642
2,328,427
2,328,659
2,328,658
2,328,729
2,328,599
2,328,692
2,328,581
2,328,511
2,328,729
2,328,598
2,328,160
2,328,343
2,328,760
2,328,368
2,328,379
2,328,759
2,328,378
2,328,800
2,328,159
2,328,660
2,328,799
Elevation
(m)
1,278
1,283
1,244
1,278
1,271
1,295
1,265
1,289
1,289
1,265
1,338
1,312
1,265
1,351
1,302
1,218
1,302
1,252
1,362
1,362
1,218
1,337
1,353
1,347
1,383
1,218
1,337
1,318
1,320
1,210
1,318
1,311
1,210
1,311
1,204
1,318
1,325
1,204
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
0
50
0
0
0
50
50
0
50
50
0
0
50
50
50
50
50
50
0
50
50
50
50
50
0
0
50
50
50
50
50
0
0
50
0
50
0
Dip
(°)
-70
-90
-45
-90
-90
-90
-50
-65
-90
-72
-50
-90
-90
-45
-45
-45
-80
-60
-45
-90
-70
-45
-45
-45
-45
-90
-90
-70
-45
-65
-45
-45
-90
-90
-45
-90
-45
-90
Length
(m)
159
295
271
216
171
213
139
219
227
164
99
257
249
84
144
145
102
144
73
81
167
66
56
44
105
197
72
303
51
126
72
86
175
113
110
427
72
141
Page 555 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGAG-22-070
LRGAG-22-071
LRGAG-22-072
LRGAG-22-073
LRGAG-22-074
LRGAG-22-075
LRGAG-22-076
LRGAG-22-077
LRGAG-22-078
LRGAG-22-079
LRGAG-22-080
LRGAG-22-081
LRGAG-22-082
LRGAG-22-083
LRGAG-22-084
LRGAG-22-085
LRGAG-22-086
LRGAG-22-087
LRGAG-22-088
LRGAG-22-089
LRGAG-22-090
LRGAG-22-091
LRGAG-22-092
LRGAG-22-093
LRGAG-22-094
LRGAG-22-095
LRGAG-22-096
LRGAG-22-097
LRGAG-22-098
LRGAG-22-099
LRGAG-22-100
LRGAG-22-101
LRGAG-22-102
LRGAG-22-103
LRGAG-22-104
LRGAG-22-105
LRGAG-22-106
LRGAG-22-107
610,104
609,627
610,082
609,625
610,050
609,555
609,916
609,957
609,554
609,449
609,969
609,449
609,915
609,969
609,612
609,958
609,611
609,958
609,404
609,403
609,966
609,947
609,808
609,808
610,142
609,963
609,913
610,139
609,963
609,965
609,962
609,964
609,942
609,823
609,822
609,942
610,034
609,820
2,328,675
2,328,823
2,328,687
2,328,822
2,328,702
2,328,880
2,328,450
2,328,623
2,328,879
2,328,961
2,328,426
2,328,960
2,328,449
2,328,426
2,328,849
2,328,623
2,328,848
2,328,624
2,329,006
2,329,005
2,328,360
2,328,639
2,328,145
2,328,145
2,328,156
2,328,104
2,328,448
2,328,150
2,328,103
2,328,359
2,328,103
2,328,357
2,328,282
2,328,439
2,328,438
2,328,282
2,328,128
2,328,437
Elevation
(m)
1,326
1,199
1,325
1,199
1,325
1,180
1,243
1,268
1,179
1,164
1,252
1,164
1,243
1,252
1,196
1,268
1,195
1,268
1,160
1,160
1,252
1,268
1,297
1,297
1,330
1,325
1,243
1,329
1,325
1,252
1,325
1,252
1,272
1,231
1,231
1,272
1,317
1,231
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
0
50
50
50
0
0
50
50
0
50
0
50
50
0
50
50
0
50
0
50
50
50
50
0
0
50
50
0
0
50
50
50
0
50
0
Dip
(°)
-45
-45
-45
-90
-45
-45
-55
-90
-90
-45
-75
-90
-75
-90
-45
-70
-90
-45
-45
-90
-60
-90
-45
-55
-70
-60
-90
-90
-75
-75
-90
-90
-67
-50
-65
-90
-55
-90
Length
(m)
80
100
102
170
102
80
236
231
104
65
239
103
235
297
93
167
131
132
61
68
270
218
443
443
255
323
348
246
377
312
446
327
273
306
351
386
274
386
Page 556 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGAG-22-108
LRGAG-22-109
LRGAG-22-110
LRGAG-22-111
LRGAG-22-112
LRGAG-22-113
LRGAG-22-114
LRGAG-22-115
LRGAG-22-116
LRGAG-22-117
LRGAG-22-118
LRGAG-22-119
LRGAG-22-120
LRGAG-22-121
LRGAG-22-122
LRGAG-22-123
LRGAG-22-124
LRGAG-22-125
LRGAG-22-126
LRGAG-22-127
LRGAG-22-128
LRGAG-22-129
LRGAG-22-130
LRGAG-22-131
LRGAG-22-132
LRGAG-22-133
LRGAG-22-134
LRGAG-22-135
LRGAG-22-136
LRGAG-22-137
LRGAG-22-138
LRGAG-22-139
LRGAG-22-140
LRGAG-22-141
LRGAG-22-142
LRGAG-22-143
LRGAG-22-144
LRGAG-22-145
610,126
610,126
610,062
610,126
609,931
610,060
609,404
610,035
610,068
610,067
610,066
610,050
610,014
610,141
609,993
610,141
610,057
609,956
610,056
609,975
609,974
610,045
610,073
610,005
610,090
609,991
610,103
609,989
610,091
610,009
610,055
610,015
610,199
610,055
610,015
610,014
610,278
610,099
2,328,072
2,328,072
2,328,411
2,328,072
2,328,050
2,328,410
2,328,744
2,328,259
2,328,476
2,328,475
2,328,475
2,328,304
2,328,144
2,328,156
2,328,421
2,328,185
2,328,341
2,328,624
2,328,341
2,328,600
2,328,600
2,328,566
2,328,489
2,328,596
2,328,471
2,328,422
2,328,383
2,328,393
2,328,421
2,328,481
2,328,369
2,328,570
2,328,136
2,328,367
2,328,569
2,328,569
2,328,104
2,328,217
Elevation
(m)
1,310
1,310
1,283
1,310
1,326
1,282
1,247
1,306
1,277
1,277
1,277
1,283
1,319
1,329
1,259
1,325
1,282
1,267
1,282
1,268
1,268
1,279
1,276
1,275
1,286
1,259
1,291
1,263
1,291
1,258
1,282
1,273
1,353
1,282
1,273
1,273
1,323
1,307
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
50
50
0
50
50
50
50
0
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
Dip
(°)
-45
-65
-60
-85
-55
-90
-45
-63
-45
-70
-90
-63
-68
-56
-77
-60
-61
-78
-82
-48
-72
-46
-64
-49
-54
-65
-59
-64
-71
-74
-59
-48
-46
-82
-68
-89
-45
-79
Length
(m)
184
239
170
296
377
266
310
308
127
146
181
256
297
234
253
173
202
210
250
149
191
122
119
169
105
216
151
234
132
170
168
169
194
260
150
194
101
182
Page 557 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGAG-22-146
LRGAG-22-147
LRGAG-22-148
LRGAG-22-149
LRGAG-22-150
LRGAG-22-151
LRGAG-22-152
LRGAG-22-153
LRGAG-22-154
LRGAG-22-155
LRGAG-22-156
LRGAG-22-157
LRGAG-22-158
LRGAG-22-159
LRGAG-22-160
LRGAG-22-161
LRGAG-22-162
LRGAG-22-163
LRGAG-22-164
LRGAG-22-165
LRGAG-22-166
LRGAG-22-167
LRGAG-22-168
LRGAG-22-169
LRGAG-22-170
LRGAG-22-171
LRGAG-22-172
LRGAG-22-173
LRGAG-22-174
LRGAG-22-175
LRGAG-22-176
LRGAG-22-177
LRGG-22-208
LRGG-22-209
LRGG-22-210
LRGG-22-211
LRGG-22-212
LRGG-22-213
610,054
610,054
610,144
610,231
609,998
610,036
610,295
610,202
610,217
609,997
610,142
610,146
610,058
610,107
610,009
610,058
610,034
609,965
610,009
610,058
610,033
609,964
610,033
610,061
609,973
610,029
610,037
610,060
610,009
610,053
609,958
609,945
610,837
610,797
610,379
610,607
610,211
610,210
2,328,537
2,328,536
2,328,260
2,328,101
2,328,327
2,328,260
2,328,146
2,328,299
2,328,254
2,328,326
2,328,445
2,328,029
2,328,444
2,328,342
2,328,321
