Exhibit 96.1

MINERA PLATA REAL, S. DE R.L. DE CV

JOINT VENTURE OF SUNSHINE SILVER MINING & REFINING CORPORATION (51.5%) AND
DOWA METALS & MINING CO., LTD. (48.5%)

VALLE ESCONDIDO NO. 5500
COMPLEJO INDUSTRIAL EL SAUCITO

PUNTO ALTO E-2, SUITE 404 CP, 31125

CHIHUAHUA, MEXICO

Los Gatos Project

Chihuahua, Mexico

Project No.:  117-8302008

Effective Date:	July 1, 2020
Issue Date:	July 2020

Prepared by:	Guillermo Dante Ramírez-Rodríguez, PhD, MMSAQP
	Leonel Lopez, SME-QP
	Kira Johnson, MMSAQP
	Keith Thompson, C.P.G.
	Kenneth Smith, SME-QP
	Luis Quirindongo, SME-QP
	Max Johnson, P.E.

350 Indiana Street, Suite 500 | Golden, CO 80401

303.217.5700 | tetratech.com

TABLE OF CONTENTS

1.0 SUMMARY	1
1.1 Introduction	1
1.1.1 Property Description, Location and Infrastructure	1
1.2 History	2
1.3 Geology and Mineralization	3
1.4 Drilling	3
1.5 Mineral Resource Estimates	4
1.6 Mineable Reserve Estimate	6
1.7 Metallurgical Testing	7
1.8 Mining Methods	8
1.9 Recovery Methods	8
1.10 Infrastructure	11
1.11 Marketing Studies and Contracts	11
1.12 Environmental Studies, Permitting, and Social or Community Impact	12
1.13 Capital and Operating Costs	14
1.13.1 LOM Capital Costs	14
1.13.2 LOM Operating Costs	14
1.14 Economic Analysis	15
1.15 Conclusions and Recommendations	16
1.15.1 Geology and Resources	16
1.15.2 Mineral Reserves	17
1.15.3 Mine Planning	17
1.15.4 Mineral Processing and Metallurgy	18
1.15.5 Economics	18
2.0 INTRODUCTION	19
2.1 Terms of Reference	19
2.2 Scope of Work	19
2.3 Units of Measure	19
2.4 Detailed Personal Inspections	19
3.0 RELIANCE ON OTHER EXPERTS	20
4.0 PROPERTY DESCRIPTION AND LOCATION	21
4.1 Location	22
4.1.1 Mining Concession	22
4.1.2 Los Gatos and Paula Adorada Concessions	24
4.1.3 Internal Concessions not held by MPR	25
4.2 Surface Rights	26
4.3 Environmental Permitting	27
4.4 Environmental Liabilities	28

TETRA TECH July 2020

i

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	29
5.1 Accessibility	29
5.2 Climate, Vegetation, Soils, and Land Use	29
5.3 Personnel	30
5.4 Infrastructure	30
5.5 Population Centers	30
6.0 HISTORY	31
6.1 Historic Resource Estimates	31
7.0 GEOLOGICAL SETTING AND MINERALIZATION	32
7.1 Regional Geological Setting	32
7.2 Property Geological Setting	35
7.3 Mineralized Zones	38
8.0 DEPOSIT TYPES	42
9.0 EXPLORATION	45
9.1 Historic Exploration	45
9.2 Recent Exploration	45
9.3 Decline and Bulk Sample	47
10.0 DRILLING	51
11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY	54
11.1 Sample Preparation	54
11.2 Security	55
11.3 Analyses	55
11.4 Quality Assurance and Quality Control for Sample Analysis	57
11.4.1 Quality Control Sample Performance	57
12.0 DATA VERIFICATION	58
12.1 Check Sampling	58
12.2 Database Verification	60
12.3 Metallurgical Sampling	60
13.0 MINERAL PROCESSING AND METALLURGICAL TESTING	62
13.1 Samples	62
13.2 Mineralogy	63
13.3 Comminution	65
13.4 Flotation Testing	65
13.4.1 Effect of Primary Grind Size on Flotation	66
13.4.2 Effect of Depressant Addition on Lead and Silver Flotation	70
13.4.3 Effect of Longer Flotation and Higher Collector Addition on Pb/Ag Flotation	73
13.4.4 Effect of CuSO4 Dosage on Zinc Flotation and Pyrite Flotation	74
13.4.5 Effect of Na2SiO3 on Zinc Flotation	75
13.4.6 Effect of pH on Zinc Flotation	76
13.4.7 Zinc Oxide Mineral Flotation	77

ii

13.4.8 Locked Cycle Flotation Testing	78
13.4.9 Product Characterization	81
13.5 Variability Composite Flotation	83
13.5.1 Silver Flotation	84
13.5.2 Lead Flotation	84
13.5.3 Zinc Flotation	85
13.6 Cyanide Destruction	85
13.7 Fluorine Control	85
14.0 MINERAL RESOURCE ESTIMATES	89
14.1 Cerro Los Gatos Deposit	89
14.1.1 Input Data	92
14.1.2 Grade Capping	92
14.1.3 Compositing	95
14.1.4 Geologic Modeling	95
14.1.5 Vein Modeling	96
14.1.6 Estimation Methods and Parameters	103
14.1.7 Mineral Resource Classification	110
14.1.8 Dilution	111
14.1.9 Cutoff Grade and Reasonable Prospects for Economic Extraction	111
14.1.10 Statement of Mineral Resources	112
14.1.11 Model Verification	114
14.1.12 Relevant Factors	121
14.2 Amapola Zone	122
14.2.1 Amapola Drill Hole Database	122
14.2.2 Amapola Geologic Modeling	122
14.2.3 Amapola Assays and Composting	125
14.2.4 Amapola Variography and Search Orientation	126
14.2.5 Amapola Resources Estimation and Categorization	127
14.2.6 Amapola Specific Gravity	128
14.2.7 Cutoff Grade and Reasonable Prospects for Economic Extraction	129
14.2.8 Amapola Deposit Resource Estimate	130
14.3 Esther Zone	132
14.3.1 Esther Drill Hole Database	132
14.3.2 Esther Geologic Modeling	132
14.3.3 Esther Assays and Composting	135
14.3.4 Esther Variography and Search Orientation	136
14.3.5 Esther Resources Estimation and Categorization	136
14.3.6 Esther Specific Gravity	137
14.3.7 Cutoff Grade and Reasonable Prospects for Economic Extraction	138
14.3.8 Esther Deposit Resource Estimate	139
14.4 Relevant Factors	141
14.5 Conclusions and Recommendations	141
14.5.1 Geology and Resources	141
14.6 Recommendations	143
14.6.1 Standards	143
14.6.2 Blanks	144
14.6.3 Duplicates	145
14.6.4 Umpire Sampling	145

iii

15.0 MINERAL RESERVE ESTIMATES	146
15.1 Net Smelter Return	146
15.2 Ore Body Description	148
15.3 Mining Method	151
15.4 Dilution and Recovery Estimates	152
15.4.1 Dilution	152
15.4.2 Mining Recovery	154
15.4.3 Drift-and-Fill Recovery	155
15.5 Mineral Reserves	156
15.6 Conclusions and Recommendations	159
15.6.1 Conclusions	159
15.6.2 Recommendations	159
16.0 MINING METHODS	160
16.1 Geotechnical	160
16.1.1 Geomechanical Investigation	160
16.1.2 Analysis and Design	164
16.1.3 Stope Design Criteria	166
16.1.4 Development	168
16.2 Mine Access Design	172
16.2.1 Mine Access	172
16.2.2 Access Ramps	172
16.2.3 Mining Sublevels	173
16.2.4 Ventilation Raises	175
16.2.5 Secondary Egress	175
16.3 Mining Methods and Sequence	175
16.3.1 Drift-and-Fill Mining	175
16.3.2 Transverse Longhole Stoping	178
16.3.3 Longitudinal Longhole Stoping	183
16.3.4 Backfilling	184
16.3.5 Transverse Longhole Stopes	184
16.3.6 Longitudinal Stopes	185
16.3.7 Drift-and-Fill Stopes	185
16.3.8 Production Rate	185
16.3.9 Stope Design Parameters	186
16.4 Development and Production Schedules	187
16.4.1 Development Productivity Rates	187
16.4.2 Development Schedule	187
16.4.3 Production Planning Criteria	189
16.4.4 Production Schedule	189
16.5 Mine Equipment	191
16.5.1 Underground Mobile Equipment	191
16.5.2 Surface Mobile Equipment	194
16.5.3 Fixed Equipment	195
16.6 Ventilation	198
16.6.1 Ventilation Method and Design Criteria	199
16.6.2 Airflow Requirements	200
16.6.3 Mine Air Cooling	202

iv

16.7 Backfill	205
16.7.1 Distribution	206
16.8 Conclusions and Recommendations	206
16.8.1 Conclusions	206
16.8.2 Recommendations	207
17.0 RECOVERY METHODS	208
17.1 Primary Crushing	210
17.2 Crushed Ore Conveying, Transport and Storage	210
17.3 Grinding	210
17.4 Lead Flotation and Regrind	211
17.5 Zinc Flotation and Regrind	212
17.6 Lead Concentrate Dewatering	212
17.7 Zinc Concentrate Dewatering	213
17.8 Tailing Dewatering	213
17.9 Cyanide Destruction (SO2/Air Process)	213
17.10 Reagent	213
17.11 Water System	213
17.11.1 Fresh Water	213
17.11.2 Process Water	214
18.0 PROJECT INFRASTRUCTURE	215
18.1 Existing Infrastructure and Services	215
18.1.1 Location	215
18.1.2 Site Access Roads	215
18.1.3 Buildings	215
18.1.4 Communications	216
18.1.5 Personnel	217
18.1.6 Power Supply	217
18.1.7 Power Distribution	217
18.2 Site Development	219
18.2.1 Mine Surface Facilities	219
18.3 Water Source	220
18.3.1 Potable Water Supply	220
18.3.2 Raw Water Distribution System	220
18.3.3 Process Water Supply	221
18.3.4 Sewage Waste Water Treatment	221
18.4 Waste Disposal	221
18.5 Underground Infrastructure	221
18.5.1 Mine Dewatering	221
18.5.2 Materials Handling	224
18.5.3 Electrical Power and Distribution	224
18.5.4 Compressed Air	225
18.5.5 Service Water	225
18.5.6 Service Bay	227
18.5.7 Fuels and Lubricants	229
18.5.8 Communications	229

v

18.5.9 Refuge Stations	230
18.5.10 Sanitary Facilities	230
19.0 MARKET STUDIES AND CONTRACTS	231
20.0 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT	234
20.1 Introduction	234
20.2 Regulations	234
20.3 Baseline Studies	235
20.4 Environmental Setting	237
20.4.1 Climate and Precipitation	237
20.4.2 Air Quality and Atmosphere	238
20.4.3 Geology	238
20.4.4 Soils	238
20.4.5 Regional and Site Hydrology	238
20.4.6 Flora	240
20.4.7 Fauna	240
20.4.8 Socioeconomic	240
20.5 Identification, Description, and Assessment of Environmental Impacts	241
20.6 Preventive and Mitigation Measures for the Environmental Impacts	243
20.7 Environmental Forecast	244
20.7.1 Atmosphere	244
20.7.2 Soil	244
20.7.3 Water	244
20.7.4 Geology and Geomorphology	244
20.7.5 Flora	244
20.7.6 Fauna	244
20.7.7 Ecosystem	244
20.7.8 Socioeconomic	245
20.8 Closure Plan	245
20.8.1 Tailings Storage Facility	247
20.8.2 Temporary Rock Storage	248
20.8.3 Underground Mine and Openings	248
20.8.4 Process and Ancillary Facilities	248
20.8.5 Borrow Areas and Exploration Camp	248
20.8.6 Seepage, Stormwater, Sedimentation and Cooling Basins	248
20.8.7 Road and Conveyor Corridors	248
20.8.8 Utility and Pipeline Corridor	249
20.8.9 Well Abandonment	249
21.0 CAPITAL AND OPERATING COSTS	250
21.1 Principal Assumptions	250
21.2 Life of Mine	252
21.2.1 Underground Mining	252
21.2.2 Processing	253
21.3 Capital Costs	253
21.4 Operating Costs	254
21.5 Taxes and Royalties	255
21.5.1 Royalties	255

vi

21.5.2 Taxes	255
22.0 ECONOMIC ANALYSIS	256
22.1 Net Smelting Return	256
22.2 Economic Results	257
22.3 Sensitivity	258
23.0 ADJACENT PROPERTIES	261
24.0 OTHER RELEVANT DATA AND INFORMATION	262
24.1 Hydrogeology and Mine Dewatering	262
24.1.1 Hydrogeology	262
24.1.2 Groundwater Modeling	266
24.1.3 Production Dewatering	279
24.1.4 Conclusions and Recommendations	281
24.2 Geochemistry	282
24.2.1 Waste Rock Characterization	282
24.2.2 Tailings Characterization	282
24.3 Surface Water Hydrology	283
24.3.1 Methodology	284
24.4 Site-Wide Water Balance	298
24.4.1 Introduction and Discussion	298
24.4.2 Model Components	299
24.4.3 Results	302
24.5 Tailings Management	304
24.5.1 Topography	307
24.5.2 Site Seismicity	307
24.5.3 Surface Water Hydrology	308
24.5.4 Site Investigation & Site Conditions	309
24.5.5 Borrow Material	309
24.5.6 Tailings Dam Design	309
24.5.7 Staged Construction	310
24.5.8 Liner Design	310
24.5.9 Tailings Delivery System	311
24.5.10 TSF Water Collection System	311
24.5.11 Underdrain System	311
24.5.12 Blanket Drain System	311
24.6 Surface Water Management	311
24.6.1 Monitoring	312
25.0 INTERPRETATION AND CONCLUSIONS	316
25.1 Geology and Resources	316
25.1.1 Data Verification	316
25.2 Mineral Reserve and Mine Plan	317
25.2.1 Mineral Reserve	317
25.2.2 Geotechnical Conclusion	317
25.2.3 Mine Plan	317
25.3 Mineral Processing and Metallurgy	317
25.4 Infrastructure	318

vii

25.5 Environmental and Social Impacts	318
25.6 Mine Reclamation	320
25.7 Economic Analysis	320
25.8 Groundwater Hydrology/Dewatering	320
25.9 Surface Hydrology	321
25.10 Tailings Management	321
26.0 RECOMMENDATIONS	322
26.1 Engineering, Procurement, and Construction Management	322
26.2 Geology and Resources	322
26.2.1 Standards	322
26.2.2 Blanks	323
26.2.3 Duplicates	324
26.2.4 Umpire Sampling	324
26.3 Mineral Reserve and Mine Planning	324
26.3.1 Mineral Reserve Estimate	324
26.3.2 Mining	325
26.3.3 Mine Planning	325
26.4 Metallurgy and Recovery Methods	325
26.5 Environmental, Permitting, and Reclamation	326
26.6 Geochemistry	326
26.7 Tailings Management	326
26.8 Surface Water Hydrology	327
26.9 Groundwater Hydrology/Dewatering	327
26.10 Water Balance	327
26.11 Reclamation and Closure	328
26.12 Additional Work	329
27.0 REFERENCES	330
28.0 DATE AND SIGNATURE PAGE	335

viii

LIST OF TABLES

Table 1-1: Drill Hole Count by Purpose	4
Table 1-2: Mineral Resource Estimate	4
Table 1-3: Estimated Mineral Resources Indicated and Inferred for Amapola and Esther	6
Table 1-4: Mineral Reserve	7
Table 1-5: Payable Metal Price Forecasts for Los Gatos	12
Table 1-6: LOM Capital Costs	14
Table 1-7: LOM Operating Costs	14
Table 1-8: TEM Results	15
Table 4-1: Los Gatos Project Titled Mining Concessions	22
Table 4-2: Internal Concessions	25
Table 9-1: Mineralized Grade Intercepts	47
Table 9-2: Bulk Sample and Block Model Comparison	49
Table 10-1: Drill Hole Count by Purpose	51
Table 12-1: Collar Verification by Handheld GPS	60
Table 13-1: Weights of Master Composite and Variability Composites	62
Table 13-2: Head Assays of Flotation Composites	63
Table 13-3: Comminution Test Results Summary	65
Table 13-4: Summary Conditions of LCTs	78
Table 13-5: Summary Results of Locked Cycle Test	79
Table 13-6: ICPSCAN and WRA Analysis	81
Table 13-7: Summary Conditions of CND Test	87
Table 13-8: Summary Results of CND Test	88
Table 14-1: Mineral Resource Estimate	89
Table 14-2: Drill Hole and Sample Count	92
Table 14-3: Grade Capping Statistics	93
Table 14-4: Vein Transformations	101
Table 14-5: Specific Gravity Measurements and Regressions	102
Table 14-6: Modeled Variograms for Ag	103
Table 14-7: Modeled Variograms for Pb, Zn, Au, Cu	104
Table 14-8: Block Model Setup Parameters	107
Table 14-9: Pass Parameters	107
Table 14-10: Mineral Resource Classification	110
Table 14-11: Cutoff Grade Parameters	111
Table 14-12: Mineral Resource Estimate	112
Table 14-13: Mineral Resource Estimate by Fault Block	113
Table 14-14: Alternative Estimate Difference Percent, Measured and Indicated Blocks	119
Table 14-15: Assay Statistics In Modeled Veins — Amapola	125
Table 14-16: Assay Capping — Amapola	126
Table 14-17: Cutoff Grade Parameters	129

ix

Table 14-18: Estimated Indicated Resources — Amapola	130
Table 14-19: Estimated Inferred Resources — Amapola	131
Table 14-20: Assay Statistics in Modeled Veins — Esther	135
Table 14-21: Assay Capping — Esther	135
Table 14-22: Cutoff Grade Parameters	138
Table 14-23: Estimated Indicated Resources — Esther	139
9Table 14-24: Estimated Inferred Resources — Esther	140
Table 15-1: Dilution by Mining Method	154
Table 15-2: Recovery by Mining Method	155
Table 15-3: Mineral Reserve	156
Table 15-4: Resources exclusive of Reserves	158
Table 16-1: Rock Mass Rating	160
Table 16-2: Point Load Test Compilation Results for All Mineralized Zones	161
Table 16-3: Joint Set Mean Orientation	161
Table 16-4: Comparison of UCS Lab Results to Point Load Testing Results	162
Table 16-5: Adjusted Rock Mass Properties used for Numerical Modeling per Lab Results	163
Table 16-6: In Situ Stress Tensors	163
Table 16-7: Parameters used to Establish Hr and N	167
Table 16-8: Ground Support Class Description for 5 m Wide × 5.5 m High Drift	169
Table 16-9: Support System for Spans Greater than 5 m	168
Table 16-10: Q Value for Los Gatos Selected Rock Unit	170
Table 16-11: Stope Design Parameters	186
Table 16-12: LOM Development Schedule	188
Table 16-13: Productivity Rates	189
Table 16-14: LOM Production Schedule	190
Table 16-15: Underground Mobile Equipment — Maximum Units on Site	191
Table 16-16: Surface Mobile Equipment — Maximum Units on Site	194
Table 16-17: Ventilation Design Criteria	199
Table 16-18: Full Production Airflow Determinations	202
Table 16-19: Cooling Design Criteria and Assumptions	203
Table 18-1: Mine Power Requirements by Area	217
Table 18-2: Mine Compressed Air Demand	225
Table 18-3: Underground Process Water Requirements	226
Table 19-1: Indicative Smelting and Refining Terms for Zinc Concentrate	231
Table 19-2: Indicative Smelting and Refining Terms for Lead Concentrate	232
Table 19-3: Payable Metal Price Forecasts for Los Gatos	232
Table 20-1: Summary of Closure and Reclamation Costs	249
Table 21-1: TEM Principal Assumptions	250
Table 21-2: LOM Production from July 1, 2020	252
Table 21-3: LOM Mill Feed from July 1, 2019	253
Table 21-4: LOM Capital Costs	254
Table 21-5: Mining Operating Costs	254

x

Table 22-1: NSR	256
Table 22-2: TEM Results	257
Table 24-1: Hydrogeologic Units Represented in the Model	265
Table 24-2: Faults Represented as Barriers to Flow in the Model	266
Table 24-3: Summary of Dewatering Simulation Results	270
Table 24-4: Water pumped from dewatering wells January-May 2020	280
Table 24-5: CONAGUA Design Storm Requirements for Impoundments	284
Table 24-6: TR-55 Hydrologic Soil Groups	287
Table 24-7: NOM-011-CNA-2000 Hydrologic Soil Type and K Values as a Function of Soil Use	287
Table 24-8: Basin Areas	289
Table 24-9: Lag Times	290
Table 24-10: Design Storms	290
Table 24-11: Minimum, Average and Maximum Total Monthly Rainfall for Los Gatos Site	291
Table 24-12: Average Annual Runoff Volume	293
Table 24-13: Source of Basis of Design	294
Table 24-14: Estimated Stormwater Conveyance Structure Dimensions, Capacity Analysis	295
Table 24-15: Channel Slopes and Velocities, Stability Analysis	295
Table 24-16: Freeboard Criteria	296
Table 24-17: Model Parameters and Source	298
Table 24-18: TSF Embankment Raise Schedule	302
Table 24-19: TSF Capacity	309
Table 24-20: TSF Spillway & Channel Dimensions	311
LIST OF FIGURES
Figure 1-1: Vein Solids 3D View	5
Figure 1-2: Process Plant Overall Flowsheet	10
Figure 4-1: Los Gatos General Location Map	21
Figure 4-2: Concessions Map for the Los Gatos Concessions	24
Figure 4-3: Surface Rights and Exploration Permissions (MPR)	26
Figure 7-1: Geologic Provinces Map Showing the SMO, Tectono-Stratigraphic Terranes (SGM, Campa and Coney, 1987; USGS, Hammarstrom et al, 2010)	32
Figure 7-2: Regional Geologic Map (SGM, 1:250,000 Original Scale) Property Structural Setting	34
Figure 7-3: Stratigraphic column of Los Gatos	36
Figure 7-4: Geological Map of the Cerro Los Gatos Deposit Area	38
Figure 7-5: Cross-Section 29 AgEq	40
Figure 8-1: Epithermal Textures in Drill Core	43
Figure 8-2: Idealized Section of a Bonanza Epithermal Deposit (Buchanan L.J., 1981)	44
Figure 9-1: Decline Entrance (Mine Portal)	48
Figure 9-2: Decline and Bulk Sample Area Cross-Section Looking Northwest (+/- 200 m)	48
Figure 9-3: 3D Clipped View of Bulk Sample and Block Model AgEq Grades	50

xi

Figure 10-1: Drill Hole Collar Monument	52
Figure 10-2: Drill Hole Location Map	53
Figure 11-1: Sampled Drill Core	54
Figure 11-2: Sample Analysis Flow Diagram	56
Figure 12-1: Check Sample Scatter Plot	58
Figure 12-2: Check Sample Scatter Plot	59
Figure 13-1: Cumulative Grain Size Distribution (Graph from SGS)	64
Figure 13-2: Effect of Grind Size on Lead Flotation in Rougher Tests	66
Figure 13-3: Effect of Grind Size on Silver Flotation in Rougher Tests	67
Figure 13-4: Primary Grind Size on Lead and Silver Flotation in Cleaner Tests	68
Figure 13-5: Effect of Grind Size on Zinc Flotation in Rougher Tests	69
Figure 13-6: Effect of Depressant on Lead Rougher Flotation	70
Figure 13-7: Effect of Depressant on Silver Flotation	71
Figure 13-8: Effect of ZnCN on Lead and Silver Flotation	72
Figure 13-9: Lead Cleaner Tests — Grade and Recovery Curves	73
Figure 13-10: Effect of CuSO4 Dosage on Zinc Rougher Flotation	74
Figure 13-11: Effect of Sodium Silicate on Zinc Flotation	75
Figure 13-12: Effect of pH on Zinc Rougher Flotation	76
Figure 13-13: Effect of pH on Zinc Cleaner Flotation	77
Figure 13-14: Silver Flotation Results of Variability Composites	83
Figure 13-15: Lead Flotation Results of Variability Composites	83
Figure 13-16: Zn Flotation Results of Variability Composites	84
Figure 14-1: Vein Solids 3D View	90
Figure 14-2: Estimated AgEq Block Grades within Grade-Shell 3D View	91
Figure 14-3: Block Resource Classification 3D View	91
Figure 14-4: Drill Hole Plan and Cross-Section Index	92
Figure 14-5: Histogram for Capping Analysis Ag	93
Figure 14-6: Histogram for Capping Analysis Pb	93
Figure 14-7: Histogram for Capping Analysis Zn	94
Figure 14-8: Histogram for Capping Analysis Au	95
Figure 14-9: Histogram for Capping Analysis Cu	95
Figure 14-10: Geologic Model Solids 3D View	95
Figure 14-11: Geologic Model Cross-Section 27 Looking NW	96
Figure 14-12: Vein Solids 3D View	97
Figure 14-13: Long-Section Vein Intervals Assignments	98
Figure 14-14: Long-Section AgEq Grade of Vein Intervals	98
Figure 14-15: Long-Section Ag Grade of Vein Intervals within Grade-Shell	98
Figure 14-16: Long-Section Pb Grade of Vein Intervals within Grade-Shell	99
Figure 14-17: Long-Section Zn Grade of Vein Intervals within Grade-Shell	99
Figure 14-18: Long-Section Au Grade of Vein Intervals within Grade-Shell	99
Figure 14-19: Long-Section Cu Grade of Vein Intervals within Grade-Shell	100

xii

Figure 14-20: Long-Section Thickness of Vein Intervals within Grade-Shell	100
Figure 14-21: SG Regression within Vein Interpretation	102
Figure 14-22: Experimental and Modeled Variography Ag	104
Figure 14-23: Experimental and Modeled Variography Pb	105
Figure 14-24: Experimental and Modeled Variography Zn	105
Figure 14-25: Experimental and Modeled Variography Au	106
Figure 14-26: Experimental and Modeled Variography Cu	106
Figure 14-27: Estimated AgEq Block Grades within Grade-Shell 3D View	108
Figure 14-28: Estimated Ag Block Grades within Grade-Shell 3D View	108
Figure 14-29: Estimated Pb Block Grades within Grade-Shell 3D View	108
Figure 14-30: Estimated Zn Block Grades within Grade-Shell 3D View	109
Figure 14-31: Estimated Au Block Grades within Grade-Shell 3D View	109
Figure 14-32: Estimated Cu Block Grades within Grade-Shell 3D View	109
Figure 14-33: Mineral Resource Classification 3D View	110
Figure 14-34: Stacked Histogram of Nearest Composite Sample to Classified Blocks	111
Figure 14-35: Grade Tonnage Curve Measured and Indicated Resources	113
Figure 14-36: Grade Tonnage Curve Inferred Resources	114
Figure 14-37: Cross-Section 13 AgEq	115
Figure 14-38: Cross-Section 13 Ag	116
Figure 14-39: Cross-Section 13 Pb	117
Figure 14-40: Cross-Section 13 Zn	118
Figure 14-41: Alternative Estimate Comparison Tonnage Curve	119
Figure 14-42: Alternative Estimate Comparison AgEq Grade Curve	120
Figure 14-43: Alternative Estimate Comparison Ag Grade Curve	120
Figure 14-44: Alternative Estimate Comparison Pb Grade Curve	120
Figure 14-45: Alternative Estimate Comparison Zn Grade Curve	121
Figure 14-46: Plan view of the Amapola Veins	124
Figure 14-47: Amapola Vein Wireframe Solids Looking Northwest	125
Figure 14-48: Amapola Vein Block Classification Looking West	128
Figure 14-49: Grade Tonnage Curve Indicated Amapola Deposit	130
Figure 14-50: Grade Tonnage Curve Inferred Amapola Deposit	131
Figure 14-51: Esther Zone Wireframes Looking Northwest	134
Figure 14-52: Esther Zone Wireframes Looking North	135
Figure 14-53: Esther Zone Block Classification Looking North	137
Figure 14-54: Grade Tonnage Curve Indicated Esther Deposit	139
Figure 14-55: Grade Tonnage Curve Inferred Esther Deposit	140
Figure 14-56: Additional drilling and total AgEq oz per estimate	142
Figure 15-1: NSR Value / Tonnage Plot	147
Figure 15-2: Orthogonal View showing Flattening Veins	148
Figure 15-3: Plan View of Mineralized Veins	149
Figure 15-4: Block Model Section	150

xiii

Figure 15-5: NWZ Longhole Stope Dilution	152
Figure 15-6: Drift-and-Fill Overbreak	153
Figure 15-7: Drift-and-Fill Dilution	153
Figure 15-8: Isometric View of Stope	154
Figure 15-9: Recovery Profile (Drift-and-Fill)	155
Figure 15-10: LOM plan showing mining, development, and the 2019 Resource model, with Measured blocks in red and Indicated blocks in green	157
Figure 16-1: Most Favorable Azimuth for Underground Excavation	162
Figure 16-2: Geometry Used in the 3D Model Looking towards the Hanging Wall	164
Figure 16-3: Raise Wall Stability	171
Figure 16-4: Raise Face Stability	172
Figure 16-5: Typical Level Development Northwest Zone	173
Figure 16-6: Typical Level Development CZ and SEZs	174
Figure 16-7: Typical Waste Development Drill Patterns	174
Figure 16-8: Typical Drift-and-Fill Stope Schematic	176
Figure 16-9 : Typical Ore Development Drill Patterns	177
Figure 16-10: Typical Transverse Stope Schematic View	179
Figure 16-11: Transverse Stope Dimensions and Drilling	180
Figure 16-12: Typical Slot Raise	180
Figure 16-13: Slot Ring Drilling	181
Figure 16-14: Production Ring and Slash Drilling	181
Figure 16-15: Slot Complete and Undercut Started	182
Figure 16-16: Production Blasting Retreat	182
Figure 16-17: Typical Paste Backfilling	183
Figure 16-18: Longitudinal Stope Dimensions and Drilling	184
Figure 16-19: Design Span for Drift-and-Fill	186
Figure 16-20: Early Production Ventilation Arrangement	200
Figure 16-21: Full Production Ventilation Arrangement	201
Figure 16-22: Heat Model Equipment Locations	204
Figure 16-23: Heat Simulation Results	205
Figure 17-1: Process Plant Overall Flowsheet	209
Figure 18-1: Borehole Sump	222
Figure 18-2: Level Dewatering Sump	223
Figure 18-3: Main Dewatering Sump	223
Figure 18-4: Service Bay Location	227
Figure 18-5: Service Bay Layout	228
Figure 22-1: Sensitivity Analysis for Silver Price	258
Figure 22-2: Sensitivity to Gold Price	258
Figure 22-3: Sensitivity to Zinc Price	259
Figure 22-4: Sensitivity to Lead Price	259
Figure 22-5: Technical Economic Model	260

xiv

Figure 24-1. Hydraulic Conductivity Summary	264
Figure 24-2: Groundwater Model Location	267
Figure 24-3. Simulated Mine Drains and Dewatering Wells	269
Figure 24-4. Predicted Mine Inflow and Dewatering Well Pumpage for All Simulated Scenarios	271
Figure 24-5: Predicted Mine Inflows and Dewatering Well Pumping Rates for All Dewatering Scenarios	273
Figure 24-6: Predicted Drawdown at Pozos 5 and 6	274
Figure 24-7: Predicted Drawdown at Pozos 11 and 12	275
Figure 24-8: Maximum Predicted Drawdown	276
Figure 24-9. Drawdown at End of Mining	277
Figure 24-10: Pumping wells in relation to the overall mine plan	279
Figure 24-11: Water pumped at the Los Gatos site, blue line represents mine infiltration and orange represents the dewatering wells	280
Figure 24-12: Soils Map of the Project Area	286
Figure 24-13: Site Photo Showing Vegetative Cover	289
Figure 24-14: Type II Temporal Distribution (excerpt from Urban Hydrology of Small Watersheds)	292
Figure 24-15: TSF Stage Storage	294
Figure 24-16: Channel Section Views	297
Figure 24-17: Well and Mine Working Dewatering	300
Figure 24-18: TSF Embankment Raise Schedule	302
Figure 24-19: Well Water, Use and Discharge	303
Figure 24-20: Tailing Storage Facility Plan View	305
Figure 24-21: Tailing Storage Facility Section View	306
Figure 24-22: Seismic Regions of Mexico	308
Figure 24-23: Hydrologic Zones of Mexico	308
Figure 24-24: TSF capacity curve	310
Figure 24-25: Underdrain for Stage 1, Phase 1	313
Figure 24-26: Tailing Deposition Plan — Stage 1, Phase 2 (Part 1)	314
Figure 24-27: Tailing Deposition Plan — Stage 1, Phase 2 (Part 2)	315

xv

ACRONYMS AND ABBREVIATIONS

Acronym	Definition
ABA	Acid Base Accounting
ABCC	Acid Buffering Characteristic Curve
ANFO	Ammonium Nitrate/Fuel Oil
AP	Area of influence of the Project
ARD	Acid Rock Drainage
ASI	Asesores en Impacto Ambiental y Seguridad
BBWI	Bond Ball Mill Work Index
bgs	Below Ground Surface
BMP	Best Management Practice
BRWI	Bond Rod Mill Work Index
CCTV	Close Circuit Television
CDN	CDN Resource Laboratories
Ce	Runoff Coefficient
CFE	Comisión Federal de Electricidad
CIM	Canadian Institute of Mining, Metallurgy and Petroleum
CO2	Carbon Dioxide
CONABIO	Comisión Nacional para el Conocimiento y Uso de la Biodiversidad
CONAGUA	Comisión Nacional del Agua
CPU	Computer Processing Unit
CRF	Cemented Rock Fill
CRP	Closure and Reclamation Plan
EHS	Environmental Health & Safety
EIS	Environmental Impact Study
EMS	Environmental Management System
ES	Environmental System
FICA	Federal Insurance Contributions Tax
FoS	Factor of Safety
FUI	Federal Unemployment Tax
G&A	General and Administration
GSI	Geological Strength Index
HDPE	High-density polyethylene
Hr	Hydraulic Radius
IES	Illumination Engineering Society
IFC	International Finance Corporation
INEGI	National Institute of Statistics and Geography (Mexico)

xvi

Acronym	Definition
LGEEPA	Ley General del Equilibrio Ecológico y la Protección al Ambiente
LOM	Life of Mine
Ltd.	Limited
Ma	Million Years Ago
masl	Meters Above Sea Level
MIA	Manifestación de Impacto Ambiental
ML	Metal Leachate
MPL	Maximum Permissible Limits
MPR	Minera Plata Real, S. de R. L. de C.V.
MSO	Maptek Stope Optimizer
N	Stability Number
NA	Not Applicable
NAF	Non Acid Forming
NAG	Net Acid Generation Potential
NOM	Norma Oficial Mexicana
NOMs	Normas Oficiales Mexicanas
NSR	Net Smelter Return
OD	Outside Diameter
OEM	Original Equipment Manufacturer
PA	Project Area
PAF	Potential Acid Forming
PAF-LC	Potential Acid Forming-Low Capacity
PAG	Potentially Acid Generating
PGA	Peak Ground Acceleration
PGM	Plant Growth Medium
PLC	Programmable Logic Controller
PMLU	Post Mining Land Use
PMP	Probable Maximum Precipitation
PN	Neutralization Potential
PNN	Potential of Net Neutralization
PPE	Personal Protective Equipment
PROFEPA	Procuraduría Federal Protección al Ambiente
QA	Quality Assurance
QC	Quality Control
REIA	Regulation on Environmental Impact Assessment
REPDA	Republic Register of Water Rights
RFQ	Request for Quote

xvii

Acronym	Definition
RMR	Rock Mass Rating
ROM	Run of Mine
SAG	Semi-Autogenous Grinding
SCADA	Supervisory Control and Data Acquisition System
SCS	Soil Conservation Service
SEMARNAT	Secretaría de Medio Ambiente y Recursos Naturales
SG	Specific Gravity
SGM	Servicio Geológico Mexicano
SMC	SAG Mill Comminution
SPI	SAG Power Index
SSMRC	Sunshine Silver Mining & Refining Corporation
SUI	State Unemployment Insurance
TEM	Technical-Economic Model
tc	Time of Concentration
tlag	Time of Lag
TRS	Temporary Rock Storage
TSF	Tailings Storage Facility
Tt	Tetra Tech, Inc.
UCS	Unconfined Compressive Strength
UPS	Uninterruptable Power Supply
Ve	Average Annual Runoff Volume
Vp	Average Annual Rainfall Volume
WAD	Weak Acid Dissociable

xviii

LIST OF UNITS

Abbreviation	Definition	Abbreviation	Definition
µm	Micrometer	lb	Pound
AgEq g/t	Grams per Tonne Silver Equivalent	m	Meters
BCM	Banked Cubic Meter	M	Million
C	Degrees Centigrade	m/d	Meters per Day
Cm	Centimeters	m/s	Meters per Second
DC	Direct Current	m3	Cubic Meters
g	Gram	m3/s	Cubic Meters per Second
g/cm3	Grams per Cubic Centimeter	mg/kg	Milligram per Kilogram
g/t	Grams per Tonne	mg/L	Milligram per Liter
Ha	Hectares	mm	Millimeter
hm3	Hectare-meter	Mpa	Mega Pascals
HP	Horse Power	Mt	Millions Tonnes
Hr	Hour	MVA	Mega Volt Amp
Kcfm	Thousand cubic feet per minute	MW	MegaWatt
K	Thousand	pH	Hydrogen Ion Concentration
Kg	Kilogram	Oz	Ounce
KIL	Kilo Pound	ppm	Parts per Million
Km	Kilometers	S.U.	Standard Units
km2	Square Kilometers	t/m3	Tonnes per Cubic Meter
kN/m3	Kilonewton per Cubic Meter	toz	Troy Ounce
Koz	Kilo Ounce	tpd	Tonnes per Day
kV	Kilovolt	usgpm	US Gallons per Minute
kVA	Kilovolt Amp	V	Volt
kW	Kilowatt	VAC	Volts Alternating Current
kWh/t	Kilowatt Hour per Tonne	W/mC	Thermal Conductivity
L	Liter
L/s	Liters per Second

xix

ABBREVIATIONS OF THE PERIODIC TABLE

actinium = Ac	aluminum = Al	americium = Am	antimony = Sb	argon = Ar
arsenic = As	astatine = At	barium = Ba	berkelium = Bk	beryllium = Be
bismuth = Bi	bohrium = Bh	boron = B	bromine = Br	cadmium = Cd
calcium = Ca	californium = Cf	carbon = C	cerium = Ce	cesium = Cs
chlorine = Cl	chromium = Cr	cobalt = Co	copper = Cu	curium = Cm
dubnium = Db	dysprosium = Dy	einsteinium = Es	erbium = Er	europium = Eu
fermium = Fm	fluorine = F	francium = Fr	gadolinium = Gd	gallium = Ga
germanium = Ge	gold = Au	hafnium = Hf	hahnium = Hn	helium = He
holmium = Ho	hydrogen = H	indium = In	iodine = I	iridium = Ir
iron = Fe	juliotium = Jl	krypton = Kr	lanthanum = La	lawrencium = Lr
lead = Pb	lithium = Li	lutetium = Lu	magnesium = Mg	manganese = Mn
meltnerium = Mt	mendelevium = Md	mercury = Hg	molybdenum = Mo	neodymium = Nd
neon = Ne	neptunium = Np	nickel = Ni	niobium = Nb	nitrogen = N
nobelium = No	osmium = Os	oxygen = O	palladium = Pd	phosphorus = P
platinum = Pt	plutonium = Pu	polonium = Po	potassium = K	prasodymium = Pr
promethium = Pm	protactinium = Pa	radium = Ra	radon = Rn	rhodium = Rh
rubidium = Rb	ruthenium = Ru	rutherfordium = Rf	rhenium = Re	samarium = Sm
scandium = Sc	selenium = Se	silicon = Si	silver = Ag	sodium = Na
strontium = Sr	sulphur = S	technetium = Tc	tantalum = Ta	tellurium = Te
terbium = Tb	thallium = Tl	thorium = Th	thulium = Tm	tin = Sn
titanium = Ti	tungsten = W	uranium = U	vanadium = V	xenon = Xe
ytterbium = Yb	yttrium = Y	zinc = Zn	zirconium = Zr

xx

UNITS OF MEASURE

All dollars are presented in U.S. dollars unless otherwise noted. Common units of measure and conversion factors used in this report include:

Weight:

1 oz (troy) = 31.1035 g

Analytical Values:

	percent	grams per metric ton
1%	1%	10,000
1 g/t	0.0001%	1.0

xxi

1.0  SUMMARY

1.1  Introduction

Minera Plata Real, S. de R.L. de C.V. (MPR) commissioned Tetra Tech, Inc. (Tetra Tech) of Golden, Colorado to produce an independent Mineral Resource estimate to update the project resource with drilling completed since the release of the project Feasibility Study (FS) and Technical Report (TR) for the Los Gatos Project in 2017.  This Resource estimate includes infill drilling from both surface and underground to define mineralization and to upgrade the mineral classification where possible.  Drilling has been performed in the Central Zone, but most of the program was focused on the North West zone and its possible extension to the north.  Reserves have been calculated from the new mine plan, which was based on the updated Resource (September 2019) information included in this report.