2,328,443
2,328,438
2,328,563
2,328,368
2,328,443
2,328,438
2,328,562
2,328,437
2,328,412
2,328,501
2,328,417
2,328,361
2,328,411
2,328,500
2,328,537
2,328,621
2,328,511
2,327,470
2,327,528
2,327,666
2,327,635
2,328,051
2,328,051
Elevation
(m)
1,274
1,274
1,312
1,334
1,264
1,305
1,344
1,319
1,331
1,265
1,312
1,291
1,278
1,293
1,268
1,278
1,270
1,261
1,269
1,278
1,269
1,260
1,269
1,282
1,254
1,270
1,279
1,282
1,261
1,274
1,268
1,250
1,252
1,255
1,261
1,278
1,311
1,312
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
Dip
(°)
-62
-87
-69
-50
-72
-53
-51
-72
-72
-76
-71
-56
-60
-61
-75
-74
-67
-52
-75
-84
-74
-72
-80
-45
-66
-78
-80
-76
-73
-76
-75
-68
-50
-85
-65
-61
-56
-75
Length
(m)
129
155
175
160
271
258
115
110
128
304
105
193
126
177
294
156
162
165
279
187
176
187
195
159
181
227
260
199
162
135
187
199
101
135
260
158
220
187
Page 558 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGG-22-214
LRGG-22-215
LRGG-22-216
LRGG-22-217
LRGG-22-218
LRGG-22-219
LRGG-22-220
LRGG-22-221
LRGG-22-222
LRGG-22-224
LRGG-22-225
LRGG-22-226
LRGG-22-227
LRGG-22-228
LRGG-22-229
LRGG-22-230
LRGG-22-231
LRGG-22-232
LRGG-22-233
LRGG-22-235
LRGG-22-236
LRGG-22-237
LRGG-22-238
LRGG-22-239
LRGG-22-241
LRGG-22-242
LRGG-22-243
LRGG-22-244
LRGG-22-245
LRGG-22-246
LRGG-22-247
LRGG-22-248
LRGG-22-249
LRGG-22-250
LRGG-22-251
LRGG-22-252
LRGG-22-253
LRGG-22-254
610,473
610,254
610,251
610,478
610,698
610,307
610,401
610,220
610,289
610,410
610,220
610,306
610,440
610,432
610,243
610,337
610,312
610,384
610,396
610,427
610,327
610,218
610,311
610,360
610,277
610,266
610,484
610,240
610,276
610,392
610,256
610,278
610,589
610,411
610,207
610,578
610,156
610,419
2,327,777
2,327,822
2,328,176
2,327,617
2,327,577
2,327,836
2,327,885
2,327,768
2,327,820
2,327,788
2,327,768
2,327,803
2,327,718
2,327,584
2,327,712
2,327,788
2,328,038
2,327,771
2,327,552
2,327,746
2,328,009
2,327,887
2,327,773
2,327,817
2,328,012
2,327,866
2,327,564
2,327,843
2,328,011
2,327,710
2,327,791
2,327,846
2,327,583
2,327,725
2,327,821
2,327,607
2,327,906
2,327,707
Elevation
(m)
1,293
1,251
1,362
1,279
1,256
1,257
1,293
1,234
1,256
1,276
1,233
1,254
1,270
1,263
1,236
1,256
1,299
1,264
1,251
1,271
1,294
1,250
1,250
1,264
1,285
1,265
1,261
1,249
1,285
1,263
1,247
1,259
1,263
1,266
1,239
1,268
1,274
1,273
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
Dip
(°)
-56
-54
-50
-58
-45
-58
-46
-56
-57
-45
-70
-56
-56
-53
-56
-60
-44
-45
-55
-52
-45
-54
-60
-59
-45
-45
-51
-50
-85
-63
-57
-59
-51
-61
-52
-53
-50
-65
Length
(m)
177
197
132
224
130
162
125
261
198
138
320
185
177
270
164
190
77
150
305
137
57
184
212
142
87
180
228
175
112
226
230
171
137
201
216
175
247
213
Page 559 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGG-22-255
LRGG-22-256
LRGG-22-257
LRGG-22-258
LRGG-22-259
LRGG-22-260
LRGG-22-261
LRGG-22-262
LRGG-22-263
LRGG-22-264
LRGG-22-265
LRGG-22-266
LRGG-22-267
LRGG-22-268
LRGG-22-269
LRGG-22-270
LRGG-22-271
LRGG-22-272
LRGG-22-273
LRGG-22-274
LRGG-22-275
LRGG-22-276
LRGG-22-277
LRGG-22-278
LRGG-22-279
LRGG-22-280
LRGG-22-281
LRGG-22-282
LRGG-22-283
LRGG-23-284
LRGG-23-285
LRGG-23-286
LRGG-23-287
LRGG-23-288
LRGG-23-289
LRGG-23-290
LRGG-23-291
LRGG-23-292
610,283
610,533
610,730
610,556
610,156
610,684
610,812
610,684
610,183
610,651
610,586
610,316
610,250
610,310
610,806
610,823
610,859
610,768
610,282
610,764
610,246
610,524
610,524
610,523
610,591
610,324
610,614
610,423
610,665
610,272
610,722
610,571
610,300
610,272
610,699
610,248
610,300
610,428
2,327,782
2,327,604
2,327,536
2,327,585
2,327,906
2,327,505
2,327,478
2,327,505
2,327,997
2,327,508
2,327,550
2,328,106
2,328,049
2,328,071
2,327,506
2,327,456
2,327,485
2,327,467
2,328,044
2,327,510
2,328,081
2,327,627
2,327,627
2,327,626
2,327,520
2,328,087
2,327,542
2,327,672
2,327,547
2,327,935
2,327,532
2,327,568
2,327,962
2,327,935
2,327,512
2,328,048
2,327,928
2,327,742
Elevation
(m)
1,247
1,260
1,246
1,257
1,275
1,247
1,246
1,247
1,279
1,245
1,252
1,328
1,306
1,312
1,252
1,243
1,264
1,232
1,300
1,242
1,318
1,270
1,270
1,270
1,242
1,321
1,253
1,280
1,251
1,278
1,245
1,254
1,281
1,278
1,246
1,306
1,284
1,270
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
Dip
(°)
-59
-49
-46
-53
-68
-50
-56
-63
-66
-50
-46
-45
-45
-44
-59
-56
-51
-54
-45
-48
-45
-44
-57
-71
-61
-45
-53
-53
-53
-51
-49
-52
-59
-67
-49
-55
-64
-64
Length
(m)
237
164
77
185
260
167
85
189
198
182
205
101
101
99
90
86
65
112
111
107
114
193
171
179
237
84
177
191
156
129
125
183
104
164
153
104
149
160
Page 560 of 577
TABLE F-1
DRILL HOLE LOCATION DATA
Drill Hole ID
Easting
Northing
LRGG-23-293
LRGG-23-294
LRGG-23-295
LRGG-23-296
LRGG-23-297
LRGG-23-298
LRGG-23-299
LRGG-23-300
LRGG-23-301
LRGG-23-302
LRGG-23-303
LRGG-23-304
LRGG-23-305
LRGG-23-306
LRGG-23-307
LRGG-23-308
LRGG-23-309
LRGG-23-310
LRGG-23-311
LRGG-23-312
LRGG-23-313
LRGG-23-314
LRGG-23-315
LRGG-23-316
LRGMI-23-001
LRGMI-23-002
LRGTO-22-001
LRGTO-22-002
LRGTO-22-003
LRGTO-23-004
LRGTO-23-005
LRSGT-23-001
LRSGT-23-002
610,422
610,272
610,317
610,207
610,223
610,223
610,281
610,604
610,604
610,318
610,444
610,444
610,367
610,350
610,255
610,426
610,437
610,329
610,183
610,183
610,604
610,548
610,324
610,324
611,488
611,487
611,609
611,718
611,718
611,736
611,735
609,992
610,367
2,327,934
2,327,902
2,327,812
2,327,916
2,327,766
2,327,766
2,327,717
2,327,563
2,327,563
2,327,943
2,327,886
2,327,885
2,328,148
2,328,133
2,328,050
2,327,675
2,327,744
2,328,131
2,327,997
2,327,997
2,327,564
2,327,620
2,327,703
2,327,703
2,327,467
2,327,467
2,327,526
2,327,562
2,327,562
2,327,590
2,327,590
2,328,411
2,327,772
Elevation
(m)
1,312
1,278
1,256
1,251
1,233
1,233
1,237
1,258
1,258
1,286
1,303
1,303
1,350
1,345
1,306
1,282
1,274
1,339
1,280
1,280
1,259
1,268
1,246
1,246
1,386
1,386
1,399
1,384
1,384
1,372
1,372
1,260
1,261