Tetra Tech has adopted certain provisions from previous TR’s by Tetra Tech and Behre Dolbear as appropriate.  Where content from previous TR’s has been included, the authors have not relied upon previous authors and are taking responsibility for the sections indicated in the certificate of qualified persons of this TR.

1.1.1  Property Description, Location and Infrastructure

The Cerro Los Gatos Deposit is located in Northern Mexico in the South-central part of the State of Chihuahua, within the Municipality of Satevó.  It is roughly centered on Latitude 27° 34’ 17” N, Longitude 106° 21’ 33” W.  The mining concessions cover 103,086.83 hectares (ha) within a block of 17 continuous with 16 grouped mineral concessions in addition to one concession (Paula Adorada) that has been subsequently purchased within the area Northwest of the town of San José del Sitio.  Other mineral deposits (Amapola and Esther) have been identified within the Company’s properties with limited exploration and drilling showing geologic potential for further exploration and development.

The project is approximately 120 km South of the state capital of Chihuahua City, and approximately 100 km North/Northwest of the historic mining city of Hidalgo del Parral.  The site can be reached via Federal Highway 24 from either city in about two hours.

The Cerro Los Gatos deposit is part of the Los Gatos project owned by MPR (a joint venture company of Sunshine Silver Mining & Refining Corporation (51.5%) and Dowa Metals & Mining Co., Ltd. (48.5%).  A camp has been constructed onsite for housing and feeding workers for the mining operations.  Water resources in the region are mostly related to the Conchos River Basin, which includes the San Pedro, San Francisco de Borja, and Satevó River Sub-Basins.  Locally, there are significant amounts of water, with shallow groundwater recorded from most exploration drilling conducted by MPR.

Power to the Cerro Los Gatos site is supplied via a 115-kV utility transmission line.  This originates from the ‘San Francisco de Borja’ substation in Satevó (Chihuahua), where a new 115 kV connection has recently been installed.

MPR is the concession owner of a series of titled concessions encompassing 103,086.83 ha.  Titled Mining concessions are summarized in Table 4-1.  MPR also holds the rights to the Paula Adorada concession (Table 4-1) through exploration agreements with purchase options that have been duly executed and recorded in the Mexican Public Registry of Mines.  A detailed account of MPR’s obligations for the purchase of Los Gatos and Paula Adorada concessions is included in Section 4.2 .

1

Regarding the mining concessions, MPR has provided copies of the titles of the concessions, as well as a copy of the Legal Opinion regarding MPR Concessions, by VHG Servicios Legales, S.C., related to titles documentation, tax payments, and assessment works, presented on November 5, 2019 to MPR.  This legal opinion stated that all claims are in full force and effect.  MPR’s Los Gatos group of concessions have a period of validity that reaches to March 3, 2058 which is the period corresponding to the oldest concession and title for the group of 16 concessions, independent of the Paula Adorada additional concession.  According to the Mexican Law the validity of the concessions is subsequently renewable in periods of 50 years.  MPR’s information indicates that all concessions have complied with their obligations as of the report effective date, and mining duties have been paid.

MPR has arranged permission to enter and perform exploration activities in several private ranch and community owned (ejido) properties in the project area.  MPR has purchased surface lands covering the Cerro Los Gatos, Esther, Amapola, and Rodeo zones covering 5,478.9 ha.  All current and planned development activities for the Cerro Los Gatos deposit are located within surface lands owned by or where access is controlled by MPR.

In 2008, MPR obtained environmental permits for drilling, road construction, and surface access rights to local ranches.  Subsequently, in January 2009 and September 2009, MPR submitted the corresponding notice of activities to SEMARNAT to cover the development of access roads and drill sites required to drill 50 holes and proposed a request to increase the number of drill holes to 250, following the Regulation NOM-120-SEMARNAT-1997.  A new report filed on December 5, 2011 expanded the permit to 600 drill holes.  A further study for Change of Land Use — Technical Justification Study (Cambio de Uso de Suelos — Estudio Técnico Justificativo) was completed in 2014 and permission received in early 2015 for underground development activities associated with the construction of an exploration decline to intersect the mineralization of the Cerro Los Gatos deposit.  Copies of these reports and filing documents can be found in MPR’s files.  Environmental baseline data collection began in May 2010, to prepare for the development of future environmental studies (EIS) required for the project.  The Environmental Impact Study (Manifestación de Impacto Ambiental) for the development of the Cerro Los Gatos project was filed with the Mexican Environmental Regulatory authorities (SEMARNAT) on December 12, 2016 and was approved in 2017.

1.2  History

The Los Gatos project has been the subject of limited historical prospecting and mineral exploration, including the development of shallow workings, limited production, and preliminary exploration activities by Consejo de Recursos Minerales, now Servicio Geológico Mexicano (SGM) at the Esther, Amapola, Gavilana (Paula), and San Luis zones with references to the occurrence of silver, lead, and zinc.

The project was initially identified by La Cuesta International Inc. reconnaissance activities in 2005, and later offered to Los Gatos Ltd. (parent of MPR).  An initial letter of agreement for exploration work on the project was negotiated and a final contract was ratified in April 2006 between MPR (initially a Mexican subsidiary of Los Gatos Limited) and La Cuesta International S.A. de C.V. (Mexican subsidiary of La Cuesta International Inc.).  Only minor field work was conducted during 2006-2007 on the Los Gatos project during the waiting period for the initial concessions to be titled, and formal exploration activities and drilling were conducted by MPR from 2008.

2

1.3  Geology and Mineralization

The Cerro Los Gatos deposit is located in the contact zone between the Sierra Madre Occidental volcanic province of western Mexico and the Mesozoic Chihuahua basin, largely sedimentary, to the East.  It is also located in the general contact zone between the Sierra Madre Occidental (SMO), Chihuahua, and Parral Tectonostratigraphic Terranes.

The area is largely characterized by a thick sequence of Tertiary volcanic rocks that are generally dissected by a strong North-Northwest bearing fault system that divides the area into the plateau and barranca sections.  This sequence is subdivided in two major units, the Lower Volcanic Group and the Upper Volcanic Group.  The area is one of the largest known epithermal, precious-metal metallogenic provinces and is host to well-known gold-silver producing mining districts, including:  Concheño, Ocampo, Batopilas, San Dimas-Tayoltita, Topia, Guanaceví, Bacís, San Francisco, Santa Bárbara, Velardeña, San José del Oro, Cosalá, Mulatos, La Ciénega, El Sauzal, Pinos Altos and Candameña Mining Districts currently in operation and/or exploration, known since the 1600’s, and other projects and old mining areas such as the Guadalupe Los Reyes, Cordero and Lluvia de Oro.

The oldest rocks of the area are Mesozoic (Cretaceous) age sedimentary rocks belonging to the Chihuahua Platform/Mesozoic Mexican Sea (open basin environment) with predominant silver-lead-zinc mineralization which characterizes this region, and is commonly related to skarn, limestone replacement and Mississippi Valley type deposits such as Naica, Santa Eulalia, Bismark, San Pedro Corralitos, Parral-La Prieta and Sierra Mojada.

The dominant rocks of the Cerro Los Gatos Project area are thick accumulations of intermediate tuffs and lavas, with lesser felsic rocks, with an age of deposition from 39-35.5 Ma; and Upper Cretaceous (Cenomanian) to Lower Paleocene age sandstones, shales, and limestones correlative with the Mezcalera formation, locally metamorphosed to phyllites, quartzites, and marbles near areas of igneous activity.  Rocks of this oldest sedimentary sequence occur within a small horst block located to the Southwest of the Cerro Los Gatos Deposit, with prominent high-angle fault boundaries on the North and South, parallel to the regional trend of faulting.  Intruding and deposited on the entire section are locally important rhyolite flows, flow domes, and dikes, usually strongly silicified, that have all the varied textures expected with the development of flow domes, including breccia zones, flow banding, and intrusive/extrusive transitions.  Each of the rocks in the section contains observable hydrothermal alteration, suggesting that mineralization in the area probably occurred late in the history of the development of the volcanic section.  It is important to stress, however, that economic grades of mineralization have only thus far been identified in the andesite and dacite sections (Behre Dolbear, 2011)

1.4  Drilling

As of July 2019, 484 drill holes relevant to the Cerro Los Gatos Deposit had been completed by MPR, for a total of 150,842 m.  The project database contains drilling on other prospects that are not applicable to this report, and it shows drilling associated with the Cerro Los Gatos deposit.  Table 1-1 tabulates the drilling by purpose.

3

Table 1-1:  Drill Hole Count by Purpose

Purpose	Count	Length (m)
Surface Exploration	399	137,220
Underground Exploration	32	3,567
Exploration and Metallurgical testing	6	1,733
Underground Bulk Sample Targeting	4	415
Metallurgical Testing	5	1,693
Geotechnical	18	4,134
Tailings Geotechnical	14	280
Hydrologic Study	6	1,800
Total	484	150,842

Drilling was initiated at the Cerro Los Gatos project in October 2008 and continued until 2012.  Drilling recommenced in 2015 following the joint venture agreement with DOWA and has continued through 2019.

An additional 255 drillholes have been completed within the Los Gatos project, but outside of Cerro Los Gatos deposit.  This comprises an additional 108,218 meters of drilling, which are not included in the Resource estimate that is the subject of this report.  The totals of the entire project are 739 holes for 259,060 meters.

1.5  Mineral Resource Estimates

Mineral resources have been estimated for the epithermal veins of the Cerro Los Gatos deposit by multi-pass Ordinary Kriging (OK) of capped and composited drill hole samples.  The Cerro Los Gatos deposit is currently the main deposit, and the subject of the 2017 FS and current mine development.  Estimated Measured, Indicated and Inferred Mineral Resources for the Cerro Los Gatos deposit, effective September 6, 2019, are shown in Table 1-2 at a 150 AgEq g/t cutoff grade.  Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Table 1-2:  Mineral Resource Estimate

Classification	Tonnes	AgEq g/t	Ag g/t	Pb %	Zn %	Au g/t	Cu %	AgEq toz M	Ag toz M	Pb lbs M	Zn lbs M	Au toz K	Cu lbs M
Measured	5,774,314	652	324	2.9	5.8	0.39	0.11	121	60	375	744	72	13
Indicated	4,586,507	489	202	2.5	5.2	0.28	0.11	72	30	251	528	42	12
Measured and Indicated	10,360,822	576	269	2.7	5.5	0.34	0.11	193	90	626	1,272	114	25
Inferred	3,717,063	361	107	2.8	4.0	0.28	0.14	43	13	231	330	34	12

NOTES:

1) 150 AgEq g/t cutoff grade has been calculated using $18/toz Ag, $0.92/lbs Pb, and $1.01/lbs Zn,
2) Columns may not total due to rounding,
3) Mineral Resources are stated as undiluted, and are inclusive of Mineral Reserves.
4) One troy ounce (toz) is equal to 31.1035 grams (g) and one Tonne is equal to 2,204.62 lbs.

4

Mineral Resources were estimated from 2,356 samples intersecting modeled veins, sourced from 426 diamond drill holes.  Capping was analyzed for each metal estimated using histograms and probability plots to determine where high-grade distribution tails deviated from lognormal.  Sampled intervals were composited to 2 m.  Composite intervals initiated and terminated at the top and bottom of the vein contacts.

Vein model solids were constructed in MicroMine™ modeling software; the resulting solids are shown in Figure 1-1, below.  Grade-shells were used to further isolate +150 AgEq g/t grade population for estimation.

Figure 1-1:  Vein Solids 3D View

Blocks and composites from each vein and post mineral fault block domain were independently transformed, realigned and made relative to the footwall and hanging-wall for estimation.  Realignment allowed for estimation to occur across post-mineral fault blocks approximating pre-fault orientation of the veins.  Estimations relative to footwall and hanging-wall position allowed for better data honoring across the dip of the vein.

Only composites within the same vein were permitted to estimate blocks of a given vein domain; because of the transformation and realignment estimation was permitted across post mineral fault block areas with the same vein code.

In addition to the Cerro Los Gatos deposit, exploration targets with favorable results include the Amapola and Esther deposits.  These deposits were included in Tetra Tech’s August 2012 estimates and additional information has not been produced for these deposits since this time.  These estimates should be considered project resources, and therefore have been brought forward in this report and are reported below.

5

Table 1-3 shows estimated, Indicated and Inferred Mineral Resources for the Amapola and Esther deposits, based on 100 grams/tonne cutoff grade, located along the Los Gatos mineral trend has only been partially investigated by drilling.  The effective date of this estimate is December 21, 2012.

Table 1-3:  Estimated Mineral Resources Indicated and Inferred for Amapola and Esther.

Cutoff, AgEq g/t	Tonnes	Ag Eq g/t	Ag, g/t	Ag Ounces	Au, g/t	Pb, %	Zn, %	Cu, %
Deposit	Amapola — Indicated Resources
100	250,000	154	135	1,100,000	0.10	0.10	0.30	0.02
Deposit	Amapola — Inferred Resources
100	3,440,000	164	140	15,500,000	0.10	0.20	0.30	0.03
Deposit	Esther — Indicated Resources
100	460,000	217	133	2,000,000	0.04	0.70	2.10	0.02
Deposit	Esther — Inferred Resources
100	2,290,000	243	98	7,200,000	0.12	1.60	3.00	0.05

Reference: Tetra Tech Technical Report Los Gatos Final-2012, 1221.

Note 1.- Figures may not total due to rounding of significant figures.

Note 2.- Inferred Resources are not defined or recognized by US SEC Industry Guide 7 but are acceptable in proposed new Guidelines.

Original exploration activities at the Esther vein system indicated the presence of a narrow quartz vein, less than 1 m, with minor veining and silicification and noticeable lack of calcite.  However, the presence of a small high-grade mineralized-shoot, probably 60 m deep, attracted interest in the area.  Drilling of this area also resulted in the discovery of the Esther mineral-concentration, which has a known length of 800 m for the main mineralized-shoot, as well as up to 1,200 meters of additional mineralized vein.  The height of the mineralized interval is indicated by drill holes ES-06 and ES-07 to be in the order of 100 m; most mineral intersections range in the order of 2 m to 8 m, with a probable average slightly over 3 m.  It has been interpreted the top of the favorable horizon at Esther is generally located about 120 m below the surface.

Geological, exploration, and drilling information has identified a Mineral Resource in three of the targets explored, namely Cerro Los Gatos, Amapola and Esther.

Other vein systems detected during the early exploration stages included Mezcalera, Cieneguita, San Luis, Paula, Boca de León, El Lince, El Rodeo, Los Torunos and San Agustín, all of which host results of interest.  The only area drilled with variable results of interest was Ocelote.

1.6  Mineable Reserve Estimate

Updated Mineral Reserves were estimated for this project report.  Information can be found on mining in the Mining Methods section of this report.

The Mineral Reserve for the Cerro Los Gatos deposit is the economically mineable portion of the Measured and Indicated Mineral Resource from September 2019 that can justify economic extraction.  The Mineral Reserve defined herein includes the application of mining factors determined in this report

6

for stope design criteria, dilution and recovery factors.  The Mineral Reserve is supported by a mine plan based on detailed stope layouts.

The Mineral Reserve, effective July 1, 2020, for Los Gatos is presented in Table 1-4.  Reserves are calculated at an NSR cutoff of $75 and exclude Resources that have been mined through the end of June 2020.

Table 1-4:  Mineral Reserve

Zone	Classification	Tonnes	Ag (g/t)	Au (g/t)	Pb (%)	Zn (%)
NWZ	Proven	2,587,684	359	0.43	3.09	5.88
	Probable	492,892	333	0.34	2.86	5.88
CZ	Proven	3,767,456	314	0.31	2.55	5.32
	Probable	1,772,921	299	0.44	2.32	5.82
SEZ	Proven	5,751	148	0.16	3.69	7.23
	Probable	569,380	148	0.16	3.69	7.23
SEZ2	Probable	421,547	118	0.17	3.11	4.16
Total	Proven	6,360,890	332	0.36	2.77	5.55
Total	Probable	3,256,740	254	0.34	2.74	5.86
Total	Proven + Probable	9,617,631	306	0.35	2.76	5.65

1.7  Metallurgical Testing

The Cerro Los Gatos deposit is a lead-zinc-silver deposit with relatively complex mineralogy.  Upon review of the metallurgical testing data, Los Gatos mineralization responded well to conventional sequential lead-silver-zinc flotation processing.

Zinc rougher flotation tailing and zinc 1st cleaner scavenger tailing is combined to become the final tailing. Tailing thickener underflow (100%) is pumped to a cyanide destruction facility.  Currently all tailings are disposed of in the tailings storage facility.  Once the backfill plant is in operation, and after detox, forty percent (40%) of final tailings will be pumped to backfill plant and the remaining (60%) will be pumped to the tailings storage facility.  Operational results have indicated that additional removal of fluorine will be required.  Additional floatation cells have been added to the lead and zinc circuits, with a goal of reducing the fluorine in the final concentrates.