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Azimuth
(°)
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
50
55
50
50
50
50
45
45
65
65
65
65
65
65
65
50
53
Dip
(°)
-60
-66
-56
-52
-61
-50
-65
-45
-49
-51
-45
-67
-45
-45
-42
-45
-47
-57
-59
-74
-46
-41
-47
-70
-45
-70
-63
-45
-55
-57
-71
-45
-51
Length
(m)
65
186
170
186
290
250
302
144
136
116
82
110
59
78
59
216
188
77
179
221
139
169
234
306
56
103
435
138
167
117
246
257
126
Page 561 of 577
APPENDIX G
DATA
DRILL HOLE SIGNIFICANT INTERSECTIONS, VEINS AND ASSAYS
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-20-001
LRGAG-20-002
LRGAG-20-003
LRGAG-20-004
LRGAG-20-005
LRGAG-20-006
LRGAG-21-007
LRGAG-21-008
LRGAG-21-009
LRGAG-21-010
LRGAG-21-011
LRGAG-21-012
LRGAG-21-013
LRGAG-21-014
Area /
Vein
Eagle
including
Eagle
Eagle
Eagle
including
Eagle
Eagle
including
Eagle
including
including
Eagle
Eagle
including
Eagle
including
Eagle
including
Eagle
including
including
including
including
Eagle
including
including
Eagle
including
including
From
(m)
138.2
167.3
160.5
47.9
94.8
97.8
49.6
74.7
83.6
148.5
155.5
188.6
24.7
35.6
42.8
16.8
38.0
52.1
64.7
111.3
140.0
146.8
150.5
172.9
90.0
108.6
114.2
144.3
173.1
198.8
To
(m)
170.9
168.9
177.2
51.3
113.8
100.0
54.9
107.6
85.6
191.0
156.5
190.2
26.7
73.8
49.8
51.5
39.0
91.3
66.2
179.7
141.0
147.5
154.5
174.5
121.7
109.5
115.0
216.9
176.2
210.7
Length1
(m)
32.7
1.6
16.7
3.4
19.0
2.2
5.4
26.9
2.0
42.5
1.0
1.6
2.0
38.2
7.1
34.7
1.0
39.2
1.5
66.1
1.0
0.8
4.1
1.7
31.7
0.9
0.8
72.6
3.1
11.9
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
1.11
19.09
0.24
0.08
0.45
1.66
0.68
0.49
1.60
1.64
7.42
6.83
0.58
0.43
1.30
0.56
7.88
0.68
2.87
1.29
14.60
14.95
7.36
1.84
0.65
1.30
2.49
5.13
4.86
28.58
Ag
(g/t)
35.1
122.6
51.4
21.3
57.0
186.4
51.9
57.0
197.6
137.5
493.0
1,269.5
93.4
34.6
123.8
47.9
263.0
58.1
133.7
55.4
1,080.0
87.0
80.8
127.7
64.5
301.0
274.0
76.0
659.6
115.9
AuEq2
(g/t)
1.58
20.72
0.93
0.360
1.21
4.14
1.37
1.25
4.23
3.47
13.99
23.76
1.82
0.89
2.95
1.20
11.39
1.45
4.65
2.03
29.00
16.11
8.44
3.54
1.51
5.31
6.14
6.14
13.66
30.13
AgEq2
(g/t)
118.7
1,554.0
69.4
27.0
90.7
310.6
102.6
93.8
317.5
260.6
1,049.5
1,781.8
136.7
66.7
221.3
90.1
854.0
108.9
348.7
152.0
2,175.0
1,208.3
632.8
265.5
113.5
398.5
460.8
460.6
1,024.1
2,259.8
Page 562 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-21-015
LRGAG-21-016
LRGAG-21-017
LRGAG-21-018
LRGAG-22-019
LRGAG-22-020
LRGAG-22-021
LRGAG-22-022
LRGAG-22-023
LRGAG-22-024
LRGAG-22-025
LRGAG-22-026
LRGAG-22-027
LRGAG-22-028
LRGAG-22-029
LRGAG-22-030
Area /
Vein
including
including
including
Eagle
including
and
Eagle
including
Eagle
including
Eagle
including
including
and
Eagle
and
Eagle
Eagle
including
including
Eagle
Eagle
including
including
including
Eagle
Eagle
Eagle
Eagle
including
Eagle
Eagle
and
including
including
Eagle
including
including
From
(m)
208.6
208.6
210.0
146.5
146.5
177.8
175.8
180.3
121.7
125.7
127.6
156.9
156.9
177.9
40.2
50.6
33.9
72.0
88.4
88.4
40.0
41.5
57.5
60.3
60.3
85.5
45.0
187.2
101.7
108.7
44.8
59.0
74.0
75.5
75.5
42.9
51.9
54.1
To
(m)
210.7
209.3
210.7
163.9
147.2
179.3
182.6
181.3
145.1
126.3
163.0
159.3
157.6
178.6
41.3
53.2
35.3
94.5
90.0
89.2
59.5
96.6
65.6
62.6
61.5
86.1
51.0
188.2
110.7
109.7
48.8
63.5
94.6
77.5
76.5
79.5
61.0
54.8
Length1
(m)
2.1
0.7
0.7
17.4
0.7
1.4
6.8
0.9
23.4
0.6
33.6
2.4
0.8
0.7
1.1
2.6
1.4
22.5
1.6
0.8
19.5
55.1
8.1
2.3
1.2
0.6
6.0
1.0
9.0
1.0
4.0
4.5
20.6
2.0
1.0
36.7
9.1
0.7
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
146.03
266.00
168.50
1.02
11.75
0.99
0.88
3.42
2.25
25.40
1.57
10.51
15.00
36.20
1.17
1.19
0.07
2.38
24.60
27.40
0.09
0.82
2.12
4.76
4.89
3.48
0.10
0.39
1.54
5.79
0.66
1.20
1.52
8.00
14.15
1.77
6.00
14.00
Ag
(g/t)
176.8
319.0
140.0
36.6
284.0
69.1
82.4
308.0
133.0
437.0
85.5
436.5
1,155.0
1,500.0
28.5
35.9
26.6
24.4
49.0
40.8
9.8
99.8
519.9
1,471.5
2,650.0
6.0
13.1
111.0
62.5
33.2
63.7
60.3
217.5
1,865.5
3,520.0
156.2
553.3
1,260.0
AuEq2
(g/t)
148.39
270.25
170.37
1.51
15.54
1.92
1.98
7.53
4.02
31.23
2.71
16.33
30.40
56.20
1.55
1.67
0.42
2.70
25.25
27.94
0.22
2.15
9.06
24.39
40.22
3.56
0.28
1.87
2.37
6.23
1.51
2.01
4.42
32.87
61.08
3.85
13.38
30.80
AgEq2
(g/t)
11,128.9
20,269.0
12,777.5
113.1
1,165.2
143.7
148.1
564.5
301.7
2,342.0
203.5
1,225.0
2,280.0
4,215.0
116.5
125.3
31.7
202.7
1,894.0
2,095.8
16.8
161.2
679.3
1,828.9
3,016.7
267.0
20.7
140.3
177.7
467.5
112.9
150.6
331.3
2,465.5
4,581.3
289.0
1,003.2
2,310.0
Page 563 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-031
LRGAG-22-032
LRGAG-22-033
LRGAG-22-034
LRGAG-22-035
LRGAG-22-036
LRGAG-22-037
LRGAG-22-038
LRGAG-22-039
LRGAG-22-040
LRGAG-22-041
Area /
Vein
Eagle
including
including
including
including
Eagle
including
including
including
Eagle
including
Eagle
Eagle
including
including
including
including
Eagle
including
including
including
Eagle
including
including
including
Eagle
including
including
including
also
including
including
Eagle
including
including
Eagle
Eagle
including
From
(m)
58.0
78.0
80.0
80.0
84.0
48.1
85.8
88.5
88.5
190.9
215.5
130.5
69.0
129.0
130.4
130.4
131.0
87.0
109.0
119.5
119.5
122.2
146.4
149.9
150.9
88.0
106.9
108.4
108.4
To
(m)
126.0
113.9
95.0
86.0
86.0
113.8
113.8
92.9
89.1
222.0
219.3
134.0
179.6
140.5
133.8
131.9
131.9
128.0
128.0
120.9
120.3
167.8
155.0
151.6
151.6
124.0
116.0
109.6
109.0
Length1
(m)
68.0
35.9
15.0
6.0
2.0
65.8
28.1
4.3
0.6
31.1
3.8
3.5
110.6
11.5
3.5
1.5
0.9
41.0
19.0
1.5
0.8
45.7
8.7
1.7
0.7
36.0
9.2
1.3
0.7
Au
(g/t)
4.25
7.83
15.61
20.96
46.80
1.36
2.96
7.17
25.00
0.66
1.92
0.57
1.64
11.31
30.24
63.24
104.50
1.26
2.11
9.00
12.55
0.82
3.44
12.95
26.00
1.30
4.81
6.76
7.70
Ag
(g/t)
109.4
184.9
273.3
472.8
272.5
107.7
216.6
1,012.8
3,020.0
64.6
170.6
9.6
264.6
2,198.3
6,392.5
12,729.3
13,742.5
83.9
143.8
509.3
782.0
69.9
200.5
181.7
115.0
75.7
256.1
1,271.6
2,030.0
AuEq2
(g/t)
5.71
10.29
19.26
27.26
50.43
2.79
5.85
20.68
65.27
1.53
4.19
0.70
5.17
40.62
115.48
232.96
287.73
2.38
4.02
15.79
22.98
1.76
6.11
15.37
27.53
2.31
8.23
23.72
34.77
AgEq2
(g/t)
428.4
771.8
1,444.4
2,044.5
3,782.5
209.6
439.0
1,550.7
4,895.0
114.4
314.5
52.2
387.5
3,046.7
8,660.8
17,472.0
21,580.0
178.5
301.8
1,184.3
1,723.3
131.7
458.3
1,152.6
2,065.0
173.0
616.9
1,778.9
2,607.5
112.6
115.0
2.4
12.64
163.6
14.82
1,111.7
114.2
149.7
156.8
158.0
152.6
109.5
134.0
115.0
169.0
164.4
163.4
154.4
145.5
143.5
0.8
19.4
7.6
5.4
1.8
36.0
9.5
31.60
0.93
1.87
2.37
0.39
1.46
4.55
70.0
65.8
145.0
168.9
65.8
134.5
389.9
32.53
1.81
3.81
4.62
1.27
3.26
9.75
2,440.0
135.7
285.6
346.6
95.1
244.3
731.1
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 564 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-042
LRGAG-22-043
LRGAG-22-044
LRGAG-22-045
LRGAG-22-046
LRGAG-22-047
LRGAG-22-048
LRGAG-22-049
LRGAG-22-050
LRGAG-22-051
LRGAG-22-052
LRGAG-22-053
LRGAG-22-054
LRGAG-22-055
LRGAG-22-056
LRGAG-22-057
LRGAG-22-058
LRGAG-22-059
LRGAG-22-060
LRGAG-22-061
LRGAG-22-062
Area /
Vein
including
also
including
including
Eagle
Eagle
and
including
including
including
Eagle
including
including
Eagle
including
Eagle
including
Eagle
Eagle
Eagle
Eagle
Eagle
Eagle
Eagle
Eagle
including
Eagle
Eagle
Eagle
Eagle
Eagle
including
including
Eagle
Eagle
including
Eagle
including
From
(m)
134.0
To
(m)
134.8
Length1
(m)
0.8
Au
(g/t)
6.01
Ag
(g/t)
577.0
AuEq2
(g/t)
13.70
AgEq2
(g/t)
1,027.7
137.0
143.5
6.6
5.62
453.6
11.67
875.3
141.0
142.2
35.03
2,627.5
146.0
166.0
176.5
182.0
182.0
138.0
138.8
142.2
22.0
28.0
64.0
67.0
94.9
72.0
139.5
26.0
42.0
154.2
186.7
185.8
184.5
182.7
171.6
147.0
142.8
37.0
29.0
70.4
69.8
96.4
84.0
141.0
28.0
43.1
1.72
4.20
8.44
21.21
45.60
1.50
3.92
18.21
0.88
5.55
2.19
3.27
2.58
0.62
1.34
0.90
1.01
129.0
315.1
633.1
1,591.1
3,420.0
112.8
293.8
1,365.5
65.7
416.5
164.3
245.4
193.7
46.4
100.7
67.4
75.7
26.2
24.5
24.5
10.8
27.0
31.5
26.5
15.7
0.31
1.14
2.90
0.72
23.5
85.8
217.2
53.8
101.0
102.5
0.37
27.4
249.5
250.5
254.5
22.7
6.8
10.5
17.5
18.5
269.5
259.7
255.5
23.6
12.0
12.0
20.3
19.3
1.2
17.10
1,345.0
No significant mineralization
8.2
0.67
78.4
20.7
1.32
215.8
9.3
2.73
428.2
2.5
7.89
999.1
0.7
17.20
2,130.0
33.6
0.63
65.6
8.3
1.78
160.2
0.7
9.30
668.0
15.0
0.31
42.8
1.0
2.94
196.0
6.3
1.15
77.9
2.8
1.85
106.3
1.4
1.37
90.8
12.0
0.21
30.4
1.5
0.31
77.7
2.0
0.33
42.8
1.1
0.34
50.6
No significant mineralization
0.9
0.06
19.1
7.1
0.34
60.2
2.0
0.96
145.3
4.9
0.19
39.7
No significant mineralization
1.5
0.04
24.7
No significant mineralization
20.0
0.72
41.9
9.1
1.57
85.7
1.0
5.33
94.3
1.0
0.40
56.2
5.3
0.66
41.7
1.5
1.18
74.6
2.8
1.09
40.5
0.8
3.56
106.0
1.28
2.71
6.59
1.15
1.22
2.17
1.63
4.97
96.0
203.1
494.1
86.4
91.2
162.7
122.0
373.0
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 565 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-063
LRGAG-22-064
LRGAG-22-065
LRGAG-22-066
LRGAG-22-067
LRGAG-22-068
LRGAG-22-069
LRGAG-22-070
LRGAG-22-071
LRGAG-22-072
LRGAG-22-073
LRGAG-22-074
LRGAG-22-075
LRGAG-22-076
LRGAG-22-077
LRGAG-22-078
LRGAG-22-079
LRGAG-22-080
LRGAG-22-081
LRGAG-22-082
LRGAG-22-083
Area /
Vein
and
Eagle
including
including
Eagle
Eagle
including
Eagle
Eagle
Eagle
including
and
Eagle
Eagle
Eagle
Eagle
Eagle
including
Eagle
Eagle
Eagle
Eagle
including
including
Eagle
and
including
Eagle
including
including
including
Eagle
Eagle
and
including
Eagle
Eagle
Eagle
From
(m)
44.7
26.0
28.3
30.8
11.0
73.0
75.0
83.0
58.8
363.5
364.4
375.7
50.0
1.5
49.5
11.1
33.0
35.4
To
(m)
49.9
41.7
33.7
31.6
17.0
76.0
76.0
86.4
61.6
377.0
366.6
377.0
51.0
4.5
51.0
12.6
51.0
40.7
26.3
130.2
149.0
151.9
111.3
147.0
163.5
31.7
38.6
41.6
44.6
10.8
172.2
206.7
207.9
15.2
132.0
212.5
29.5
164.0
164.0
152.7
115.3
168.5
164.5
52.7
49.2
45.3
45.3
13.9
173.2
213.5
209.4
16.7
154.9
215.4
Length1
Au
Ag
AuEq2
(m)
(g/t)
(g/t)
(g/t)
5.2
0.87
20.0
1.14
15.7
1.21
46.7
1.83
5.4
2.99
108.5
4.43
0.8
16.35
339.0
20.87
6.0
0.28
41.4
0.83
3.0
5.00
19.7
5.26
1.0
13.15
17.0
13.38
3.4
1.02
63.9
1.87
2.8
0.06
14.2
0.24
13.5
0.15
67.8
1.06
2.2
0.53
127.5
2.23
1.4
0.30
204.6
3.03
1.0
0.61
55.4
1.35
3.0
0.55
50.4
1.22
1.5
0.47
54.7
1.20
1.5
1.11
95.0
2.38
18.0
0.33
48.2
0.98
5.4
0.97
118.3
2.54
No significant mineralization
No significant mineralization
3.3
0.34
90.1
1.54
33.8
0.76
45.5
1.37
15.0
1.30
79.5
2.36
0.8
14.70
640.0
23.23
4.0
0.78
65.7
1.66
21.5
0.66
67.7
1.56
1.0
4.95
460.0
11.08
21.0
0.43
79.4
1.49
10.6
0.70
122.8
2.34
3.8
1.12
196.1
3.73
0.8
1.96
512.0
8.79
3.2
0.47
78.6
1.52
1.0
1.30
36.4
1.78
6.8
2.43
23.6
2.74
1.5
10.60
25.0
10.93
1.5
0.69
59.2
1.48
22.9
0.27
16.8
0.50
2.8
1.94
42.1
2.51
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
AgEq2
(g/t)
85.3
137.4
332.5
1,565.3
62.5
394.4
1,003.3
140.3
18.3
79.2
167.3
227.0
101.2
91.4
90.0
178.3
73.1
190.8
115.6
102.5
176.8
1,742.5
124.2
117.1
831.3
111.6
175.6
279.8
659.0
113.8
133.5
205.6
820.0
111.2
37.4
187.9
Page 566 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-084
LRGAG-22-085
LRGAG-22-086
LRGAG-22-087
LRGAG-22-088
LRGAG-22-089
LRGAG-22-090
LRGAG-22-091
LRGAG-22-092
LRGAG-22-093
LRGAG-22-094
LRGAG-22-095
LRGAG-22-096
LRGAG-22-097
LRGAG-22-098
LRGAG-22-099
LRGAG-22-100
LRGAG-22-101
LRGAG-22-102
LRGAG-22-103
LRGAG-22-104
LRGAG-22-105
Area /
Vein
Eagle
Eagle
including
Eagle
Eagle
including
Eagle
Eagle
Eagle
including
Eagle
and
Eagle
Eagle
Eagle
including
including
Eagle
Eagle
Eagle
including
Eagle
including
Eagle
including
and
Eagle
including
including
Eagle
Eagle
and
Eagle
Eagle
including
Eagle
including
and
From
(m)
14.0
111.0
120.0
1.5
94.5
94.5
To
(m)
48.5
128.5
121.5
5.8
103.5
97.5
28.8
148.5
203.0
114.9
163.1
312.0
383.9
116.8
118.5
122.6
254.3
33.8
216.1
212.1
124.5
165.9
313.4
388.2
135.0
128.5
123.6
257.3
148.9
148.9
318.6
324.2
256.0
256.0
272.6