JK-drop weight and SMC test results show the samples are softer when compared to JKTech database while the SPI test results categorize the samples from moderately soft to moderately hard per SGS database.  The Ai and BWI tests results describe the samples as from abrasive to very abrasive and from moderately hard to hard.

Very good lead and silver flotation results have been achieved from SGS Vancouver Metallurgical Lab.  The final lead cleaner concentrates of the LCTs averaged 60.9% Pb at 5,404 g/t Ag at average recoveries of 89% lead and 68.7% silver.  The zinc cleaner concentrates averaged 54.2% Zn at an average recovery of 66% due to the high willemite content.

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1.8  Mining Methods

The underground mine design has been created to support a steady-state production rate of 2,500 tpd of ore.  The sequence of mining has been planned to begin with the Central Zone (CZ), which has already been accessed via the existing decline developed down to the 1400 Level.  The mine plan includes using cut and fill methods for the entirety of the CZ, in a top down sequence, starting in the central portion.

The NorthWest Zone (NWZ) has been planned to be mined concurrently with the CZ via longhole stoping methods with sublevels developed at 20 m vertical intervals.  Portions of the NWZ thicker than 9 m (footwall to hanging wall) will be mined using transverse longhole mining.  Areas less than 9 m in width will be mined using longitudinal longhole mining.

Production has started at the Los Gatos Project.  A combined overhand and underhand approach have been chosen in order to optimize the mining sequence based on the existing excavation extents in order to meet the production and grade requirements of the mine.  A mix of overhand and underhand mining has been implemented, starting on the 1390 Levels in both the NWZ and the CZ.

Limited production began in November 2018 and has been able to continue at a constant ramp-up through to the present time.  Due to the underhand approach, sill pillars are designed at necessary intervals in both cut-and-fill (CAF) and longhole stoping (LHS) blocks.  The underhand longhole stoping areas require re-mining or undercutting of the bottom sill drifts compared to the overhand approach.

Modern trackless mobile equipment is being utilized for most mining activities.  LHDs and dedicated underground trucks are used for ore/waste loading and transport from the underground workings through an internal ramp system and portal that connects all levels to surface.

Based on the deposit geometry and anticipated geomechanical conditions, underground mining of the Los Gatos Resource incorporates both longhole and drift-and-fill mining methods.  The existing exploration decline from surface has been extended to provide primary access and delivery of services.  The ramp is also used for haulage of ore and waste from the underground operations.

1.9  Recovery Methods

The expected grades and recoveries for lead, zinc, and silver to individual flotation concentrates were further investigated by a pilot plant program at SGS Lakefield using a sample composed of a bulk sample accessed by an underground decline into the orebody.  The results of that program are being finalized but preliminary indications generally support the study grade and recovery estimates provided herein.

The Project processing facility is designed to treat 2,500 tpd of lead, zinc and silver material at an operational availability of 92 percent.  The processing flow sheet for the Project is a standard flow sheet that is commonly used in the mining industry, including conventional flotation recovery methods typical for lead-zinc material.  Figure 1-2 below is the overall process plant flowsheet.  M3 engineering, has revised and updated the process design based on the results of 2016 SGS Vancouver metallurgical testing programs, and MPR’s input.

Run-of-mine (ROM) material is crushed in a primary jaw crusher located adjacent to the underground mine portal.  From there it is conveyed to the processing facilities where and is ground to 80 percent finer than 45 microns in a semi-autogenous grinding (SAG) and ball milling circuit.

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The ore is further processed in a flotation circuit consisting of lead flotation followed by zinc flotation.  The majority of the silver is recovered in the lead flotation circuit and some silver is also collected in the zinc flotation circuit.

Lead sulfide is recovered in a rougher flotation bank, producing a concentrate that is upgraded to smelter specifications in three stages of cleaning.  Tails from the lead flotation section are then be conditioned for zinc sulfide flotation.  The process scheme for zinc flotation also includes a rougher bank and five stages of cleaning to produce smelter-grade zinc concentrates.  For both lead and zinc sections, the rougher flotation concentrates are reground to 80 percent finer than 25 microns prior to cleaner flotation to liberate the sulfides for further upgrading.  The plant is currently operating without re-grinding, final primary grind is currently achieving the final desired grind size, during start-up.  Re-grind will be implemented as necessary when the plant reaches full design capacity and a greater load is fed to the primary grinding circuit.

An additional deep-froth flotation cell has been added to the zinc circuit, and one additional deep-froth flotation cell will be added to the zinc and lead circuits during 2020, to remove more fluorine, a deleterious mineral for sales.  Both final lead and zinc concentrates are thickened, filtered, and stored in concentrate storage facilities prior to loading in trucks for shipment.

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1.10  Infrastructure

Los Gatos Project is located in the Municipality of Satevó, Chihuahua, Mexico, approximately 160 kilometers, southwest of the state capital of Chihuahua City and about 8 kilometers west of San José del Sitio, Chihuahua.

The access road from Chihuahua, Chihuahua, México is newly paved.  A portion of the road from San José del Sitio has been rerouted to the mine site to minimize interference with the river that runs by the mine property.

The camp consists of a structural steel pre-engineered building capable of housing 350 people.  The camp includes a kitchen and cafeteria, laundry, infirmary, and other buildings required to maintain this facility.  Emergency power is designed for the camp to provide 100% backup in the event of a power outage.

The camp is currently serviced by a satellite dish-based internet and TV connection.  Mobile phone service at the mine site has been improved.  This upgraded system provides communication capabilities for surface and underground personnel.

Power to the site is supplied via a 115 kV utility transmission line.  This line originates from the San Francisco de Borja substation and is approximately 66 km long.  Based on the process and mine equipment, a total electrical load of approximately 24 MW is required for the project.  The diesel generators used during power line construction remain on site for back-up power as required.

Raw water to meet potable and non-potable water demand is supplied by groundwater pumped from dewatering wells.  The well water is cooled from 50°C to 40°C prior to use.  Groundwater from an existing well is suitable quality for potable and non-potable uses.

Sewage water treatment systems were designed to handle waste as required on the project.

Storage and management of landfill disposal is in a building and separated into two zones, one for non-hazardous waste and a second zone for hazardous waste.  The hazardous waste is collected and disposed by a certified and authorized company per Mexican regulations.

1.11  Marketing Studies and Contracts

MPR has secured contracts and agreements for the following for the development of the project:

·Minera Plata Real, S. de R.L. de C.V. (MPR), along with Operaciones San Jose de Plata S. de R.L. de C.V. and Servicios San Jose de Plata S. de R.L. de C.V., form the Los Gatos Joint Venture, owned by Sunshine Silver Mining & Refining Corporation (51.5%) and Dowa Metals & Mining Co., Ltd. (48.5%).

·Royalty agreement with La Cuesta International S.A. de C.V. under the terms of the document, Contrato de Exploración, Explotación y Promesa — La Cuesta International, S.A. de C.V. and Minera Plata Real, S.A. de C.V., dated April 2006.  Under the terms, MPR pays a royalty payment of US $40,000 per year during the preproduction period.  When production was initiated a net smelter return (NSR) royalty began at 2% on production from the Los Gatos concession.  This is reduced to 0.5% upon all payments reaching $10 million.  The maximum royalty payment for this agreement is set at US $15 million.

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·MPR holds the rights to 17 mining concessions totaling 103,086.8 ha.  These have been duly executed and recorded in the Mexican Public Registry of Mines (Vazquez, Sierra, and García, S.C) for the mining concessions title in register number 231498 dated March 4, 2008, and updated Title Opinion by VHG Servicios Legales, S.C. dated on November 5, 2019, Minera Plata Real Concessions.

·Mine and land access agreement with the Ejido La Esperanza, Contrato de Usufructo dated April 13, 2012 with an annual cost of US$ 11,200.

·MPR has an easement contract for the access road on land owned by Ejido San José.  This easement contract is for 30 years.

Indicative smelting and refining terms have been prepared for the zinc and lead concentrates and were used in the economic analysis.  Details can be found in Section 19.

MPR has obtained metal price forecasts for the four payable metals from Los Gatos from nine financial institutions:  Table 1-5 contains the average long-term forecast that were used in the economic analysis.

Table 1-5:  Payable Metal Price Forecasts for Los Gatos

Metal	Units	Consensus Long-Term Price
Gold	$/Ozs	1,472
Silver	$/Ozs	18.99
Lead	$/lb	0.87
Zinc	$/lb	1.09

MPR has no lease agreements for the Los Gatos project.

1.12  Environmental Studies, Permitting, and Social or Community Impact

According to the environmental system inventory, no relevant or critical areas were found, nor were protected natural areas or conservation areas established by National Commission for the Knowledge and Use of Biodiversity (Comisión Nacional para el Conocimiento y Uso de la Biodiversidad — CONABIO) identified, except for the Priority Hydrological Region 39, identified as the Upper Conchos River Basin.  The basin could be directly affected by the Project because the runoff may be affected by the construction of mining infrastructure, by the forest removal, and the modification of the original structure of the soil.

On December 12th, 2016 the Environmental Impact Statement (EIS or MIA) was submitted to the Secretary of Environment and Natural Resources (SEMARNAT).  The SEMARNAT regulates the environmental aspects of mining projects and issues the permits once the EIS is approved, according to Art. 28, Frac. III, VII and X of the General Law of the Ecologic Equilibrium and Environmental Protection, and by the Art- 5 Section L), Frac. I and III, Section O) Frac. I and Section R), Frac. I of the Regulations for Environmental Impact Statement.  The MIA was approved in 2017.

The project is in the municipality of Satevó, in the state of Chihuahua, with a population of 452 people, according to the official record of year 2010.  The current land uses are livestock, agriculture and recently mining due to the execution of the Los Gatos mining exploration and production that has been carried out since 2009.

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The following Federal, State and Municipal Permits issued for the Cerro Los Gatos Mining Operation have been applied for and issued by Federal, State and Municipal authorities as required for exploration, preparation, development and operations during the 2018 and 2019, have been obtained by Minera Plata Real:

·SEMARNAT — Environmental Impact Statement (Manifestación de Impacto Ambiental — Modalidad Regional) (MIA-R) issued on 2017, No. SGPA/DGIRA/DG/05121-2017.

·SEMARNAT 1st.  Modification to increase Impacted area from 211.0841 ha to 268.8450 ha and from 95 to 99 mine workings, No. SGPA/DGIRA/DG/01914, March 15, 2018.

·SEMARNAT 2nd.  Modification to increase Impacted area from 268.8450 ha to 325.0860 ha and from 99 to 133 mine workings, No. SGPA/DGIRA/DG/09272, November 28, 2018.

·SEMARNAT — Exemption to present MIA for Extension of Road from San José to Los Gatos mine. No. SG.IR.08-2018/097, May 4, 2018.

·SEMARNAT — Modification of trajectory for “Power Line 115 KV Los Gatos”.  No. SG.IR.08-2018/093, May 4, 2018.

·SEMARNAT — Authorization of Preventive Report for Direct Mining Exploration, diamond Drilling in Los Gatos NW-CE-SE, Cascabel Fault and El Valle Vein”.  No. SG.IR.08-2019/070, May 7, 2019.

·CONAGUA — Residual waters discharge from Los Gatos into Santo Toribio Creek, 8.0 l/s. No. 06CHI141265/24FADL16, August 31, 2018.

·CONAGUA — Authorization for residual waters discharge from Los Gatos into Santo Toribio Creek, 8.0 l/s. No. 06CHI141265/24FADL16, August 31, 2018.

·CONAGUA — Authorization for increment of residual waters discharge and change of point from Los Gatos into Santo Toribio Creek, 8.0 l/s to 120 l./s. No. 06CHI141265/24FADL16, July 16, 2019.

·CONAGUA — GASIR — Authorization for construction and operation of Tailings Storage Facilities No. 1 with capacity for 7.6 Mm3 to be built is four stages and a period of 9 years for construction. No. 4494, January 18, 2019.

·SEMARNAT — Approval of Environmental Unique License (Licencia Ambiental Unica, LAU), for production of 2,500 tpd for MPR, No. LAU-CHIH-001-2019, May 27, 2019.

·DGGIMAR — SEMARNAT — Approval of Registry for Management of Dangerous Residues for production of 144.642 tonnes/year, No. 08-PMG-I-3405-2019, March 26, 2019.

·Secretaría de Desarrollo Urbano y Ecología (SEDUE) Chihuahua — Approval of Registry and Generating Corporation with Plan of Management for Residues with Special Handling, approval for production of 468.747 tonnes/year, April 9 and 11, 2019.

·During the months of June and July 2018, permits and approvals by the Municipality of Satevó, Chihuahua of included the following permits:

·Permit for the Use of Land;

·Authorization and approval for initiation of mining construction workings and infrastructure; and

·Official alignment and number.

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1.13  Capital and Operating Costs

These costs are presented as a summary, refer to Section 21 for detailed estimates.

All costs and economic results are presented in 2020 U.S. dollars.  Quantities and values are presented using standard metric units unless otherwise specified.  No escalation has been applied to capital or operating costs.  No gearing is assumed in the analysis.

Technical economic tables and figures presented in this report require subsequent calculations to derive subtotals, totals, and weighted averages.  Such calculations inherently involve a degree of rounding.  Where these occur, they are not considered to be material.

1.13.1  LOM Capital Costs

These costs are presented as a summary.  LOM capital cost requirements are estimated at $267 million, including sustaining capital, as summarized in Table 21-4.  The Initial capital of $316 million was required to commence operations and construction was completed on time and on budget.

Table 1-6:  LOM Capital Costs

Description	Units	Sustaining Capital
Direct Costs
Mine & Surface Infrastructure	$000s	266,398
Direct Costs	$000s	266,398
Indirect Costs
Mine & Surface Infrastructure	$000s	932
Indirect Costs	$000s	932
Total Sustaining Capital	$000s	267,330

1.13.2  LOM Operating Costs

LOM operating costs are summarized in Table 1-7.  These costs are based on the current operations.

Table 1-7:  LOM Operating Costs

Description	Unit Cost ($/meter)	Unit Cost ($/t-milled)
Mine, Surface and G&A	26.47	83.58
LOM Operating	—	83.58

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1.14  Economic Analysis

The economic analysis is presented on an unlevered, post-tax, present value (PV) basis.   Valuation estimates presented in this technical report should be adjusted for existing LGJV current liabilities, receivables and long-term indebtedness.  Economic results are summarized in Table 22-2.  The analysis suggests the following conclusions, assuming no gearing:

·Mine Life: 11 years;

·Pre-tax present value (PV5.0%): $764 million;

·Post-tax present value (PV5.0%): $653 million;

·Taxes Paid: $148 million; and

·Sustaining project capital of $267 million

Table 1-8:  TEM Results

Description	Unit Cost ($/t-milled)			LOM Value ($000s)
Net Smelting Return	$	214.04		2,058,579
La Cuesta Royalty as of June 2020	$	(1.50	)	(14,415	)
Net Revenue	$	212.54		2,044,164
Operating Costs
Mine, Surface, and G&A	$	(83.58	)	(803,835	)
Operating Costs	$	(83.58	)	(803,835	)
Operating Margin	$	128.96		1,240,329
Capital Costs
Sustaining Capital Costs	—			(267,330	)
Capital Costs	—			(267,330	)
Pre-Tax Cash Flow
Cash Flow	—			978,867
PV5.0%	—			764,690
Post-Tax Cash Flow
Cash Flow	—			830,653
PV5.0%	—			653,166

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1.15  Conclusions and Recommendations

A new tradeoff study should be undertaken to evaluate the possibility of expansion to 3,000 tpd.  The crushing and grinding circuits were designed with the capacity of 3,000 tpd, and additional Resources have been identified.  This new study should evaluate the additional cost for mine and flotation expansion.

1.15.1  Geology and Resources

1.15.1.1  Conclusions

Project geologic and drill hole data for the Los Gatos project has been collected and analyzed by MPR using industry standard methods and practices and is sufficient to characterize grade and thicknesses of the deposit and to support the estimation of Measured, Indicated and Inferred Mineral Resources.  Although the deposit has been densely drilled, Resource expansion potential and project upside exist in the immediate deposit area, as well as other identified prospects throughout the land package.

1.15.1.2  Recommendations

It is recommended that:

·Additional drilling is conducted to further convert Inferred Resources and expand Resources at the Cerro Los Gatos deposit.  The current Resources are significant but additional Resource potential remains in the immediate area.  Drilling is specifically recommended:

·Down-dip and along strike to the southeast in the SE Zone;

·In the detached blocks in the hanging wall of the NW block that are currently classified as Inferred;

·Additional definition of the plunging shoot in the SE3 block and potential down-dip extent of mineralization of the SE3 block; and

·Down-dip in the Central block following up on high-grade encountered in GA-55, GA-66, and GA-243 to determine if the system continues in some other form.

·Additional infill drilling is conducted at Amapola and Esther to delineate mineralized shoots and assess full Resource potential;

·Following potentially positive results from infill drilling at Amapola and Esther, the Resources should be updated, and a scoping study conducted to determine if the two deposit areas could contribute to the economics of the Cerro Los Gatos project;

·Surface mapping and sampling is conducted in greater detail to refine and prioritize prospects in the project area; and

·Geophysical surveys are conducted following prospect prioritization, but before exploration drilling.

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1.15.2  Mineral Reserves

1.15.2.1  Conclusions

Mineral Reserves were calculated for the updated mine plan, based on the September 2019 Resource update.

1.15.2.2  Recommendations

It is recommended to continually review and update the Reserves throughout the mine life.  Increased Resource will most likely increase the life of the mine and a new mine plan should consider potential Resources discovered in the future.  It should also consider updated geotechnical information to define the pillar size requirements.

1.15.3  Mine Planning

1.15.3.1  Conclusions

Based on the deposit geometry and anticipated geomechanical conditions, as well as currently observed field conditions, the economic extraction of the Los Gatos Resource will continue to incorporate both longhole mining and drift-and-fill mining methods.  Modern trackless mobile equipment is being utilized for all development and mining activities.  The exploration decline from surface has been extended to provide primary access and delivery of services.  The ramp is also used for haulage of ore and waste from the underground operations.

Ongoing waste development to sustain the 2,500 tpd production rate averages approximately 211 m/month during the production period.

The life of mine is scheduled at approximately 2,500 tpd for a total of 11 years, with steady-state ore production reached in the second quarter of 2021.  Along with the Inferred Resources, there are indications of additional Resources along strike that, with additional drilling, may increase the mining Resource.

1.15.3.2  Recommendations

The following recommendations were made for Mine Planning:

·Refine the detailed mine design and schedule for the first five years of the project.  Look for opportunities to improve the average grade by selectively targeting higher-grade areas.

·Review the exploration development necessary to increase the Mineral inventory and incorporate it into the mine plan.

·Complete additional hydrological studies to predict ground water inflows more accurately.  Hot water inflows not only impact the dewatering system detail design, but they also increase ventilation and cooling requirements.

·An operational solution to manage/stop hot water inflows through ungrouted diamond drill holes is required.