393.8
394.8
396.8
196.1
203.6
218.4
233.0
182.0
182.0
309.0
313.0
335.4
153.0
150.0
329.6
326.6
259.4
257.3
275.0
406.3
398.8
397.8
198.0
205.2
219.4
236.0
184.0
183.0
318.5
315.5
349.4
Length1
Au
Ag
AuEq2
(m)
(g/t)
(g/t)
(g/t)
34.5
0.06
17.0
0.29
17.5
0.54
29.8
0.94
1.5
4.77
92.0
6.00
4.3
0.19
44.2
0.78
9.0
0.87
38.5
1.39
3.0
1.77
24.7
2.10
No significant mineralization
5.0
0.23
32.0
0.66
66.8
0.54
33.3
0.98
9.1
1.89
96.3
3.17
9.7
0.37
55.3
1.11
2.8
0.57
50.2
1.24
1.4
4.42
12.7
4.59
4.3
0.38
27.4
0.74
18.2
1.37
146.0
3.32
10.0
2.39
235.0
5.53
1.0
7.51
448.0
13.48
3.1
0.52
58.6
1.30
No significant mineralization
4.2
3.20
169.4
5.46
1.2
10.75
521.0
17.70
11.0
0.69
73.2
1.67
2.4
2.88
224.8
5.87
3.4
2.05
54.7
2.78
1.3
5.23
129.8
6.96
2.4
1.11
30.7
1.52
12.5
0.52
59.5
1.32
4.0
1.35
148.3
3.33
1.0
2.95
349.0
7.60
1.9
1.00
11.3
1.15
1.6
2.39
67.9
3.29
1.0
2.36
3.4
2.41
2.9
0.38
40.4
0.92
2.0
4.71
12.6
4.87
1.0
9.04
18.0
9.28
9.5
0.46
55.7
1.20
2.5
1.12
117.5
2.69
14.0
0.91
35.7
1.38
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
AgEq2
(g/t)
21.6
70.5
449.8
58.7
103.9
157.2
49.6
73.5
237.7
83.2
92.8
344.2
55.8
248.9
414.5
1,011.3
97.8
409.6
1,327.2
125.3
440.6
208.6
522.1
113.8
98.8
249.7
570.3
86.6
247.0
180.4
68.9
365.5
696.0
90.3
201.7
103.8
Page 567 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-106
LRGAG-22-107
LRGAG-22-108
LRGAG-22-109
LRGAG-22-110
LRGAG-22-111
LRGAG-22-112
LRGAG-22-113
LRGAG-22-114
LRGAG-22-115
LRGAG-22-116
LRGAG-22-117
LRGAG-22-118
LRGAG-22-119
LRGAG-22-120
Area /
Vein
Eagle
including
Eagle
Eagle
including
Eagle
Eagle
including
including
Eagle
Eagle
including
Eagle
including
including
and
including
including
Eagle
Eagle
including
including
Eagle
including
including
Eagle
including
including
Eagle
including
including
including
including
including
Eagle
including
Eagle
including
From
(m)
192.2
189.2
313.9
130.5
132.0
To
(m)
205.2
192.2
315.4
138.0
133.0
62.0
98.7
102.6
231.8
322.6
323.6
132.8
133.8
134.8
160.5
165.1
169.0
124.4
114.7
103.9
235.0
331.1
324.6
149.0
135.6
135.6
190.0
177.4
170.0
180.1
191.6
200.6
57.6
57.6
58.6
64.5
65.5
65.5
95.6
95.6
104.6
107.6
107.6
108.6
162.0
186.0
229.0
239.2
211.6
202.6
201.6
72.3
62.6
59.6
89.5
74.5
66.5
150.6
120.6
119.6
114.6
110.6
110.6
174.5
187.4
242.3
240.3
Length1
Au
Ag
AuEq2
(m)
(g/t)
(g/t)
(g/t)
13.0
0.69
71.8
1.65
3.0
0.93
162.1
3.09
1.5
0.34
61.5
1.16
7.5
0.44
78.7
1.49
1.0
1.35
179.0
3.74
No significant mineralization
62.4
1.43
42.8
2.00
16.0
4.57
70.7
5.52
1.4
17.05
189.0
19.57
3.1
0.12
42.6
0.69
8.5
0.80
114.5
2.33
1.0
1.64
365.0
6.51
16.3
3.29
357.7
8.06
1.8
20.05
2,000.0
46.72
0.8
40.00
3,490.0
86.53
29.5
2.32
57.8
3.09
12.3
5.20
72.1
6.16
1.0
21.10
170.0
23.37
No significant mineralization
31.5
1.44
75.1
2.45
11.0
3.38
147.2
5.34
1.0
12.85
64.0
13.70
14.7
1.45
68.9
2.37
5.0
2.56
162.0
4.72
1.0
5.18
257.0
8.61
25.0
1.09
133.1
2.87
9.0
1.96
306.8
6.06
1.0
8.08
1,460.0
27.55
55.0
7.80
2,152.7
36.51
25.0
16.07
4,664.2
78.26
15.0
26.74
7,755.8 130.15
7.0
55.87 16,524.6 276.19
3.0
121.97 36,953.0 614.67
2.0
145.25 52,764.5 848.78
12.5
0.89
81.5
1.98
1.4
2.77
321.0
7.05
13.4
1.44
44.9
2.04
1.2
9.92
119.7
11.52
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
AgEq2
(g/t)
123.4
231.8
86.7
111.8
280.3
149.7
413.6
1,467.7
51.9
174.8
488.0
604.8
3,503.7
6,490.0
231.7
462.4
1,752.5
183.4
400.4
1,027.8
177.6
353.9
645.5
214.9
454.1
2,066.0
2,737.9
5,869.3
9,761.3
20,714.5
46,100.5
63,658.3
148.3
529.1
152.7
863.9
Page 568 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-121
LRGAG-22-122
LRGAG-22-123
LRGAG-22-124
LRGAG-22-125
LRGAG-22-126
LRGAG-22-127
LRGAG-22-128
LRGAG-22-129
LRGAG-22-130
LRGAG-22-131
LRGAG-22-132
LRGAG-22-133
LRGAG-22-134
LRGAG-22-135
LRGAG-22-136
Area /
Vein
Eagle
Eagle
including
Eagle
Eagle
and
including
Eagle
including
including
Eagle
including
including
and
Eagle
Eagle
including
including
Eagle
including
Eagle
including
including
including
including
Eagle
Eagle
and
Eagle
including
including
Eagle
including
and
including
Eagle
including
including
From
(m)
203.7
205.1
128.7
140.2
141.0
127.9
153.9
153.9
165.7
181.5
182.5
205.2
112.8
103.5
125.0
125.0
56.6
68.5
50.5
50.5
51.3
91.5
93.3
50.8
151.4
173.6
66.6
78.9
80.7
133.9
151.8
188.4
195.8
47.0
78.1
78.1
Length1
Au
Ag
AuEq2
(m)
(g/t)
(g/t)
(g/t)
No significant mineralization
216.0
12.4
1.54
28.2
1.91
207.3
2.2
6.05
45.6
6.65
Abandoned due to technical difficulties
130.7
2.0
2.07
95.1
3.33
142.6
2.4
7.05
56.3
7.80
141.9
0.8
14.35
66.0
15.23
170.2
42.3
2.14
127.9
3.84
161.3
7.3
9.66
485.8
16.13
156.7
2.8
15.35
865.7
26.89
189.0
23.3
7.07
86.2
8.22
184.2
2.7
45.01
106.2
46.42
183.3
0.8
75.60
148.0
77.57
210.6
5.4
0.64
120.1
2.24
119.0
6.3
0.65
54.9
1.38
143.4
39.9
1.28
86.6
2.43
136.4
11.4
3.48
209.7
6.28
126.0
1.0
11.40
659.0
20.19
72.9
16.3
1.09
28.0
1.46
71.0
2.5
4.59
47.4
5.22
68.7
18.2
1.72
118.4
3.30
52.0
1.5
10.28
785.2
20.75
52.0
0.7
19.20
1,275.0
36.20
97.2
5.8
0.75
127.9
2.46
95.0
1.8
1.34
318.0
5.58
54.7
3.9
0.99
55.5
1.73
153.7
2.3
0.34
91.2
1.56
174.7
1.1
0.89
104.5
2.29
103.5
36.9
1.13
69.2
2.05
95.3
16.5
2.28
134.0
4.06
81.9
1.3
8.55
823.9
19.53
138.0
4.2
0.80
96.3
2.08
152.6
0.8
3.96
448.0
9.93
199.2
10.8
1.32
42.5
1.89
198.1
2.3
5.10
37.2
5.60
100.0
53.0
1.03
64.8
1.90
98.2
20.2
2.16
55.6
2.90
78.8
0.7
9.54
490.0
16.07
To
(m)
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
AgEq2
(g/t)
143.6
499.0
250.1
584.9
1,142.3
288.3
1,210.0
2,016.8
616.5
3,481.6
5,818.0
168.1
103.4
182.5
470.7
1,514.0
109.8
391.4
247.5
1,556.4
2,715.0
184.2
418.4
129.5
117.0
171.6
153.8
304.6
1,465.1
156.2
745.0
141.7
419.8
142.2
217.6
1,205.5
Page 569 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-137
LRGAG-22-138
LRGAG-22-139
LRGAG-22-140
LRGAG-22-141
LRGAG-22-142
LRGAG-22-143
LRGAG-22-144
LRGAG-22-145
LRGAG-22-146
LRGAG-22-147
LRGAG-22-148
LRGAG-22-150
LRGAG-22-151
Area /
Vein
Eagle
including
and
Eagle
including
also
including
also
including
Eagle
including
including
Eagle
including
Eagle
including
and
including
Eagle
including
including
Eagle
including
Eagle
including
including
Eagle
including
including
Eagle
Eagle
including
including
Eagle
Eagle
including
including
Eagle
From
(m)
120.6
121.2
142.8
109.0
110.8
To
(m)
122.8
122.0
149.8
136.3
112.2
Length1
(m)
2.3
0.8
7.0
27.3
1.4
Au
(g/t)
0.64
1.52
0.99
3.06
15.64
Ag
(g/t)
157.5
390.0
77.8
193.4
907.7
AuEq2
(g/t)
2.74
6.72
2.03
5.64
27.74
AgEq2
(g/t)
205.5
504.0
152.1
423.1
2,080.6
121.9
122.7
0.8
14.00
19.0
14.25
1,069.0
124.3
125.3
1.0
5.43
563.0
12.94
970.3
76.3
93.0
93.5
157.6
159.2
156.6
177.1
192.0
196.4
99.0
99.6
99.6
123.9
126.7
35.1
35.1
37.4
112.5
128.4
132.9
75.5
53.2
85.6
96.0
138.3
209.4
211.1
215.1
194.4
97.4
94.0
94.0
163.7
160.8
185.0
178.0
203.4
197.7
107.0
101.1
100.2
133.4
128.2
43.5
40.4
39.7
154.6
142.5
133.4
78.7
99.8
99.0
96.8
143.3
220.7
217.4
216.2
201.7
21.1
1.1
0.5
6.1
1.6
28.4
0.9
11.4
1.3
8.0
1.5
0.7
9.6
1.5
8.5
5.4
2.4
42.2
14.1
0.5
3.2
46.6
13.4
0.8
5.0
11.3
6.3
1.1
7.3
1.77
20.04
29.10
0.77
2.32
3.82
43.30
2.10
9.43
2.79
3.01
4.00
2.29
7.41
2.80
4.38
8.19
1.94
3.70
16.15
1.99
1.08
3.24
24.30
1.13
6.96
12.02
32.10
1.77
43.7
71.2
67.0
135.8
309.3
113.9
99.0
188.6
709.8
138.9
393.5
704.0
198.9
708.0
472.7
727.3
1,400.5
271.7
699.7
5,740.0
57.8
40.8
76.0
45.0
137.1
64.9
100.0
157.0
74.9
2.35
20.99
29.99
2.58
6.44
5.34
44.62
4.61
18.89
4.64
8.26
13.39
4.94
16.85
9.11
14.07
26.87
5.56
13.03
92.68
2.76
1.63
4.25
24.90
2.96
7.83
13.35
34.19
2.77
176.6
1,574.0
2,249.5
193.7
483.2
400.1
3,346.5
345.7
1,416.7
348.2
619.5
1,004.0
370.8
1,263.8
682.9
1,055.6
2,015.1
417.2
977.2
6,951.3
206.7
122.0
319.1
1,867.5
221.7
587.0
1,001.6
2,564.5
207.8
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 570 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-152
LRGAG-22-153
LRGAG-22-154
LRGAG-22-155
LRGAG-22-156
LRGAG-22-157
LRGAG-22-158
LRGAG-22-159
LRGAG-22-160
LRGAG-22-161
LRGAG-22-162
LRGAG-22-163
LRGAG-22-164
LRGAG-22-165
LRGAG-22-166
Area /
Vein
including
Eagle
including
Eagle
Eagle
including
Eagle
including
and
including
Eagle
including
Eagle
Eagle
including
including
including
Eagle
and
and
including
Eagle
including
including
Eagle
including
including
Eagle
including
Eagle
including
Eagle
including
including
including
Eagle
including
and
From
(m)
195.0
73.2
83.3
65.1
88.2
88.9
205.0
205.0
229.3
231.6
38.4
43.9
153.1
78.2
79.2
80.0
99.1
173.9
200.4
214.9
219.7
91.4
93.4
94.8
100.7
105.7
107.7
117.5
125.9
199.0
214.3
110.9
117.7
119.6
121.8
118.7
119.2
141.5
To
(m)
196.3
85.3
85.3
66.6
92.8
89.9
211.0
205.6
233.6
232.6
47.3
45.1
155.4
102.8
88.0
80.7
100.1
175.4
202.5
227.7
220.5
124.4
103.0
96.1
131.7
113.8
108.4
132.5
129.7
221.1
217.5
160.9
125.5
122.5
122.5
127.8
120.6
151.6
Length1
(m)
1.3
12.1
2.0
1.5
4.6
1.0
6.0
0.6
4.3
1.0
9.0
1.2
2.3
24.6
8.8
0.8
1.0
1.5
2.2
12.9
0.8
33.0
9.6
1.3
31.0
8.1
0.7
15.0
3.8
22.1
3.2
50.0
7.8
2.9
0.8
9.1
1.3
10.1
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
3.48
0.55
1.35
0.35
0.52
1.16
1.93
9.20
2.79
8.19
1.86
7.59
0.46
2.19
3.87
13.05
2.71
2.09
5.77
2.13
14.30
2.51
5.16
19.58
6.07
19.38
144.00
0.78
1.40
3.21
20.21
4.61
22.90
59.98
201.00
3.08
8.12
2.32
Ag
(g/t)
154.2
70.6
124.3
143.5
113.2
202.5
30.4
74.7
64.8
52.1
161.0
697.9
86.9
137.9
340.5
2,660.0
91.0
176.9
50.1
70.2
255.0
338.9
945.4
3,665.4
1,063.9
3,921.7
36,319.0
92.3
200.3
32.6
86.8
780.2
4,616.9
11,858.2
31,747.0
199.5
609.9
114.6
AuEq2
(g/t)
5.54
1.49
3.01
2.26
2.03
3.86
2.34
10.20
3.65
8.88
4.01
16.89
1.62
4.03
8.41
48.52
3.92
4.45
6.44
3.07
17.70
7.03
17.76
68.46
20.26
71.67
628.25
2.01
4.07
3.64
21.37
15.01
84.45
218.09
624.29
5.75
16.25
3.85
AgEq2
(g/t)
415.3
111.8
225.5
169.8
152.2
289.5
175.3
764.7
273.8
666.4
300.9
1,266.8
121.6
302.0
630.9
3,638.8
294.3
334.0
483.0
230.3
1,327.5
527.0
1,332.0
5,134.2
1,519.2
5,375.2
47,119.0
150.5
305.5
273.0
1,602.9
1,125.9
6,334.1
16,357.0
46,822.0
430.9
1,219.1
288.9
Page 571 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGAG-22-167
LRGAG-22-168
LRGAG-22-169
LRGAG-22-170
LRGAG-22-171
LRGAG-22-172
LRGAG-22-173
LRGAG-22-174
LRGAG-22-175
LRGAG-22-176
LRGAG-22-177
LRGG-22-208
LRGG-22-209
LRGG-22-210
LRGG-22-211
LRGG-22-212
Area /
Vein
Eagle
including
Eagle
and
Eagle
including
including
Eagle
including
Eagle
including
Eagle
including
including
Eagle
including
including
including
Eagle
including
Eagle
including
Eagle
including
Eagle
including
including
Main area
including
Main area
including
including
Main area
including
Main area
including
including
Main area
From
(m)
135.3
139.0
99.8
132.9
55.4
84.7
85.9
139.7
142.2
172.6
180.4
155.3
165.6
166.1
107.8
108.7
108.7
142.6
95.9
99.3
72.5
75.2
131.5
141.9
133.7
154.8
160.7
37.5
43.2
58.3
86.0
87.0
216.0
222.1
94.4
98.3
100.1
91.8
To
(m)
150.8
139.8
102.0
138.0
99.9
99.2
87.1
145.1
144.1
182.5
182.5
209.1
185.0
171.3
148.2
112.7
110.1
143.6
102.8
100.3
83.7
76.4
147.3
143.5
165.3
164.6
163.0
47.7
45.0
91.5
89.0
88.0
229.0
225.0
118.5
101.9
101.9
114.8
Length1
(m)
15.5
0.8
2.3
5.1
44.5
14.5
1.3
5.4
1.9
9.9
2.2
53.8
19.4
5.2
40.5
4.1
1.4
1.0
6.9
1.0
11.2
1.2
15.9
1.6
31.7
9.8
2.3
10.3
1.9
33.2
3.0
1.0
13.0
2.9
21.7
3.7
1.8
23.0
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
2.16
18.60
2.56
2.12
1.51
3.78
18.38
0.87
2.23
2.00
5.43
3.27
8.09
13.72
3.67
5.35
8.90
11.80
1.23
3.45
1.55
4.11
2.54
9.03
1.21
3.03
10.15
1.17
3.50
3.15
30.19
73.10
0.95
2.69
1.77
7.54
10.02
0.38
Ag
(g/t)
102.3
76.0
75.8
226.3
52.3
47.3
105.9
99.3
230.1
78.5
107.0
123.8
193.1
288.8
281.6
1,267.0
1,878.9
1,430.0
197.3
454.0
50.2
55.4
119.8
627.5
145.3
392.8
1,357.2
313.0
976.2
276.5
2,587.7
5,620.0
99.5
184.9
117.1
348.1
391.0
71.2
AuEq2
(g/t)
3.53
19.61
3.57
5.14
2.21
4.41