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1.15.4  Mineral Processing and Metallurgy

The Cerro Los Gatos deposit is a silver, lead, and zinc Resource.  Lead and zinc occur primarily as galena and sphalerite, respectively.  Significant amount of willemite is identified in all tested samples.  Lead oxide minerals are also identified in some of the samples, especially from South East zone samples.  The existence of lead and zinc oxide minerals impacted their flotation performance.

JK-drop weight and SMC test results show the samples are softer when compared to JKTech database while the SPI test results categorize the samples from moderately soft to moderately hard per SGS database.  The Ai and BWI tests results describe the samples as from abrasive to very abrasive and from moderately hard to hard.

Very good lead and silver flotation results have been achieved.  The final lead cleaner concentrates of the LCTs averaged 60.9% Pb at 5,404 g/t Ag at average recoveries of 89% lead and 68.7% silver.  The zinc cleaner concentrates averaged 54.2% Zn at an average recovery of 66% due to the high willemite content.

The expected grades and recoveries for lead, zinc, and silver to individual flotation concentrates utilized in this report were further investigated by a pilot plant program using material composed of a bulk sample accessed by an underground decline into the orebody.  The results of that program indicate the grade and recovery estimates provided herein.  The production plant is currently in operation.

1.15.4.1  Recommendations

During operations, it has become evident that additional removal of fluorine is necessary.  The Cerro Los Gatos Mine has reduced the fluorite content of the concentrate by providing additional cleaning stages, four and six stages in the lead and zinc floatation sections (from three and five).  This is done to minimize the recovery of the fluorite by entrainment/achieve the targeted fluorite rejection.

1.15.5  Economics

Economics have been updated for this report considering the updated Resources and Reserves.  The results of the new estimate have improved due to higher grades, additional tonnes, and actual costs.

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2.0  INTRODUCTION

Minera Plata Real, S. de R.L. de C.V. (MPR) is a Joint Venture Company owned by Sunshine Silver Mining & Refining Corporation (51.5%) and Dowa Metals & Mining Co., Ltd. (48.5%).  Tetra Tech, Inc. (Tetra Tech) has been contracted to prepare this report update for the Los Gatos Project in Chihuahua, Mexico.  This updated report contains new information on the September 2019 Resource update and Reserves based on a mine plan produced from the updated Resources.  Economics have been updated based on the new Resource and Reserve calculations.

2.1  Terms of Reference

This report has been prepared as a Technical Report for MPR by Tetra Tech.  The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in consultants’ 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.

2.2  Scope of Work

The scope of work conducted by Tetra Tech per the request of MPR is the development of a report update to include a Resource update based on drilling that has been completed since the FS was finished in 2017.  The new Resource has been utilized in the creation of an updated mine plan, which has been used to calculate updated Reserves.  Information was brought forward from the 2017 FS as appropriate, with relevant changes detailed where they have occurred in the other sections.

2.3  Units of Measure

The Metric system has been used throughout this report.  All currency is in US dollars ($) unless otherwise stated.

2.4  Detailed Personal Inspections

1)Guillermo Dante Ramírez-Rodríguez, PhD, MMSAQP, inspected the site from September 30—October 1, 2015, January 17, 2017 and August 20-21, 2019.

2)Leonel López, CPG, SME-RM visited the site on November 29-30, 2018 and August 20-21, 2019.

3)Kenneth Smith, SME QP visited the site January 17, 2017 and August 20-21, 2019.

4)Kira Johnson, MMSAQP visited the site July 17-18, 2012 and August 20-21, 2019.

5)Keith Thompson, P.G. visited the site February 23-28, 2015; July 10-14, 2015; December 12-15, 2015; January 29—February 2, 2016; and September 17-21, 2018.

6)Luis Quirindongo, SME QP visited the site June 25-26, 2015.

7)Max Johnson, P.E. has not visited the site.

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3.0  RELIANCE ON OTHER EXPERTS

Regarding the mining concessions, the author has been provided with a title opinion by the law firm VHG Servicios Legales, S.C. that related to titles documentation, tax payments, and assessment works, presented November 5, 2019 to MPR, the opinion stated that all claims are in full force and effect.  According to Title Opinion, all the Los Gatos mining concessions are grouped, except for the Paula Adorada concession.

The author has relied on the title opinion and statements by MPR that the claims and agreements are in good standing.

The mine plan has been generated by MPR staff and verified by Tetra Tech for the calculation of Reserves.

Tetra Tech has also relied on MPR’s input regarding information for economics, processing, and operations.

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4.0  PROPERTY DESCRIPTION AND LOCATION

The Los Gatos project is located in the South-central part of Mexican State of Chihuahua, within the Municipality of Satevó (Figure 4-1).  Chihuahua borders the neighboring states of Coahuila, Durango, Sinaloa and Sonora, and shares a common border with the United States of America.

Chihuahua has a long mining history with substantial production of silver, gold, lead and zinc from deposit districts such as Santa Eulalia, Naica, Santa Bárbara, San Francisco del Oro, Bismark, and new deposits such as El Sauzal, Palmarejo, and Dolores.

Figure 4-1:  Los Gatos General Location Map

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4.1  Location

The Cerro Los Gatos Deposit is approximately centered on Latitude 27° 34’ 17” N, Longitude 106° 21’ 33” W, near the town of San José de Sitio.  It is approximately 120 km South of the state capital of Chihuahua City and approximately 100 km North/Northwest of the historic mining city of Hidalgo del Parral.  The project is accessible by automobile from Mexican Federal Highway 24 to kilometer 81 and then turning West on a newly paved road for 40 km West to the community of San José de Sitio, which is situated near the Southeast end of the concession block.  Travel time by automobile is approximately two hours, either from Chihuahua City from the North or Hidalgo del Parral from the South.  The city of Valle de Zaragoza, located on federal highway 24, 35 km to the South of the turnoff to San José del Sitio is the nearest significant commercial center.

4.1.1  Mining Concession

MPR is the owner of mineral rights held by 16 titled concessions grouped under Los Gatos Group concession in addition to the concession Paula Adorada, covering 103,086.83 ha. Titled mining concessions are summarized in Table 4-1.

Table 4-1:  Los Gatos Project Titled Mining Concessions

	Lot	Holder	Surface (Hectares)	Title	Type of Concession	Term	Location
1	Los Gatos	MPR	19,711.6889	231498	Mining	March 3, 2058	Satevó, Chihuahua
2	Los Gatos 2	MPR	10,719.5765	228950	Mining	February 21, 2057	Satevó, Chihuahua
3	Los Gatos 3	MPR	27.2846	231076	Mining	January 15, 2058	Satevó, Chihuahua
4	Los Gatos 4	MPR	52,596.9673	238511	Mining	September 22, 2061	Satevó, Chihuahua
5	Mezcalera	MPR	4,991.6263	228249	Mining	October 16, 2056	Satevó, Chihuahua
6	Mezcalera 2 Fracción I	MPR	39.2621	228929	Mining	February 20, 2057	Satevó, Chihuahua
7	Mezcalera 2 Fracción II	MPR	26.1402	228930	Mining	February 20, 2057	Satevó, Chihuahua
8	Mezcalera 2 Fracción III	MPR	29.0859	228931	Mining	February 20, 2057	Satevó, Chihuahua
9	La Gavilana	MPR	10.0000	237137	Mining	November 18, 2060	Satevó, Chihuahua
10	La Gavilana Fracción I	MPR	44.0000	237461	MINING	December 20, 2060	Satevó, Chihuahua
11	Paula Adorada	MPR	40.0000	223392	Mining	December 8, 2054	Satevó, Chihuahua
12	San Luis	MPR	16.0000	236908	Mining	October 4, 2060	Satevó, Chihuahua

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	Lot	Holder	Surface (Hectares)	Title	Type of Concession	Term	Location
13	Los Estados Fracc. 1	MPR	9.0000	237694	Mining	April 25, 2061	Satevó, Chihuahua
14	Los Estados Fracc. 2	MPR	44.0000	237695	Mining	April 25, 2061	Satevó, Chihuahua
15	San Luis 3	MPR	0.0111	240452	Mining	May 22, 2062	Satevó, Chihuahua
16	San Luis 2	MPR	42.3904	238694	Mining	October 17, 2061	Satevó, Chihuahua
17	Los Veranos	MPR	14,739.8002	238573	Mining	September 22, 2061	Satevó, Chihuahua

The Cerro Los Gatos Joint Venture holds these concessions through its 100%-owned Mexican subsidiary company, Minera Plata Real S. de R.L. de C.V. (MPR).  The Los Gatos Joint Venture is 51.5% owned by Sunshine Silver Mining & Refining Corporation and 48.5% owned by Dowa Metals & Mining Co., Ltd.  MPR holds the rights to the concession of Paula Adorada through an exploration agreement with purchase option, which has been duly executed and recorded in the Mexican Public Registry of Mines (VHG Servicios Legales, S.C.).  The Los Gatos concession (Title 231498) is held subject to a royalty provision to La Cuesta International, Inc.

Ownership of the Los Gatos Joint venture partnership has been established and is registered under the following agreements:

·According to Legal Opinion issued by VHG Servicios Legales, S.C. dated on November 5, 2019, ownership of the concessions mineral rights is under “Non-possessory Pledge Agreement entered into by and between MPR and Dowa Metals & Mining Co. LTD.  Recorded on May 21, 2018, under book 129, number 137 and volume 41 of the Book of Mining Acts and Contracts”.

·Another second “Non-possessory Pledge Agreement entered into by and between MPR and Dowa Metals & Mining Co. LTD., ratified on July 31, 2019 in process of being recorded”.

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Figure 4-2 shows the Los Gatos Concessions Map.

Figure 4-2:  Concessions Map for the Los Gatos Concessions

The details of the two contracts follow, as provided by MPR.

4.1.2  Los Gatos and Paula Adorada Concessions

On the Los Gatos concession, MPR is required to make semi-annual advance royalty payments, and is required to make a production royalty payment of 2% net smelter return on production from the Los Gatos concession (reduces to 0.5% upon all payments reaching $10 million) and 0.5% net smelter return from lands within a one kilometer boundary of the Los Gatos concession to La Cuesta International S.A de C.V..  Under the terms, MPR paid a royalty payment of US $40,000 per year during the preproduction period.  When production was initiated a 2% net smelter return (NSR) royalty was to begin on production from the Los Gatos concession.  This is reduced to 0.5% upon all payments reaching $10 million) and 0.5% net smelter return from lands within a one-kilometer boundary of the Los Gatos concession.  Upon commencing production, payments under the royalty agreement were deferred until March 31, 2021 with an annual interest rate of 4.5% applied to the outstanding balance.  During the deferral period, MPR pays a royalty of $100,000 per year until January 2021.  The agreement has no expiration date; however, the

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Company may terminate the agreement upon 30-day official termination notification.  The registered ownership of the Los Gatos concession (title 231498) was transferred from La Cuesta to MPR in 2015.

The Paula Adorada concession was acquired from the Chihuahua based company Grupo Factor through an option to purchase agreement in effect from 2008-2013.  Once all obligations and payments were completed during the term of the agreement, the registered ownership of the Paula Adorada concession (title 223392) was transferred from Grupo Factor to MPR in 2014 with no remaining obligations or royalties.

4.1.3  Internal Concessions not held by MPR

There are nine small concessions within the Los Gatos project area that are not held by MPR.  MPR plans to apply for the canceled concessions when they are formally liberated.  Tetra Tech notes if there is more than one application at the moment the claim is declared free by the Chihuahua’s DGM Agency for the new concession, the winner by lottery held by the Mining Agency will be assigned the new rights.

Table 4-2:  Internal Concessions

Concession Name	Title	Surface Area	Owner	Status
Ampl. Este de San Luis	187436	42.4019	Mario Humberto Ayub Touché	Canceled
Ampliación Oeste de San Luis	187432	53.3725	Mario Humberto Ayub Touché	Current
La Calesa	187433	4.1992	Fidencio Chávez Soto	Canceled
La Concepcion	188161	10.0000	Minas de Satevó, S.A. de C.V.	Canceled
San Brígido	191780	80.0000	José Luis Chávez Cobos	Current
Ampl. de San Brígido	195307	50.0000	Fidencio Chávez Soto	Current
La Cieneguita	204861	21.0000	Minas de Satevó, S.A. de C.V.	Canceled
Margarita	171530	70.8550	Cía. Minera La Perla, S.A.	Current
El Tren	172158	54.7700	Cía. Minera La Perla, S.A.	Current

Under Mexican Mining Legislation, titled concessions must have submitted the required Surveying and Assessment Works to define their precise location and rights against any pre-existing mining claim.  Once titled, concession owners have the obligation to submit annual Assessment of Work Reports for each concession or group of concessions based on minimum investment amounts.  Tetra Tech notes all the Los Gatos concessions are grouped for legal compliance with requirements by Mexican Mining Law, except for the Paula Adorada claim, which was acquired later.

A second obligation that titled mining concessions must meet is the bi-annual payment of mining duties.  Tetra Tech notes that in Title Opinion dated November 5, 2019 by VHG Servicios Legales, S.C. for the Los Gatos Project, MPR has complied with the payment of mining duties up to the second semester of 2019.

According to the Title Opinion issued by the law firm of VHG Servicios Legales, S.C. dated November 5, 2019 for the Los Gatos Project, MPR has complied with their obligations and mining duties have been paid, including the second semester of 2019; therefore, MPR concessions are in good standing.

Titled mining concessions, following the amendments made to the Mining Law in 2005, have an effective period of 50 years counted from their registration in the Public Registry of Mines and can be renewed for

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equal periods provided there are no grounds for cancellation; therefore, MPR’s concessions have a period of validity through March 3, 2058, which corresponds to the oldest concession within the Group, Los Gatos concession, and is assigned to the Los Gatos Group, and independently to 2054 for the Paula Adorada concession, which may be also incorporated to the group upon application by MPR. (Table 4-1).

4.2  Surface Rights

MPR has arranged permissions to enter and perform exploration activities on several land properties in the project area.  Figure 4-3 shows the distribution of communal land and private property where permissions have been obtained and those under negotiation against the boundaries of mining concessions.  MPR has purchased surface lands covering the known extents of the Cerro Los Gatos, Esther and Amapola Resource areas totaling 5,478.9 ha as shown in light blue in Figure 4-3.

Figure 4-3:  Surface Rights and Exploration Permissions (MPR)

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MPR has negotiated and ratified access agreements with the two blocks of communally owned land belonging to the Ejidos of La Esperanza shown in purple on Figure 4-3 and San José del Sitio shown in light green on Figure 4-3.  These agreements allow access through the surface land holdings of both communities and are registered with the federal land registry.  MPR has worked together with both communities since the inception of the exploration project and has enjoyed strong support from the community leaders and general assemblies.

MPR has established a multi-disciplinary team to communicate the progress of the project to both La Esperanza and San José del Sitio, as well as the surrounding communities and municipalities.

4.3  Environmental Permitting

On December 12, 2016 MPR filed a MIA to SEMARNAT, the environmental authority in Mexico.  According to the Mexican regulations, the Project “Los Gatos” contemplates the execution of activities considered risky, which must be in compliance with the provisions of Article 147 of the Ley General del Equilibrio Ecológico y la Protección al Ambiente (LGEEPA) (General Law of Ecological Equilibrium and Environmental Protection).  To comply with this regulation, an environmental risk study has been filed along with the MIA.  The Environmental Unique License was approved May 27, 2019.

On July 17, 2017 SEMARNAT approved the MIA-Regional with document No. SGPA/DGIRA/DG/05121-2017 for mining exploitation project in Satevó, Chihuahua to develop work and activities (site preparation, construction, operation and others) needed to develop mining exploitation.  The approved area to be directly impacted by works and activities is 211.084 hectares.  The authorized permit period is for 24 years (from July 17, 2017 to July 17, 2041) starting the day of notification of this resolution, with the possibility to extend for a period similar to the authorized.

On September 4, 2017 SEMARNAT approved the MIA-Particular with document No. SG.IR.08-2017/251 to develop work and activities to construct and operate the project named Línea Eléctrica 115 KV Los Gatos, which consists of the opening of an 18-m wide corridor and 58.0 kilometers, which affects an area of 105 hectares to collocate 249 posts to support the power lines in the municipalities of San Francisco de Borja and San Javier Satevó, Chihuahua.  The authorized permit period is for 20 years (from September 4, 2017 to September 4, 2037) starting the day of notification of this resolution.

According to article 28, section VII of the LGEEPA, land use changes in forested areas as well as in jungles and arid zones are subject to the environmental impact assessment.  The construction of mining and support infrastructure required by the Mining Project “Los Gatos”, implies a CUSTF (Change of Land Use) in an area of 390.37 ha covered by desert microphyllous scrubland, according to Article 14 of the environmental impact assessment regulations, and the information on land use change is included in the MIA.

On November 1, 2017 the SEMARNAT approved the Estudio Técnico Justificativo para Ejecutar el Cambio de Uso de Suelo de Terrenos Forestales with document No. SG.CU.08-2017/310 to remove forest vegetation and fertile soil in an area of 390.6972 hectares to prepare the site and construct the infrastructure required for the mining exploitation of Los Gatos project.  The authorized permit period is for 36 months (from November 1, 2017 to November 1, 2020) with the possibility to extend the permit period.

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4.4  Environmental Liabilities

The author is not aware of any environmental liabilities to which the property is subject.

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5.0  ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1  Accessibility

San José del Sitio is accessible by an improved gravel road from the turnoff of Federal Highway 24 at the KM 81 marker between the cities of Chihuahua and Hidalgo del Parral.

The access road from Chihuahua, Chihuahua, Mexico has been recently paved.  A portion of the road from San José del Sitio has been rerouted to the mine site to minimize interference with the drainage that runs in the valley where the Cerro Los Gatos deposit is located.

5.2  Climate, Vegetation, Soils, and Land Use

The project area is in the Sierras y Llanuras del Norte Physiographic Province near the boundaries between the Gran Meseta, Cañones, and Sierras, and Llanura Tarahumara Sub Provinces.  The general physiography of the Los Gatos area is characterized by low to middle rolling volcanic hills with local escarpments and flat valley floors.  Altitudes vary with between 1,550 masl at the base of the Santo Toribio Creek and 1,780 masl at the top of the Los Gatos Hill, one of the highest peaks of the project area.

The climate of the Los Gatos project is typical of desert areas of Northwest Mexico, extreme semi-arid, with a maximum temperature in the order of 41.7°C and a minimum recorded at minus 14°C; annual average temperature is 18.3°C.  Annual rainfall averages 363.9 millimeters (mm) over an average of 61 days, mostly during the rainy season of June through September, and relative humidity is 50%, with a dominant Northeastward wind.  There is abundant sunshine and little cloud cover during most of the year.  Snow is a rare occurrence in Southern Chihuahua but has been recorded on occasion.  Exploration and mining activities are seldom interrupted by adverse weather conditions, except for short-lived storms producing floods and damage to access roads.

Vegetation is characterized by a semi-desert landscape, with typical low brush vegetation in the slopes including lechuguilla, ocotillo, sotol, yucca, sage, bear grass, and other types of indigenous grasses.  Larger brush and trees are common along the main watercourses, with the presence of oak, cypress, cottonwood, poplar, huizache, and mesquite, among others.

The soils of the area are sandy to rocky and are composed of detrital material from the local volcanic and sedimentary rocks classified as lithosols and yermosols.  The lack of flat areas with regular water sources and good soils results in only small areas useful for crops, but there is sufficient growth of native grasses and desert plant life to support the principal economic activity of the region, cattle grazing.  Land tenure in the municipality of Satevó is 25% communal (Ejido); and 52% private property, with predominantly cattle grazing and other agricultural use.