19.80
2.19
5.29
3.04
6.86
4.92
10.67
17.57
7.43
22.24
33.95
30.87
3.86
9.50
2.22
4.85
4.14
17.40
3.15
8.27
28.25
5.35
16.52
6.84
64.69
148.03
2.28
5.15
3.33
12.18
15.23
1.33
AgEq2
(g/t)
264.6
1,471.0
267.6
385.3
165.5
331.1
1,484.7
164.2
397.1
228.1
514.6
369.3
800.1
1,317.7
556.9
1,668.0
2,546.5
2,315.0
289.9
712.8
166.6
363.6
310.5
1,304.8
236.0
619.9
2,118.4
401.0
1,238.7
512.8
4,851.7
11,102.5
171.1
386.6
249.5
913.5
1,142.5
100.0
Page 572 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGG-22-213
LRGG-22-214
LRGG-22-215
LRGG-22-216
LRGG-22-217
LRGG-22-218
LRGG-22-219
LRGG-22-220
LRGG-22-221
LRGG-22-222
LRGG-22-224
LRGG-22-225
LRGG-22-226
LRGG-22-227
LRGG-22-228
LRGG-22-229
LRGG-22-230
LRGG-22-231
LRGG-22-232
LRGG-22-233
LRGG-22-234
LRGG-22-235
LRGG-22-236
Area /
Vein
including
Main area
Main area
Main area
Main area
including
Main area
Main area
including
including
including
Main Area
Main Area
including
Main Area
including
including
Main Area
including
Main area
including
Main area
Main area
Main area
including
including
Main Area
Main Area
Main Area
Main Area
Main Area
including
Main Area
including
Main Area
Main Area
including
Main Area
From
(m)
110.5
125.3
117.5
141.9
99.0
102.1
179.8
57.0
57.0
60.5
61.2
126.3
63.8
70.7
194.9
202.0
203.9
127.7
134.3
97.6
99.8
254.4
136.5
124.0
130.6
132.0
235.7
To
(m)
114.8
131.5
128.8
145.0
105.7
102.8
198.9
74.1
63.9
63.0
62.2
128.8
79.4
74.1
224.2
206.4
204.9
136.3
134.9
102.8
102.8
260.2
137.9
153.0
137.0
132.7
251.5
30.0
112.6
113.4
280.2
281.1
34.3
112.5
118.5
27.8
32.0
122.8
114.2
299.0
282.0
50.3
126.4
122.0
38.7
Length1
Au
Ag
AuEq2
(m)
(g/t)
(g/t)
(g/t)
4.3
1.72
289.4
5.58
6.3
0.14
20.8
0.42
11.3
0.54
69.4
1.47
3.1
0.55
76.2
1.56
6.7
0.36
62.9
1.20
0.7
1.58
267.0
5.14
15.1
0.52
109.6
1.98
15.0
3.98
529.2
11.04
4.8
12.43
1,625.6
34.10
2.5
20.32
2,774.1
57.31
0.9
37.90
4,250.0
94.57
2.5
0.65
66.6
1.54
15.6
0.50
98.9
1.82
3.4
1.34
156.5
3.42
29.4
0.85
138.3
2.69
4.4
4.98
748.7
14.97
0.9
12.70
2,370.0
44.30
8.7
0.94
104.8
2.34
0.6
7.60
592.0
15.49
5.2
1.51
145.0
3.44
3.0
2.32
193.0
4.89
5.8
0.50
91.0
1.71
1.4
1.10
153.1
3.14
29.0
1.28
105.2
2.68
6.3
4.84
352.0
9.54
0.7
24.00
335.0
28.47
11.9
1.30
120.1
2.90
Hole abandoned above target
No significant mineralization
2.0
0.59
95.3
1.86
10.2
2.60
60.9
3.41
0.8
24.60
238.0
27.77
18.9
0.51
78.2
1.56
0.9
1.84
400.0
7.17
16.0
0.18
44.2
0.77
13.9
3.63
144.6
5.56
3.5
12.34
350.7
17.01
10.9
0.50
94.4
1.76
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
AgEq2
(g/t)
418.3
31.2
110.0
117.0
90.1
385.5
148.3
827.8
2,557.6
4,298.2
7,092.5
115.6
136.2
256.8
202.1
1,122.5
3,322.5
175.6
1,162.0
258.3
366.8
128.2
235.8
200.9
715.2
2,135.0
217.5
139.8
255.7
2,083.0
116.7
538.0
57.8
417.0
1,275.9
132.2
Page 573 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGG-22-237
LRGG-22-238
LRGG-22-239
LRGG-22-241
LRGG-22-242
LRGG-22-243
LRGG-22-245
LRGG-22-246
LRGG-22-247
LRGG-22-248
LRGG-22-249
LRGG-22-250
LRGG-22-251
LRGG-22-252
LRGG-22-253
LRGG-22-254
LRGG-22-255
LRGG-22-256
LRGG-22-257
LRGG-22-258
LRGG-22-260
LRGG-22-261
LRGG-22-262
Area /
Vein
including
Main Area
including
Main Area
Main Area
including
Main area
including
Main area
including
Main area
Main area
Main area
including
Main area
including
Main area
Main area
Main area
including
Main area
Main area
including
Main area
Main area
including
and
Main area
including
Main area
including
Main area
including
Main area
Main area
including
Main area
Main area
From
(m)
31.7
148.0
148.0
164.3
112.7
112.7
53.2
54.5
128.4
133.8
195.4
79.2
185.0
193.6
167.6
175.0
120.8
125.4
151.2
152.3
161.0
123.6
126.1
210.4
167.7
170.5
189.4
174.8
183.0
135.3
138.0
60.5
71.4
144.8
115.3
119.6
50.6
135.0
To
(m)
32.7
149.8
148.7
166.3
118.0
114.2
59.6
55.5
144.6
136.2
199.8
81.8
199.0
194.6
185.0
180.0
122.2
135.3
157.5
154.1
162.6
133.2
126.8
212.4
179.6
171.1
192.9
184.3
184.3
142.0
139.0
72.0
72.0
151.0
137.7
120.1
56.2
140.9
Length1
(m)
1.0
1.8
0.7
2.1
5.3
1.5
6.3
1.0
16.2
2.4
4.5
2.6
14.0
1.0
17.5
5.0
1.4
9.9
6.3
1.8
1.6
7.4
0.7
2.0
11.9
0.6
3.4
9.5
1.3
5.4
1.1
11.6
0.7
6.2
22.5
0.5
5.6
5.8
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
3.83
0.87
1.15
1.08
0.77
2.28
0.59
1.93
0.70
2.29
0.76
0.90
2.90
31.80
0.60
1.28
0.66
0.95
1.54
3.12
1.09
1.39
4.81
0.94
1.28
7.06
0.45
1.08
4.50
1.97
5.28
1.50
8.64
1.14
1.20
11.50
0.67
2.27
Ag
(g/t)
686.0
139.8
264.0
102.0
163.0
374.0
86.0
321.0
72.0
94.9
142.9
239.8
84.6
261.0
93.1
160.0
78.3
333.9
101.5
213.0
128.5
167.6
429.0
146.2
183.1
1,830.0
166.1
185.8
727.8
136.1
336.0
164.7
261.0
162.5
124.1
427.0
58.1
41.6
AuEq2
(g/t)
12.98
2.73
4.67
2.44
2.94
7.27
1.74
6.21
1.66
3.56
2.67
4.10
4.03
35.28
1.84
3.41
1.70
5.40
2.89
5.96
2.80
3.62
10.53
2.88
3.72
31.46
2.67
3.55
14.20
3.79
9.76
3.69
12.12
3.31
2.85
17.19
1.45
2.82
AgEq2
(g/t)
973.2
204.7
350.2
182.7
220.6
545.0
130.5
465.7
124.7
266.7
199.9
307.4
302.3
2,646.0
138.1
255.9
127.8
404.9
217.1
447.0
210.2
271.6
789.7
216.3
279.2
2,359.5
200.1
266.6
1,065.3
284.0
732.0
277.0
909.0
248.3
213.8
1,289.5
108.6
211.8
Page 574 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGG-22-263
LRGG-22-264
LRGG-22-265
LRGG-22-267
LRGG-22-268
LRGG-22-269
LRGG-22-271
LRGG-22-272
LRGG-22-273
LRGG-22-274
LRGG-22-275
LRGG-22-276
LRGG-22-277
LRGG-22-278
LRGG-22-279
LRGG-22-280
LRGG-22-281
LRGG-22-282
LRGG-22-283
LRGG-23-284
LRGG-23-285
Area /
Vein
including
Main area
including
Main area
including
including
Main area
including
including
and
Main area
including
Main area
Main area
including
Main area
Main area
Main area
Main area
Main area
Main area
including
Main area
Main area
Main area
Main area
including
including
Main area
including
Main area
including
Main area
including
Main area
including
Main area
and
From
(m)
135.0
150.2
153.0
131.6
149.5
153.0
134.7
136.0
137.1
151.9
57.0
66.0
17.7