Locally, the surface lands are mostly owned by private individuals as small cattle ranches, with average sizes of 1,000 to 2,000 hectares.  Many of these ranches are unimproved grazing lands with no structures; however, a few ranch houses exist in the scattered areas.  Some landowners live locally in the community of San José del Sitio or surrounding communities, while others live in the surrounding cities of Zaragoza, Parral, and Chihuahua.

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Two communities hold parcels of surface lands as agrarian communes or “Ejidos.”  These are the communities of La Esperanza and San José del Sitio, which have corporate ownership of their respective surface lands (Behre Dolbear, 2011).

5.3  Personnel

As a part of the company’s commitment to adding value to the local communities and building capacity in Mexico, over 99% of the operations workforce is from Mexico.  Most of the workforce is being sourced from local towns, with skilled labor from the nearby cities of Chihuahua and Hidalgo del Parral.

5.4  Infrastructure

Water Resources in the region are mostly related to the Conchos River Basin, including the San Pedro, San Francisco de Borja, and Satevó River Sub-Basins.  A larger supply of surface water is associated with the Conchos River, located 7 km to the South of the main exploration areas.  The Conchos River is dammed in several locations, including La Boquilla, a major hydro power plant in the region.  Scattered ranch houses within the project area are normally serviced by generators and small wells or capture ponds from surface runoff waters.

Locally, there are significant groundwater deposits.  Water from dewatering and mine inflows are being cooled and used for onsite purposes.  Water is reclaimed from the tailings storage facility for use in the process facility.  MPR has conducted detailed hydrogeologic studies over the entire area.

Power to the Cerro Los Gatos Mining project site is supplied via a 115-kV utility transmission line.  This originates from the ‘San Francisco de Borja’ substation in Satevó (Chihuahua), where a 115-kV connection has recently been installed.

A camp and supporting facilities have been constructed onsite for workers and contractors.

5.5  Population Centers

The mine is located nearby San José del Sitio, categorized as a Municipal Section of the Satevó Municipality.  It is a community of approximately 264 persons, with electrical and water services, elementary school, and basic health services available.  Regular daily bus services connect the town with the capital city of Chihuahua.  The city of Valle de Zaragoza, 45 km to the East-Southeast of San José del Sitio, located on Federal Highway 24, 35 km to the South of the turnoff to San José del Sitio, is the nearest significant commercial center.

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6.0  HISTORY

The Los Gatos project area is considered an underexplored region with only small showings of precious and base metals mineralization.  It has been the subject of only very limited historical prospecting and mineral exploration.  Various maps from the Servicio Geológico Mexicano (SGM — Mexican Geologic Survey) and its predecessors show prospect locations at the Esther, Gavilana (Paula), and San Luis zones with references to the occurrence of silver, lead, and zinc.  Reports in the archives of the SGM reference field review of the Tren and Margarita prospects (Baca Carreón, 1964) in the Northwest part of the Los Gatos concession as well the Santa Rita prospect (Ramírez, 1976), located to the East of San José del Sitio outside the Southeastern limit of MPR’s concessions block.

There are small prospect pits and minor workings in the Esther, San Luis, Tren/Margarita, and Paula zones.  Local verbal accounts suggest that most of this development occurred in the period of 1920-1950, although there is reference to minor work at the Esther zone as recently as the 1960’s.  From the limited development observed in each of these zones, it is unlikely that there was ever any record keeping of production.  Surface work by MPR has not uncovered any evidence of modern prospecting activities in the area such as drill hole collars, survey points, or earlier sample locations.

The project was initially recognized by reconnaissance activities by Perry Durning and Frank Hillemeyer of La Cuesta International Inc. (La Cuesta) in 2005 while under contract with Silver Standard Resources. La Cuesta applied for the original Los Gatos concession in 2005 and recommended the target to Silver Standard for acquisition.  Silver Standard geologists visited the project in 2005 but rejected further work and freed La Cuesta to promote their project to other entities.

Los Gatos Ltd. (LGL) (then parent of MPR prior to the merger into Sunshine Silver Mines Corporation in 2011) was contacted later in 2005 and the project was visited by LGL representative Jon Gelvin.  An initial letter of agreement for exploration work on the project was negotiated between La Cuesta and Los Gatos Ltd. in 2006, and a final contract was ratified in April 2006 between MPR and La Cuesta.  Only minor field work was conducted during 2006-2007 on the Los Gatos project during the waiting period for the initial concession to be titled (P. Pyle, 2010).

Considering the lack of important mine workings and previous drilling or any other direct exploration, there are no other records of historical Mineral Resource and Mineral Reserve estimates.  Any silver, lead, and zinc production that might have been carried out from the Esther, Gavilana (Paula), San Luis, Tren, and Margarita prospects was probably limited to a few hundreds of tonnes with irregular silver and lead-zinc concentrations (Behre Dolbear, 2011).

6.1  Historic Resource Estimates

No historic Resource estimates were completed before MPR’s involvement with the project.

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7.0  GEOLOGICAL SETTING AND MINERALIZATION

7.1  Regional Geological Setting

The Los Gatos project is located in the transition zone between the Sierra Madre Occidental volcanic province of Western Mexico and the Mesozoic Chihuahua basin, largely sedimentary, to the East.  It is also located in the general contact zone of the Sierra Madre Occidental (SMO), Chihuahua, and Parral Tectonostratigraphic Terranes Figure 7-1.

Figure 7-1:  Geologic Provinces Map Showing the SMO, Tectono-Stratigraphic Terranes
(SGM, Campa and Coney, 1987; USGS, Hammarstrom et al, 2010)

The zone extends from the younger Trans-Mexican volcanic belt in the state of Jalisco in central Mexico Northward through the states of Durango, Sinaloa, Chihuahua, and Sonora and partially into the Southwestern United States.  It is largely characterized by a thick sequence of Tertiary volcanic rocks that are generally dissected by a strong North-Northwest bearing fault system that divides the area into the plateau and barranca sections.  The sequence of volcanic rocks is subdivided in two major units, the Lower Volcanic Group and the Upper Volcanic Group.

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Lower Volcanic Group (LVG): Characterized by a predominant pile of andesitic volcanoclastic rocks that characteristically outcrop at the bottom of the deep barrancas.  The group is generally massive in nature and shows extensive propylitic alteration, commonly due to the alteration effects of coarse-grained to porphyritic intrusive rocks.  Even though the volcanism is predominantly andesitic, the upper parts, toward the contact with overlying volcanic, tend to become more felsic, and thick beds of rhyodacite and rhyolite and are found intercalated with andesite and dacite.

Upper Volcanic Group (UVG): Characterized by a thick sequence of felsic volcanoclastic rocks, predominantly ignimbrites, that shows well-defined bedding and tuffaceous horizons.  These rocks form most of the high scarps and cliffs that characterize this province.  Normal extension faulting creates a series of large, gently dipping blocks with almost no signs of alteration.  Volcanism in this area started 30 to 32 million years ago with the beginning of a sudden bimodal, calcalkalic event.

The SMO Province is one of the largest known epithermal, precious-metal metallogenic provinces and is host to well-known gold-silver producing mining districts, including Concheño, Ocampo, Batopilas, San Dimas-Tayoltita, Topia, Guanaceví, and Bacís and recent discoveries such as Mulatos, La Ciénega, El Sauzal, and Pinos Altos.  It has been well established that most of these districts are enclosed within the LVG sequence, although few of them show mineralization transecting up to the base of the UVG, as evidenced at La Ciénega and Ocampo.

The oldest rocks of the area are Mesozoic (Cretaceous) aged sedimentary rocks belonging to the Chihuahua Platform.  Predominant silver-lead-zinc mineralization characterizes this region, and mineralization is commonly related to skarn, limestone replacement, and Mississippi Valley type deposits. Some of the best examples include Santa Eulalia, Naica, Bismark, San Martín, Velardeña, La Negra, La Encantada, Concepción del Oro, Charcas, and many others.

Volcanic rocks of the SMO volcanic province were erupted upon an irregularly folded and faulted surface of these older sedimentary rocks.

Pre-volcanism compression (80-40 Ma) recorded in the underlying sedimentary sequence was followed by North-Northeast extension (<29 Ma) in the region (Ferrari, et al., 2007).  Ferrari, Valencia-Moreno, and Bryan (2007) state that the “Sierra Madre Occidental consists of five main igneous complexes:  (1) Late Cretaceous to Paleocene plutonic and volcanic rocks; (2) Eocene andesites and lesser rhyolites, traditionally grouped into the Lower Volcanic Complex; (3) silicic ignimbrites mainly emplaced during two pulses in the Oligocene (ca. 32-28 Ma) and Early Miocene; (4) basaltic-andesitic lavas that erupted toward the end of, and after, each ignimbrite pulse, which have been correlated with the Southern Cordillera Basaltic Andesite Province of the Southwestern United States; and (5) post-subduction volcanism consisting of alkaline basalts and ignimbrites emplaced in the late Miocene, Pliocene, and Pleistocene, directly related to the separation of Baja California from the Mexican mainland” (P. Pyle, 2010).

The regional geology in the area of the Los Gatos project is shown in Figure 7-2.  The dominant rocks of the Los Gatos project area would be classified as belonging to the Lower Volcanic Group placing the ages of deposition from 39-35.5 Ma (P. Pyle, 2010).

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Figure 7-2:  Regional Geologic Map (SGM, 1:250,000 Original Scale) Property Structural Setting

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7.2  Property Geological Setting

The oldest rocks of the project area are Upper Cretaceous (Cenomanian) to Lower Paleocene aged sandstones, shales, and limestones correlative with the Mezcalera formation, deposited in the limits between the Aldama (limestone) Platform and the Mesozoic Mexican Sea (open basin environment).  These rocks have been locally metamorphosed to phyllites, quartzites, and marbles near areas of igneous activity, including at the Santa Rita skarn area East of San José del Sitio adjacent to the Southeast of the Los Gatos claim block. Rocks of this oldest sedimentary sequence occur within a small horst block located to the Southwest of the Cerro Los Gatos deposit, with prominent high-angle fault boundaries on the North and South, parallel to the regional trend of faulting (Figure 7-2).

A stratigraphic column representing the regional geology is shown in Figure 7-3. Rocks of the series locally referred to as the “Lower Volcanic Series” (Units Tv2, Tv3, and Tv4 from McDowell 2007) dominate the geology of the Los Gatos project area.  The oldest of these units are composed of andesitic lava flows and pyroclastic breccias (Unit Tv2) that were deposited on irregular topographic surfaces and have variable thicknesses.  The general orientation is relatively flat, with a low regional dip to the Southeast of approximately 8 degrees.  There are many exceptions to this orientation in outcrop, however, due to the irregular topography onto which the flows were deposited.  Overlying and interbedded with the older andesitic flows are flows and tuffs of dacitic composition (Tv3).  Volcanoclastic sandstones and sedimentary breccias occur in the Northeastern portions of the Los Gatos project area in fault contact with the andesitic and dacitic flow rocks.  It is possible that these rocks are correlative with the Ts unit of McDowell 2007, which is described as “coarse, generally lithified clastic deposits, associated with Northwest trending linear basins.  These rocks are derived from the andesitic and dacitic local terrain and occasionally contain fragments of hydrothermally altered material and vein fragments.

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Figure 7-3: Stratigraphic column of Los Gatos

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Intruding and deposited on the entire section are locally important rhyolite flows, flow domes, and dikes that are likely correlative with unit Tv4 of McDowell 2007.  These rocks are usually strongly silicified and have all the varied textures expected with the development of flow domes, including breccias, flow banding, and intrusive/extrusive transitions.

Each of the rocks in the section contains observable hydrothermal alteration, suggesting that mineralization in the area probably occurred late in the history of the development of the volcanic section.  It is important to stress, however, that economic grades of mineralization have only thus far been identified in the andesite and dacite sections (Tv2 and Tv3) (P. Pyle, 2010).

The Los Gatos District hosts a series of quartz, quartz-calcite, and calcite veins in at least fifteen separate vein systems that are exposed along a strike length of approximately 30 km and an outcrop belt width of approximately 5 km.  Vein width is generally in the order of 1 m with local wide zones up to 8 m.

Structurally, the veins form two sets with North and Northwest strikes and mostly steep dips, consistent with formation as oblique-slip normal faults.  Slicken line rakes indicate dominant normal faulting but with some significant dextral-slip components.  A structural model is proposed in which veins formed in dextral-normal faults, with North-striking veins predicted to be thicker with dominant normal slip kinematics and Northwest-striking veins predicted to have oblique-slip kinematics.  A dextral component is consistent with horsetail structures (e.g., Mezcalera Vein) and dilatant bends and jogs (cymoid loops; e.g., San Luis system).  District-scale East dip of volcanic units suggests a master normal fault to the East hidden below conglomeratic cover and a potential exploration target; geophysical techniques may be useful in such exploration (E.P. Nelson, 2007).

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On a regional scale, both West-Northwest and North-Northwest trending structures are very common.  The West-Northwest trending fault and fracture zones probably reflect reactivated basement structures, whereas many of the North-Northwest trending faults probably are associated with basin-and-range extension. In the region, epithermal mineralization is associated with both phases of extension, so both trends have exploration potential; however, the basement architecture has overall control on the distribution of magmatic centers and hydrothermal systems.  This can be seen in the deflection of later basin-and-range faults and Laramide fold-thrust structures into the pre-existing West-Northwest structural corridors. Because of this, the main West-Northwest trending fault zones are considered to be higher-priority targets (T. Starling, 2010).

7.3  Mineralized Zones

Mineralization at the Cerro Los Gatos deposit is associated with a series of West-Northwest trending veins hosted in volcanic rocks on the footwall side of a listric normal fault contact.  The hanging wall of the fault is comprised of epi-clastic sediments.  Figure 7-4 is a surficial geologic map of the deposit.

Figure 7-4:  Geological Map of the Cerro Los Gatos Deposit Area

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Economic mineralized grades are not present at surface; however, epithermal alteration textures are present and aided in the discovery of the deposit.  The general Northwest trending East dipping Cerro Los Gatos vein system is persistent with a mapped extension in the order of 10 km, true widths of as much as 30 m at depth as demonstrated by diamond drilling, and local associated veining up to 50 m wide.  Banded quartz veins and breccias are cemented by quartz, calcite, and abundant manganese oxides.

A study based on geological characteristics and silver-lead-zinc (arsenic-antimony-mercury) anomalous sections of the vein resulted in the discovery of the Cerro Los Gatos listric-shaped mineralized horizon hosting steeply to shallowly dipping mineralized-shoots at depth.  Mineralization of interest occurs for approximately 2,500 m in length, between an elevation varying roughly between 1,200 masl and 1,400 masl through a mineralized vertical extension of between 50 m and 250 m and an estimated average in the order of 200 m.  The reported average drilled width of the structure is in the order of 8.9 m.  It is noted that some sections of the vein required deeper drilling and some holes intersected mineralization of interest.

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Figure 7-5 is a cross section through the center of the deposit, section line 29, showing Ag equivalent (AgEq) in the drill hole traces and the block model.

Figure 7-5:  Cross-Section 29 AgEq

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The top of the mineralized horizon at Cerro Los Gatos is generally located at an elevation of 1,400 masl.  The surface is in the order of 1,570 masl ± 50 masl.

Mineralization of interest and high-grade mineralization have been identified in the different vein systems at the Los Gatos project.  Lead, zinc, and silver have been identified from epithermal quartz veins at the surface and from drilling intersections, while smaller, but important quantities of gold and copper associated with the veins have been intersected.  Anomalous values have, thus far, been identified in the Cerro Los Gatos, Esther, Amapola, Cieneguita, San Luis, Paula, Rodeo, San Agustín, Boca de León, Lince and Mezcalera zones.  Drilling has identified a continuous geometry of the mineralization in the Cerro Los Gatos, Esther, and Amapola zones.

Lead mineralization occurs primarily as galena and lead oxide minerals of varying grain sizes that are disseminated in quartz vein material, as open-space filling in cavities, and as replacements in the andesitic and dacitic flow units.

Zinc mineralization occurs as sphalerite and zinc silicate minerals of variable grain sizes disseminated in quartz vein material, as open-space filling in cavities, and as replacements in the andesitic and dacitic flow units.  Sphalerite ranges from yellow to brown in color and is deposited in a similar style but is not always associated with the galena mineralization.

Silver mineralization occurs as acanthite (argentite) and native silver and has been detected in thin sections as proustite as small inclusions within galena grains.

Copper mineralization occurs as chalcopyrite and occasional native copper disseminated within quartz veins.  Gold mineral species have not been identified visually but are present in small quantities in assay results.

The veins themselves display variable gangue mineralization, depending on the depth of exposure within the epithermal environment.  It is common to observe calcite or manganese oxide mineralization at high levels within the epithermal system, which transitions to barite, fluorite, and quartz at lower levels.  Adularia, albite, and alunite have also been observed within the veins but only in small percentages and usually at high elevation levels.  Within the mineralized portions of the veins, it is common to see quartz with minor fluorite and occasional minor calcite associated with lead, zinc, silver, copper, and gold mineralization.  The veins are typically rhythmically banded on a scale of 1 mm to 10 mm per band, with repeated pulses of quartz carrying the metals and other gangue minerals.  It is common to see multiple pulses of mineralization where small veins crosscut each other.  It is also common to see various coloration of quartz in the multiple pulses, ranging from milky white to vitreous gray to amethystine purple (P. Pyle, 2010).

It is apparent that most of the economic mineral values are associated with sulfide mineralization.  Oxide mineralization is limited at depth, and is commonly related to fracture, breccia zones, and open spaces within the veins.

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8.0  DEPOSIT TYPES

Veins in the Cerro Los Gatos deposit show textures and gangue mineralogy (local chalcedony and calcite, and quartz-replaced lattice texture calcite) that indicate a relatively high-level hydrothermal system in the boiling environment.  Breccia with clasts of vein quartz indicates a protracted hydrothermal system during multiple faulting events, a positive sign for economic epithermal veins.  It has been interpreted that mineralized-ore shoots may extend relatively far down dip, possibly to at least 250 m.

Mineralization at Cerro Los Gatos is characterized by, silver, lead, zinc, and copper sulfides and their corresponding oxides, along with fluorite, manganese, barite, and traces of gold associated with quartz and calcite veins.  The veins vary in orientation from West-Northwest to Northwest to North-Northwest to North-Northeast and vary in thickness from 1 m to 8 m in outcrop but displaying much greater true thickness at depth.  Study of the veins in hand specimens and thin sections suggest they are epithermal in origin and are likely of intermediate sulfidation composition.

The exploration model for these types of veins was put forward in a landmark paper by Dr. Larry J. Buchanan (1981) that set the basis for the understanding and interpretation of epithermal deposits that has been widely used in exploration, Figure 8-2 below.  Dr. Buchanan now serves as a special consultant to SSMRC and has been instrumental in the recognition of the importance of the mineralogy of the veins and the expected transitions of the veins in the sub-surface.

Exploration of epithermal veins at Los Gatos is mainly focused on the interpretation of geological, structural, mineralogical, and alteration features in order to identify those areas within where mineral deposition was most likely to occur due to paleo-boiling surfaces at depth.  These specific levels within the veins are where economic concentrations of lead, zinc, silver, copper, and gold-particularly bonanza-grade mineralization can be expected.  Additional exploration is being targeted to other areas where mineralization can be concentrated, such as in the hanging wall fracture zones to the veins, the flanks of the flow domes, and structural intersections within the vein trends.

Other deposit types in the region suggest that higher-temperature mineralization can also occur, such as the skarn setting identified at the Santa Rita prospect located to the Southeast of the concession block.  These higher-temperature analogues have not yet been identified within the concession block (P. Pyle, 2010).

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Figure 8-1 shows epi-thermal textures encountered in drill hole GA-132 at 392 m down the hole on the left, and GA-175 at 273 m down the hole on the right.

Figure 8-1:  Epithermal Textures in Drill Core

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Figure 8-2:  Idealized Section of a Bonanza Epithermal Deposit (Buchanan L.J., 1981)

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9.0  EXPLORATION

9.1  Historic Exploration

Apart from small pits and workings in the Gavilana (Paula), San Luis, Tren/Margarita, and Esther zones, there has been little to no previous exploration work within the Los Gatos project area.  No evidence of systematic prospecting, sampling, or drilling has thus far been identified.  The only record of historic sample collection is from Baca Carreón, 1964, where 16 samples along the Tren/Margarita zone were taken.

9.2  Recent Exploration

MPR began its first phase of exploration at the Cerro Los Gatos deposit in 2007 with a program of surface geologic mapping and rock sampling covering approximately 60% of the original Los Gatos concession within the core of the claim block.  This work was conducted by a local Mexico-based consulting group, Grupo Azta, and is detailed in a report by J. Islas (2008).

Work by Grupo Azta identified more than 100 km of strike length of quartz and calcite veins, many of which contained anomalous lead, zinc, and silver mineralization.  Grupo Azta collected 1,217 rock samples from surface outcrops of vein and wall rocks.  Of the 1,217 samples, 200 samples contained values in excess of 10 grams of silver per tonne.

As the mapping by Grupo Azta progressed, it became clear that there were several orientations of prospective veins within the Los Gatos project area.  Consultant Eric Nelson was brought in to review the structural geology of the area and make recommendations on the more prospective trends.  The results of his work, contained in Nelson (2007), suggested that the most favorable vein trend was North-Northwest, and an initial program of drilling was proposed (J. Islas, 2008).

MPR expanded its program at the Los Gatos project in June 2008, employing its own local technical staff, under the supervision of Philip Pyle, Jon Gelvin, and Dr. Larry J. Buchanan.  During the months from June 2008 to October 2008, environmental permits were obtained, proposed drill areas were re-mapped and re-sampled, surface access rights were negotiated with local ranches, and drill access roads were constructed.

Drilling began with one rig in October 2008, at the Paula zone, and transitioned in early 2009 to the Cerro Los Gatos zone.  The initial significant identification of silver was from hole GA-04 in April 2009, where 73.6 g/t silver was found over 4 m from 152 m to 156 m depth.  This was quickly followed by significant intercepts in hole GA-06 and then GA-09, which contained 34 m of 414 g/t silver, 2.0% lead, and 4.85% zinc.  At this point in the drilling program, the geometry and the preferred level for mineral deposition was identified, and a series of holes were drilled that Indicated a continuous mineralized body of apparent high grade mineralization with lead, zinc, and silver mineralization over a strike length in excess of 2.5 km, a dip extent in excess of 250 m, and an average thickness of 8.9 m, within the Cerro Los Gatos banded vein complex.

Also, in early 2009, drilling in the Esther zone commenced, with one rig moving back and forth between the Cerro Los Gatos and Esther zones.  Significant mineralization was identified in hole ESO4, with 14 m containing 79.8 g/t silver from 102 m to 116 m depth.  This was followed with significant offsets in

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mineralization holes ES05 and ES06, proving a continuous geometry of mineralization for over 1 km of strike length, with an average thickness of more than 3.4 m and a minimum down dip extent of 200 m.

A decision was made to replace the original drilling rig in June 2009 with one of larger capacity, followed quickly by the addition of a second and third drilling rig during the summer of 2009.

Drilling resumed in connection with underground development at Cerro Los Gatos in August-December 2018 from designed chambers within the underground workings.  The drilling was conducted by Major Drilling with NQ size drilled from the footwall side of the mineralized zone.

Drilling shifted back to the surface in January -July 2019 using Major 5000 rigs from Major Drilling Company.  This work focused on improving confidence in the Resources of the NW and Central zones in areas where underground development had not yet reached.  As in previous surface drilling programs, holes were pre-collared with a tri-cone bit and core collected using HQ size, reducing to NQ size if poor drilling conditions were encountered.

Detailed soil geochemistry programs were conducted over the Esther zone and the area between the Cerro Los Gatos and Esther zones.  Results of the sampling identified new veins in the Esther zone and revealed four separate structures between the Esther and Cerro Los Gatos zones.

Detailed topographic mapping was created using Photosat, a Canadian contractor.  The topography was created at 1 m, 5 m, 10 m, and 50 m contours from Geoeye® satellite coverage captured exclusively for the survey.  Survey control points were established on the surface, with coordinates by total station in order to guarantee the accuracy of the survey.

A detailed 3D IP survey was conducted during July 2010 using SJ Geophysics, a contractor from Canada.  Lines were initially spaced at 100 m with stations every 25 m and later tightened to 50 m by 25 m.  Data from the survey was processed using the UBC inversion algorithms, and the results suggest a correlation between vein mineralization at the Cerro Los Gatos zone and zones of high chargeability and low resistivity.  In addition, the vein mineralization at the Esther zone suggests a similar relationship of high chargeability and low resistivity.  The first holes to test the trends of mineralization from these surveys have successfully extended mineralization in both zones.  As a result of the good correlation with mineralization, extensions of the surveys were begun in November 2010 in both the Cerro Los Gatos and Esther zones.  Additionally, data were collected in the Amapola and San Agustín zones to determine the signature of mineralization in these areas for drilling.

Detailed geologic mapping has been conducted over approximately 60% of the Los Gatos concession utilizing both local staff from MPR and independent contractors.  Regional-scale mapping has taken place on the remaining 40%.

A second review of structural geology of the Los Gatos and surrounding areas was prepared in March 2010 by consultant Tony Starling of Telluris Consulting.  His work suggests a relationship between mineralization and the presence of the younger dome rocks and identifies the younger Northeast trending cross faults as a potentially important conduit for fluid flow during the mineralization phase.  In the report (Starling, 2010) also identifies several other favorable zones from Geoeye, Spot, and Aster imagery, which may serve as loci for mineralization, that are not well exposed in outcrop.

A detailed study of the local geology in the San Agustin zone was prepared during October 2010 and detailed in a report by Byington (2010).  This work suggests a strong preference for North-Northeast trending veins, and a series of drilling recommendations were made.  A test of the drill proposals was made with mostly negative results.

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Geological, exploration, and drilling information has identified a Mineral Resource in three of the targets explored, Amapola, Cerro Los Gatos, and Esther.  MPR produced preliminary Resource estimates internally and through Rowearth Consulting of the Esther and Cerro Los Gatos zones (Behre Dolbear, 2011).

It is the author’s opinion that the samples collected for exploration purposes are representative of project area.  Soils and surface rock chip samples were assayed and used for location purposes only and were not incorporated in Mineral Resource estimates.  Assayed drill core was the only source of grade and thickness data incorporated in the Mineral Resource estimate.

In addition to the Cerro Los Gatos deposit, the Esther deposit and the Amapola deposit, SSMRC has identified nine other mineralized zones defined by high-grade drill intersections in the Los Gatos District.  Grade intercepts from such mineralized zones are shown Table 9-1 below.

Table 9-1:  Mineralized Grade Intercepts

Mineralized Zones	Length(m)	Ag(g/t)	Pb(%)	Zn(%)
Boca de León	2.2	90.6	5.0	0.8
Cieneguita	1.3	62.4	5.4	0.9
El Lince	4.0	62.2	0.0	0.1
El Rodeo	0.8	61.5	3.4	4.0
Los Torunos	1.8	34.2	2.6	0.9
Mezcalera	2.0	59.4	0.1	0.1
La Paula	4.0	180.0	0.1	0.1
San Agustín	1.3	148.0	1.2	2.3
San Luis	2.0	271.0	0.3	0.1

NOTE:  Does not include Ocelote and Wall-E/Ava zones, as sufficient drilling has not been completed at these zones.

9.3  Decline and Bulk Sample

Driving of a 1.3 km decline at a 15 percent grade commenced in July of 2015 and reached the upper extent vein in the Central block in September of 2016.  Figure 9-1 is a photo of the entrance of the decline taken in August of 2015.

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Figure 9-1:  Decline Entrance (Mine Portal)

A 1,215 tonne bulk sample was mined on the 1,400 m level from five vein drift blast rounds each with 4m advance along the strike of the lower vein.  The material was blended and 50 tonnes were sent to SGS in Lakefield, Ontario, Canada for pilot plant metallurgical testing.

Figure 9-2 is a cross-section showing the path of the decline and the location of the bulk sample.  Four drill holes pierced the vein prior to the advance of the decline, samples were collected from the blast rounds.

Figure 9-2:  Decline and Bulk Sample Area Cross-Section Looking Northwest (+/- 200 m)

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Table 9-2 compares the planned grade and tonnage of six drift 4 m rounds on the vein and the round average and the assumed tonnage of five rounds.  The additional drilling and bulk sample suggest that locally the upper extent of the vein is richer in Ag, Pb and Zn than estimated in the model and lower in tonnage.  The vein was encountered where it was modeled and expected.  Drilling from the decline demonstrated Resource grade material exists outside of the model up-dip to the Northwest.  The drill holes and bulk sample data have not been included in the block model and could potentially expand the tonnage and improve the grade in the up-dip area of the Central block if included.

Table 9-2:  Bulk Sample and Block Model Comparison

	Ag g/t	Pb %	Zn %	AgEq g/t	Mass
Model	314	1.70	3.29	500	2,216	(1)
Round 1	1,083	10.97	11.82	1,922
Round 2	739	11.4	11.22	1,571
Round 4	571	5.95	6.86	1,044
Round 5	309	3.41	5.55	642
Round 6	473	4.64	4.73	817
Round Average	635	7.27	8.04	1,199	1,215	(2)
Comparison % Round Average/Model	202	429	244	240	69	(3)

NOTES1: Planned for 6 rounds with 4 m advances;
   2: Bulk sample grades only contain 5 rounds.
   3: Model mass corrected for 5, 243 tonne rounds instead of 6.

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Figure 9-3 shows the bulk sample drilling as well as the blast round AgEq grades compared to the model AgEq grade within the high-grade +150 AgEq g/t shell.

Figure 9-3:  3D Clipped View of Bulk Sample and Block Model AgEq Grades

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10.0  DRILLING

As of July 2019, 484 drill holes relevant to the Cerro Los Gatos Deposit had been completed by MPR, for a total of 150,842 m.  The project database contains drilling on other prospects that are not applicable to this report.  Table 10-1 shows drilling associated with the Cerro Los Gatos deposit and tabulates the drilling by purpose.

Table 10-1:  Drill Hole Count by Purpose

Purpose	Count	Length (m)
Surface Exploration	399	137,220
Underground Exploration	32	3,567
Exploration and Metallurgical testing	6	1,733
Underground Bulk Sample Targeting	4	415
Metallurgical Testing	5	1,693
Geotechnical	18	4,134
Tailings Geotechnical	14	280
Hydrologic Study	6	1,800
Total	484	150,842

Drilling was initiated at the Cerro Los Gatos project in October 2008 and continued until 2012.  Drilling recommenced in 2015 following the joint venture agreement with DOWA, with four rigs simultaneously drilling until February 2016.  Drilling began with a Mexican contractor, Minera Gavilán, but most of the drilling has been completed by Major Drilling Company with Major 5000 rigs.  Drilling is conducted using a wire line rig with diamond core capabilities.  Holes begin with HQ size and are reduced, if necessary, to NQ and very rarely to BQ, if difficult drilling conditions are encountered.  Drilling from the 2015-2016 program were pre-collared with tri-cone bit.

Drilling resumed in connection with underground development in August-December 2018 from designed chambers within the underground workings.  The drilling was conducted by Major Drilling with NQ size drilled from the footwall side of the mineralized zone.

Drilling shifted back to the surface in January -July 2019 using Major 5000 rigs from Major Drilling Company.  This work focused on improving confidence in the Resources of the NW and Central zones in areas where underground development had not yet reached.  As in previous surface drilling programs, holes were pre-collared with a tri-cone bit and core collected using HQ size, reducing to NQ size if poor drilling conditions were encountered.

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Holes are surveyed with a Flexit EZ trac device at 50 m intervals, as the holes are completed.  Accuracy of the Flexit EZ trac is reported by its manufacturer to be 0.25 degrees in calculation of both the azimuth and inclination.  Surveys of hole collar coordinates are completed by a local contract topographer using a Topcon Total Station GTS-236W.  All collar and survey information are stored in a master database in Microsoft Access® (Behre Dolbear, 2011).  Most collars have been cemented and annotated with the drill hole name, Figure 10-1 few monuments have been inadvertently destroyed by vehicles.

Figure 10-1:  Drill Hole Collar Monument

Early on drill holes were positioned to intersect the vein at nearest to perpendicular to strike and dip as possible however, recent infill drilling has utilized shared drilling pads to limit surface disturbance and preparation.  Figure 10-2 shows the location of drill holes.

An additional 255 drillholes have been completed within the Los Gatos project, but outside of Cerro Los Gatos deposit.  This comprises an additional 108,218 meters of drilling, which are not included in the Resource estimate that is the subject of this report.  The totals of the entire project are 739 holes for 259,060 meters.

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Figure 10-2:  Drill Hole Location Map

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11.0  SAMPLE PREPARATION, ANALYSES AND SECURITY

Data summarized in this section and utilized for Resource estimation has been collected by MPR. The sample preparation, analyses and security procedures implemented by MPR meet standard practices in most cases.  Refinements to several protocols are required to achieve the best possible quality control (QC) of sampling.  The data collected is of adequate quality and reliability to support the estimation of Mineral Resources.

No historic sampling by previous operators has been utilized by MPR and has not been described in this section.  Descriptions and quantities of samples are limited to drilling within the immediate Cerro Los Gatos deposit Resource area (project ID “GA” in the database); various surface sampling and drilling outside of deposit area are not considered relevant to this section.

11.1  Sample Preparation

Diamond drill core is transported from the rig to the core preparation site in the town of San José del Sitio, by truck.  Following geotechnical logging by field assistants, geologists log the core and select sample intervals.  Sample intervals are selected only where the geologist anticipates mineralization to exist. In practice, the core is extensively sampled above the hanging wall and below footwall on either side of the mineralized zone.  Samples are constrained to a minimum length of 80 cm and maximum of 10 m.  The mode sample length is 2 m, 79% of samples are 2 m in length, 0.5% are greater than 2 m and 20.5% are less than 2 m.

During the process of sample selection, the geologist draws a centerline to guide the core cutter.  A sample sheet is provided to the core cutter containing sample numbers and from, to intervals.  In addition to a cut sheet, a sample tag booklet system is used.  The booklet contains two removable tags, one is stapled to the plastic corrugated core box and the other is placed in the sample bag along with the sample, the book is retained.  The sample ID is also transcribed on the white plastic core box using a red marker as shown in Figure 11-1.  Sample numbering begins where the previous sample batch left off.

Figure 11-1:  Sampled Drill Core

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11.2  Security

Core preparation is completed in a converted residential double lot in the town of San José del Sitio.  The buildings and fence gates are locked at night.  Sample batches waiting to be transported to Chihuahua City are stored in a secure building adjacent to the core logging area.  When each hole is completed, samples are transported to Chihuahua City, typically three times a week, where they are immediately delivered to ALS Chemex’s preparation facility in Chihuahua City (ALS Chihuahua).  A Chain of Custody form is used to track the samples once they leave the security of the core facility in San José del Sitio.  Only project level staff are involved with the selection, preparation and delivery of samples to the laboratory.

11.3  Analyses

Samples are prepared by drying, crushing, splitting, and pulverizing at ALS Chihuahua and pulps are shipped to Vancouver, British Columbia, Canada (ALS Vancouver) for analysis. ALS Chemex is independent of Sunshine and MPR and is ISO 17025 accredited, the accreditation of ALS Vancouver encompasses preparation processes completed at ALS Chihuahua.

Samples are initially analyzed for Ag, Pb, Zn, Cu, and 37 additional elements using aqua regia inductively coupled plasma — atomic emission spectroscopy (ICP41) with re-run for values exceeding 100 g/t Ag, and 1% Pb, Zn, Cu analyzed by ore grade aqua regia inductively coupled plasma — atomic emission spectroscopy (OG62).  Values further exceeding 1,500 g/t Ag are re-run using fire assay with gravimetric finish (GRA21)

Samples are also initially analyzed for Au using fire assay with atomic absorption spectroscopy finish (AA23) with a re-run for values exceeding 10 g/t Au using fire assay with gravimetric finish (GRA21).

Analysis flow is further shown in graphic form in Figure 11-2.

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Figure 11-2:  Sample Analysis Flow Diagram

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11.4  Quality Assurance and Quality Control for Sample Analysis

MPR’s quality assurance (QA) measures involve the use of standard practice procedures for sample collection as described above; and include oversight by experienced geologic staff during data collection.  QC measures implemented by MPR include in-stream sample submittal of standard reference material, blank material and field duplicate sampling.

11.4.1  Quality Control Sample Performance

QC sample performance is tracked by a dedicated-on site database manager by visually comparing results returned from the lab to control samples in an established work order.  In few instances, issues were cited and re-run by the laboratory.  QC sample performance trends over time are not reviewed on site but is recommended that they be graphically tracked on a continual basis for Ag as well as Pb, Zn, Cu and Au.

QC performance reviewed as part of this report indicate:

·     Standard performance is good, with most initial failures attributable to clerical errors and true failures from certified values near the detection limit.  Clerical errors once discovered should be corrected as soon as possible;

·     Blank performance for Ag is good, but poor performance for Pb, Zn, and Cu should be investigated.  Disparities between Ag performance and the poor performance of Pb, Zn, and Cu is related to the proportion of the detection limit to relevant Resource grade.  Uncertified blanks could contain base levels of Pb, Zn and Cu several times greater than the detection limit, or samples are possibly being contaminated during preparation or analysis and can further complicate investigation.  The use of non-certified blanks has been discontinued.  Observed levels of possible contamination are too low to have a material effect on the estimation of Resources.  This is supported by good performance seen for standards, significant contamination would likely have biased standard analyses as well, and this was not observed; and

·     Field duplicate performance is reasonable, but protocols should be altered to test intervals within the high-grade portions of the mineralized zone more often.

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12.0  DATA VERIFICATION

The following describes steps taken by the qualified person, Tetra Tech personnel, and previous authors to verify data provided by MPR.  The author visited the site November 29-30, 2018, and August 20-21, 2019.  A Tetra Tech staff member audited the project database and visited ALS Chemex Vancouver on August 29, 2012.

Data verification conducted during the 2015 site visit included:  observations of drill hole collar locations and orientations, inspection of drill core compared to logs and analytical results, observations of the drill rig and core collection, observations core intake, visits to outcrops, discussions with MPR geologist including reviews of working maps and cross-sections.  Data verification by the qualified person indicates the data collected by MPR meets industry standard practice and is enough to support the estimation of Mineral Resources.