42.1
50.2
39.5
79.9
40.2
55.6
67.5
135.0
137.5
138.0
170.2
179.5
1.2
10.1
10.1
135.8
136.6
175.1
186.8
106.1
106.1
91.1
94.6
68.7
90.2
To
(m)
136.0
155.6
153.5
159.0
156.5
154.7
141.8
139.0
138.0
156.2
73.5
68.5
20.6
59.4
51.0
43.3
84.4
49.5
66.0
71.8
148.7
140.0
142.2
176.3
185.5
16.7
12.4
10.7
147.5
138.0
190.8
188.8
118.0
107.1
96.3
96.3
74.8
92.6
Length1
(m)
1.0
5.4
0.5
24.9
7.0
1.7
7.1
3.0
0.9
4.3
16.5
2.5
2.9
17.3
0.9
3.8
4.5
9.3
8.3
4.3
13.7
2.5
4.2
6.1
6.0
14.0
2.3
0.6
11.7
1.4
13.0
1.9
12.0
1.1
5.3
1.7
6.1
2.4
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
7.91
1.50
0.87
1.01
2.46
8.78
1.25
2.69
5.70
0.86
1.32
6.44
3.54
2.53
26.70
2.19
1.54
0.74
0.63
0.51
1.48
4.26
1.45
1.02
1.44
1.77
10.30
22.70
0.93
5.36
0.99
2.59
0.67
5.06
1.62
2.98
1.01
1.84
Ag
(g/t)
28.2
85.7
472.0
104.5
247.5
746.8
416.6
905.0
2,300.0
215.8
153.2
571.5
264.6
159.8
1,215.0
197.0
73.5
92.7
100.6
170.4
213.4
439.7
254.4
76.7
199.2
573.7
3,308.2
7,580.0
387.3
2,780.4
216.0
474.0
80.2
169.0
144.6
180.9
173.4
366.0
AuEq2
(g/t)
8.29
2.64
7.16
2.40
5.76
18.74
6.81
14.76
36.37
3.73
3.36
14.06
7.07
4.66
42.90
4.82
2.52
1.98
1.97
2.78
4.32
10.13
4.85
2.04
4.09
9.42
54.41
123.77
6.10
42.43
3.87
8.91
1.74
7.31
3.55
5.39
3.32
6.72
AgEq2
(g/t)
621.5
198.3
537.3
180.3
431.8
1,405.5
510.7
1,106.7
2,727.5
280.0
251.9
1,054.5
530.1
349.5
3,217.5
361.5
188.7
148.3
148.1
208.7
324.1
759.5
363.4
153.2
306.9
706.2
4,080.8
9,282.5
457.2
3,182.6
290.0
668.3
130.7
548.5
265.9
404.3
249.3
503.9
Page 575 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGG-23-286
LRGG-23-287
LRGG-23-288
LRGG-23-289
LRGG-23-291
LRGG-23-292
LRGG-23-293
LRGG-23-294
LRGG-23-295
LRGG-23-296
LRGG-23-297
LRGG-23-298
LRGG-23-299
LRGG-23-300
LRGG-23-302
LRGG-23-303
LRGG-23-304
LRGG-23-305
Area /
Vein
Main area
including
Main area
including
and
Main area
including
Main area
including
Main area
including
including
Main area
including
including
Main area
including
Main area
including
Main area
and
Main area
Main area
including
Main area
including
Main area
including
Main area
Main area
including
including
including
Main area
including
Main area
including
Main area
From
(m)
142.4
144.4
62.5
62.5
79.2
114.6
120.2
101.3
111.5
87.0
113.8
116.7
141.0
141.0
144.2
8.1
26.2
145.2
145.2
121.9
137.3
147.6
221.6
228.5
196.2
199.1
233.5
242.4
135.4
52.8
67.0
72.0
73.0
33.3
40.5
53.9
59.0
20.3
To
(m)
165.5
146.0
72.6
69.3
80.2
121.3
121.3
129.4
112.5
121.8
120.8
117.7
152.8
145.3
145.3
40.7
27.3
161.4
146.3
123.1
140.5
150.2
232.2
230.0
204.5
201.8
244.8
243.4
138.4
83.0
76.7
73.8
73.8
43.4
42.0
63.8
60.5
33.0
Length1
(m)
23.1
1.7
10.1
6.8
1.1
6.7
1.1
28.1
1.1
34.8
7.0
1.0
11.8
4.3
1.1
32.7
1.1
16.3
1.1
1.2
3.1
2.6
10.6
1.6
8.3
2.7
11.3
1.0
3.1
30.2
9.7
1.8
0.8
10.1
1.5
10.0
1.5
12.7
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Au
(g/t)
0.80
9.03
1.48
2.02
1.55
0.45
0.36
0.71
3.38
1.37
4.68
19.30
0.91
2.33
3.24
1.19
3.33
1.02
5.53
1.30
0.43
0.75
1.15
3.06
0.83
1.71
1.36
3.92
1.23
1.68
4.54
18.99
30.10
0.53
0.99
1.82
4.34
0.51
Ag
(g/t)
143.3
1,352.8
81.9
97.3
459.0
111.1
222.0
128.7
1,080.0
132.4
440.4
1,935.0
224.1
563.1
1,295.0
124.4
518.0
69.8
125.0
91.0
85.7
62.4
189.9
414.4
121.7
222.2
223.7
662.0
112.5
277.7
769.1
3,484.3
3,210.0
91.7
150.0
181.2
633.0
111.8
AuEq2
(g/t)
2.71
27.07
2.58
3.31
7.67
1.93
3.32
2.42
17.78
3.14
10.55
45.10
3.89
9.84
20.51
2.85
10.24
1.95
7.20
2.51
1.57
1.58
3.68
8.59
2.45
4.68
4.35
12.75
2.73
5.39
14.79
65.44
72.90
1.75
2.99
4.24
12.78
2.00
AgEq2
(g/t)
203.0
2,030.0
193.3
248.6
575.3
144.8
249.0
181.9
1,333.5
235.3
791.1
3,382.5
292.1
737.7
1,538.0
213.4
767.7
146.6
539.8
188.1
117.7
118.6
276.4
644.2
183.6
350.7
325.9
956.0
204.5
404.0
1,109.5
4,908.2
5,467.5
131.3
224.3
318.1
958.5
150.2
Page 576 of 577
TABLE G-1
DRILL HOLE SIGNIFICANT INTERSECTIONS AND ASSAYS
Drill Hole ID
LRGG-23-306
LRGG-23-308
LRGG-23-309
LRGG-23-310
LRGG-23-311
LRGG-23-312
LRGG-23-313
LRGG-23-314
LRGG-23-315
LRGG-23-316
LRGMI-23-001
LRGTO-22-002
LRGTO-23-005
Area /
Vein
including
Main area
including
Main area
including
Main area
including
Main area
including
Main area
Main area
and
Main area
including
Main area
including
Main area
including
Main area
including
including
Main area
Main area
including
including
including
Main area
including
including
including
including
From
(m)
27.7
36.1
37.7
159.7
164.0
115.3
122.3
39.8
47.0
137.4
157.0
165.5
125.4
131.4
123.0
130.0
193.3
196.3
249.0
255.0
261.0
33.1
91.0
91.9
92.9
93.8
137.5
137.5
138.1
140.0
140.6
To
(m)
28.9
47.8
39.2
173.2
169.0
133.9
122.9
55.0
49.8
139.1
158.0
166.5
134.6
134.6
131.2
131.2
206.7
197.6
276.3
262.7
261.9
33.7
96.2
92.8
93.8
94.4
142.3
138.1
139.0
140.6
141.2
Length1
(m)
1.2
11.7
1.5
13.6
5.0
18.6
0.6
15.3
2.8
1.7
1.0
1.0
9.2
3.2
8.2
1.2
13.4
1.3
27.3
7.6
0.9
0.6
5.2
0.9
0.9
0.6
4.8
0.6
0.9
0.6
0.6
Au
(g/t)
1.69
4.99
30.50
1.36
2.78
1.50
12.05
0.50
1.53
0.42
0.80
2.14
3.23
4.64
0.58
0.78
0.66
4.87
1.27
4.43
11.80
0.43
2.90
5.95
2.38
3.17
1.92
0.92
2.94
1.81
2.71
Ag
(g/t)
487.0
604.8
3,550.0
157.9
246.9
89.4
144.0
74.4
249.4
97.6
204.0
158.0
234.8
460.8
98.9
291.0
56.6
172.5
395.6
1,337.6
7,230.0
42.2
49.42
37.7
24.7
62.5
173.5
108.0
626.0
5.2
17.8
AuEq2
(g/t)
8.18
13.05
77.83
3.46
6.08
2.69
13.97
1.49
4.85
1.72
3.52
4.25
6.36
10.79
1.90
4.66
1.42
7.18
6.54
22.26
108.20
1.00
3.56
6.45
2.71
4.00
4.23
2.36
11.29
1.87
2.95
AgEq2
(g/t)
613.8
979.0
5,837.5
259.6
455.8
201.8
1,047.8
112.0
364.0
129.0
264.0
318.5
477.2
808.9
142.6
349.5
106.2
538.1
490.6
1,669.5
8,115.0
74.7
266.7
484.0
203.2
300.2
317.7
177.0
846.5
140.6
221.0
Notes: 1 Not true width.
2
Converted using a silver to gold ratio of 75:1 at recoveries of 100%.
P&E Mining Consultants Inc.
GoGold Resources Inc., Los Ricos South Feasibility Study, Report No. 469
Page 577 of 577
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