12.1  Check Sampling

Confirmatory sampling of drill core was completed by Tetra Tech during a site visit in July 2012.  Thirty-four previously sampled half cores were quartered and bagged under the supervision of Tetra Tech and sent to ALS Chemex to undergo the sample analysis regiment used by MPR.  Seven check samples collected by Behre Dolbear in 2009 as part of their data verification efforts were added to this comparison.  Figure 12-1 compares the original sample database values and the check samples in a box and whiskers plot.

Figure 12-1:  Check Sample Scatter Plot

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The box and whisker plots show the means, medians, and population distributions are reasonably similar for Ag, Pb, Zn, Cu and Au.  Additionally, the 41 samples are compared in Figure 12-2 as a scatter plot, with the original sample database values on the x-axis and the check samples on the y-axis.  The scatter plot shows reasonable reproducibility but does contain large differences for a handful of samples which is typical of duplicate sampling.  Both figures demonstrate the original and check data are suitably similar for the purposes of independent qualified person verification.  This spot check verification is not a substitute for umpire sampling recommended in Section 11.0 of this report, which should be made part of MPR QA/QC process.

Figure 12-2:  Check Sample Scatter Plot

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12.2  Database Verification

Verification of the database involved the following:

·Drill hole database validation checked for from and to overlaps, excessive drill hole deviations, missing intervals, and missing holes from drill hole interval files;

·Comparison of the section interpretations to the geologic logs;

·A comparison of the collar elevation to the topographic surface;

·Comparison of standard reference material to the normal intervals to determine if any mislabeled standards were present in the database;

·Spot checks of the assay certificates to the database conducted in 2012 database audit; and

·Collar location spot checks by handheld GPS, Table 12-1.

Table 12-1:  Collar Verification by Handheld GPS

Hole ID	Easting	Northing	Check Easting GPS	Check Northing GPS	Absolute Difference (m)
GA-23	368,554	3,047,578	368,549	3,047,582	6.8
GA-150	368,553	3,047,578	368,549	3,047,582	5.2
GA-154	368,553	3,047,578	368,549	3,047,582	5.2
GA-263	368,551	3,047,578	368,549	3,047,582	3.9
GA-167	369,439	3,047,228	369,431	3,047,233	9.9
GA-170	369,439	3,047,229	369,431	3,047,233	9.8
GA-140	368,638	3,047,548	368,630	3,047,556	11.1
GA-142	368,638	3,047,548	368,630	3,047,556	10.9
GA-145	368,638	3,047,549	368,630	3,047,556	10.8
GA-76	368,618	3,047,562	368,614	3,047,563	4.4
GA-183	368,642	3,047,596	368,635	3,047,597	6.4
GA-185	368,642	3,047,597	368,635	3,047,597	6.5
GA-180	368,642	3,047,596	368,635	3,047,597	6.2

12.3  Metallurgical Sampling

In April of 2016, SGS received a list of 21 variability samples selected by Sunshine Silver Mining & Refining Corporation that had been submitted to the SGS lab in Vancouver for metallurgical and hardness testing.  A review of the samples selected was completed by Amanda Landriault, P.Geo. to evaluate how geologically representative the samples were compared to the material in the mine plan.  The goal of the sample selection review was to identify any gaps or fatal flaws in the samples that were sent for analysis.  This review did not include a review of the number of, or types of tests completed on the samples.

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The conclusions and recommendations of the study illustrate that samples were well selected to respect various grades, alterations, and host rock.  No apparent metal domaining or zoning exists in the deposit to date.  Most of the samples fall within the bulk of the grade distributions and sampling of the higher-grade material has been accounted for.

SGS is of the opinion that some of the material just above the AgEq 150 ppm cut-off, for example two samples with AgEq values between 175-200 ppm, should be included in future metallurgical sampling.

The compositing recipe instructions were provided to SGS by DOWA and did not systematically include dilution material at the sample limits.  Five of the 21 samples contain sufficient dilution that could provide an idea of the metallurgical behavior of material containing wall rock.  It is critical to include wall rock surrounding the mineralized zones to study how the host rock behaves during processing and how it could affect metal extraction.  Additionally, the wall rock surrounding mineralized structures contains alteration and possibly low-grade ore which influence the comminution and flotation.  A selective mining unit (SMU) must also be considered due to the variable thickness of the veins; specifically, the smallest reasonable unit of measurement possible during the blasting and excavation process.  Proximal veins could be mined together depending on the SMU, in which case additional country rock will invariably be included.  Dilution will be refined throughout the mine life based on the mining conditions and methods.

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13.0  MINERAL PROCESSING AND METALLURGICAL TESTING

This section summarizes testwork completed by SGS Vancouver Metallurgy on 2015 test program samples originating from Los Gatos Project, Mexico.  The objective of this program was to support the feasibility study.

13.1  Samples

In total, 21 variability composites were assembled per instructions provided by MPR.  A master composite was prepared from the first 13 variability composites (Var 1-Var 13) by a 75% split-out.  The 75% splits were combined as master composite.  The remaining 25% split of Var 1-Var 13 and the 100% of Var 14-Var 21 were blended separately.  The weights of master composites and variability composites for metallurgical testwork are summarized in Table 13-1.  A total of 467 kg of master composite was prepared.

Table 13-1:  Weights of Master Composite and Variability Composites

CompName	Start Weight,kg	Master CompositeSplit 75%,kg	VariabilityComposite Split, kg
Vari 1	25.9	19.4	6.5
Vari 2	106.8	80.1	26.7
Vari 3	26.1	19.6	6.5
Vari 4	79.4	59.6	19.9
Vari 5	21.6	16.2	5.4
Vari 6	13.2	9.9	3.3
Vari 7	34.2	25.6	8.5
Vari 8	25.2	18.9	6.3
Vari 9	100.9	75.7	25.2
Vari 10	50.4	37.8	12.6
Vari 11	31.0	23.2	7.7
Vari 12	35.5	26.6	8.9
Vari 13	72.9	54.7	18.2
Vari 14	89.4		89.4
Vari 15	57.9		57.9
Vari 16	21.3		21.3
Vari 17	21.7		21.7
Vari 18	26.4		26.4
Vari 19	13.5		13.5
Vari 20	12.1		12.1
Vari 21	23.4		23.4
Master Composite		467.3

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The head assay results are summarized in Table 13-2.  The master composite assayed 0.1% Cu, 2.38% Pb, 5.27% Zn, 275 g/t Ag, 3.48% S and 2.86% F.  Silica is the primary gangue mineral ranging from 48.1% to 77.6%.

Table 13-2:  Head Assays of Flotation Composites

AnalyteUnit	Cu%	Pb%	Zn%	Fe%	Aug/t	Agg/t	S%	F%
Vari 1	0.07	5.38	3.47	2.79	0.52	206	3.26	2.45
Vari 2	0.14	1.95	5.12	2.95	0.60	566	3.35	3.14
Vari 3	0.12	1.90	6.99	2.92	0.47	163	4.11	2.02
Vari 4	0.11	1.09	2.65	2.54	0.41	235	1.59	2.87
Vari 5	0.05	2.38	2.72	2.14	0.20	88	3.06	3.48
Vari 6	0.15	2.28	5.00	5.07	0.13	128	4.97	1.43
Vari 7	0.06	1.63	3.74	1.90	0.14	133	2.05	3.04
Vari 8	0.07	1.16	3.90	2.79	0.26	67	2.51	2.03
Vari 9	0.13	2.94	7.65	4.89	0.18	311	3.85	3.74
Vari 10	0.12	4.73	9.50	3.66	0.25	310	7.26	4.29
Vari 11	0.15	3.70	6.79	3.08	0.20	185	4.52	4.25
Vari 12	0.07	1.54	5.03	3.32	0.51	161	3.91	3.02
Vari 13	0.05	1.61	3.71	2.95	0.46	83	3.19	3.10
Vari 14	0.10	1.11	4.81	4.18	0.12	90	3.58	0.70
Vari 15	0.10	3.66	6.28	4.99	0.26	183	5.14	1.38
Vari 16	0.13	6.42	9.37	3.72	0.57	1100	6.69	1.51
Vari 17	0.09	0.97	3.33	3.12	0.49	620	2.74	1.71
Vari 18	0.11	2.19	5.75	4.55	0.10	100	4.02	2.18
Vari 19	0.12	2.58	7.43	7.35	0.23	109	4.40	3.92
Vari 20	0.08	1.43	2.10	3.14	0.10	58	1.01	4.45
Vari 21	0.10	3.73	2.93	4.11	0.16	113	1.63	5.94
Master	0.10	2.38	5.27	3.56	0.51	275	3.48	2.86

13.2  Mineralogy

The identified recoverable minerals are silver minerals, galena, and sphalerite; the pyrite content varies from 0.34-4.45% with an average of 1.78%.

Quartz has been identified as the primary gangue mineral, ranging from 47.9 to 71.7%.  The samples also contain significant amount of feldspar (0.87-22.9%), mica (0.6-6.19%), chlorite (1-11%) and fluorite (1.9-11.2%).

Lead is present mainly as galena and lead oxide minerals are also identified in some of the samples; elevated oxide levels are contained in samples received from the Southeast zone.

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Zinc is present primarily as sphalerite with significant occurrences of willemite (Zn2SiO4) and other zinc silicates (as alteration phases) in all samples.  The combined amount of willemite and other zinc minerals ranges from 0.6-7.24%, but some samples contain as much as 6.93% of willemite and other zinc silicates (Var 9).  Zinc deportment should be studied and understood well as this parameter could influence the mine production plan and may impact cash flow models.

According to the SGS dataset, generated using QEMSCAN technology as displayed in Figure 13-1, galena and sphalerite are both very finely grained.  Fine primary and regrind sizes would be anticipated to facilitate separation of the two minerals into clean flotation concentrates.

Figure 13-1:  Cumulative Grain Size Distribution (Graph from SGS)

QEMSCAN analysis was performed on the master composite and 21 variability samples.  Process mineralogy was also conducted on the flotation and gravity products.  This section summarizes the QEMSCAN results.  The detailed results are presented in “A Feasibility Level Metallurgical Study of the Los Gatos PJ Project.”

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13.3  Comminution

A summary of the grindability results is provided in Table 13-3.  The number of samples tested per zone is also provided.  The following tests were performed on Los Gatos samples:

·JK Drop-weight

·SAG Mill Comminution, SMC

·SAG Power Index, SPI

·Bond Rod Mill Work Index, BRWI

·Bond Ball Mill Work Index, BBWI

·Abrasion Index

Table 13-3:  Comminution Test Results Summary

		JK Parameters						SPI		BWI	BondAi
	Relative Density (2)	A x b (1)	ta (1)	SCSE (1)(kWh/t)	A x b (2)	ta (2)	SCSE (2)(kWh/t)	CEETCi	SPI(min)	BWI(kWh/t)	Ai(g)
Average	2.79	62.4	0.51	8.30	53.6	0.50	8.87	8.7	73.2	17.4	0.661
Std. Dev.	0.09	—	—	—	7.8	0.06	0.44	3.4	14.6	1.5	0.065
Rel. Std. Dev.	3	—	—	—	15	13	5	39	20	8	10
Minimum	2.68	—	—	—	77.6	0.70	7.66	16.0	46.7	15.3	0.598
10th Percentile	2.70	—	—	—	61.6	0.56	8.40	12.6	59.0	15.4	0.604
25th Percentile	2.71	—	—	—	57.4	0.54	8.55	11.5	62.9	16.1	0.609
Median	2.77	—	—	—	51.2	0.47	9.02	7.8	71.3	17.5	0.630
75th Percentile	2.87	—	—	—	47.7	0.45	9.25	5.7	82.2	18.4	0.719
90th Percentile	2.92	—	—	—	45.8	0.43	9.32	5.4	96.0	18.9	0.741
Maximum	2.93	—	—	—	45.0	0.43	9.36	4.5	102.1	20.6	0.775

(1) from DWT

(2) from SMC; the ta value reported as part of the SMC procedure is an estimate

* Minimum and Maximum refer to softest and harder for the grindability tests, respectively

Based on these results, the ore appears to be softer per JKTech database of Drop-weight test and SMC test; the ore also appears to be of moderately soft to moderately hard per SGS database with respect to SAG milling and appears to be of moderately hard to hard with respect to ball milling.  Based on the Ai, the ore is categorized the samples from abrasive to very abrasive.

13.4  Flotation Testing

Flotation tests were performed on the master composite and 21 variability composites.

Due to the improved marketability of the separate lead and/or zinc concentrates over a bulk concentrate; a sequential flowsheet was used for the design of the plant.  The following subsections refer to the development of that sequential flowsheet.

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A total of 44 flotation tests including five locked cycle tests were completed on the master composite.  The conditions developed in 2014 program were used as a baseline.  A number of factors including primary grind size, reagent dosage and flotation time were investigated in this test program.

13.4.1  Effect of Primary Grind Size on Flotation

13.4.1.1  Lead and Silver Flotation

Six lead and zinc rougher flotation tests were performed at varying primary grinds and the results of lead recovery and zinc misplacement are presented in Figure 13-2.

Figure 13-2:  Effect of Grind Size on Lead Flotation in Rougher Tests

F1, F2 and F3:  These three tests were performed at a k80 of 73, 102 and 125 microns, respectively.  No significant difference between these selected size distributions was observed. The lead rougher recoveries were high at 89.7-90.4% at a zinc misplacement of 13.9-15.4%.

F4 and F5:  The two tests were performed at a k80 of 169 and 205 microns, respectively.  The lead recoveries were approximately 2% lower when compared to the tests F1-F3 with a zinc misplacement of 14.5-15.5%.

F11:  The test was conducted at a k80 of 46 microns and produced the best lead flotation results.  Finer primary grind produced higher lead recovery.

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The relationship between silver recovery and rougher concentrate mass pull is presented in Figure 13-3.  There was a good correlation between silver recovery and primary grind size.  Higher silver recovery was obtained at finer grind sizes and the 46-micron grind produced the highest silver recovery under tested conditions.

Figure 13-3:  Effect of Grind Size on Silver Flotation in Rougher Tests

The primary grind size on lead and silver flotation was further confirmed by cleaner tests as shown in Figure 13-3.  Those tests were performed at same ZnCN addition level (500 g/t in rougher and 230 g/t in cleaner).  Apparently, primary grind size plays a critical role in lead and silver recovery.

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F22 and F27 were conducted at a k80 of 45 microns and the best lead and silver flotation results were achieved (top two curves in Figure 13-4).

F32 was performed at a k80 of 67 microns.  While the lead recovery was high at this size distribution, the silver recovery dropped to 62.1% from 68.9-71.8% compared to F22 and F27.

The lead and silver recovery dropped further as the size became coarser as shown in F6, F20, F8, F9, and F14, especially silver recovery which was significantly impacted.

Figure 13-4:  Primary Grind Size on Lead and Silver Flotation in Cleaner Tests

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13.4.1.2  Zinc Flotation

The effect of primary grind size on zinc rougher recovery is presented in Figure 13-5.  The zinc rougher recoveries vary between 58.9-62.7% and the overall zinc recoveries (zinc rougher recovery plus zinc in lead rougher concentrate) are 74.5-77%.  Higher overall zinc recoveries were obtained at finer primary grind sizes with the highest overall zinc recovery achieved from test F11, which was conducted at a k80 of 46 microns and resulted in an overall zinc recovery of 77%.

Figure 13-5:  Effect of Grind Size on Zinc Flotation in Rougher Tests

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13.4.2  Effect of Depressant Addition on Lead and Silver Flotation

Four lead rougher flotation tests were completed to investigate the effect of depressant addition on lead and silver flotation.  The results are presented in Figure 13-6 and Figure 13-7.

The final lead rougher recoveries were close regardless of the depressant added, but better lead and zinc separation was obtained from F28 in which ZnCN was used.  The zinc misplacement was the lowest at 20.4% in this test.  However, the ZnCN addition negatively impacted silver recovery as shown in Figure 13-6.  The silver recovery was much lower when compared to the other three tests, F29-F31.

Figure 13-6:  Effect of Depressant on Lead Rougher Flotation

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Figure 13-7:  Effect of Depressant on Silver Flotation

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The impact of ZnCN addition on lead and silver flotation was further confirmed by cleaner flotation as shown in Figure 13-8.

F35 and F37 were completed without any depressant addition, while ZnCN was added in the cleaner only (F21 and F36) or in rougher and cleaner (F39).  The silver recovery was significantly lower with the ZnCN addition (F12, F36 and F38) when compared to non-depressant addition tests (F35 and F37).

Figure 13-8:  Effect of ZnCN on Lead and Silver Flotation

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13.4.3  Effect of Longer Flotation and Higher Collector Addition on Pb/Ag Flotation

Finer primary grind of 45 microns produced the highest silver recovery as discussed earlier.  The lead recovery was high at a 76-micron grind but the silver recovery was lower when compared to a 45 micron grind.  F34 was completed at a 76-micron grind while one more lead rougher stage was added to extend the flotation time.  The collector dosage was also increased in rougher in that test.  The idea was to investigate if the longer flotation time and increased collector dosage could improve the silver recovery at 76 microns of grind.  The results are presented in Figure 13-9 and compared to F20 and F22.

The results clearly showed that the silver recovery improved significantly with the extended flotation time and higher collector addition level.

Figure 13-9:  Lead Cleaner Tests — Grade and Recovery Curves

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13.4.4  Effect of CuSO4 Dosage on Zinc Flotation and Pyrite Flotation

The CuSO4 addition level on zinc rougher flotation was investigated in F15, F16 and F17.  Pyrite flotation was also performed following the two stages of zinc rougher flotation by overdosing the collector addition.  The zinc flotation results are presented in Figure 13-10.

Higher CuSO4 addition level did improve the zinc rougher recovery.  Good zinc recovery was obtained at the lower CuSO4 addition level of 250 g/t.

Figure 13-10:  Effect of CuSO4 Dosage on Zinc Rougher Flotation

Pyrite flotation was conducted following zinc flotation by adding 250 g/t of SIPX with 8 minutes of flotation.  Only 1.1-2.2% of S was recovered into the pyrite concentrate.  Pyrite content in the master composite was low and most of the pyrite was already recovered into lead and zinc rougher concentrates during flotation.

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13.4.5  Effect of Na2SiO3 on Zinc Flotation

Sodium silicate was added in zinc flotation to investigate if it could improve the selectivity of zinc flotation and concentrate grade.  The results are compared to F7 and presented in Figure 13-11.  No improvement was observed on zinc flotation with the addition of sodium silicate.

Figure 13-11:  Effect of Sodium Silicate on Zinc Flotation

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13.4.6  Effect of pH on Zinc Flotation

The effect of pH on zinc rougher flotation was tested and the results are presented in Figure 13-12.  The three tests were completed under the same conditions except pH levels were 7.5, 9.0, and 11.0.  The overall zinc recovery (Zn rougher plus Zn in lead rougher) was 76.3%, 75.3% and 76.2% respectively for the three tests, F1, F12 and F13.

The pH had no significant effect on zinc rougher flotation, and this was further confirmed by cleaner tests as shown in Figure 13-13.  Good zinc flotation results were achieved at natural pH of 7-7.8.

Figure 13-12:  Effect of pH on Zinc Rougher Flotation

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Figure 13-13:  Effect of pH on Zinc Cleaner Flotation

13.4.7  Zinc Oxide Mineral Flotation

The Var 9 sample contained significant amount of zinc oxide minerals and the zinc recovery was relatively low by sulfide flotation.  Oxide flotation tests were conducted to investigate the potential of improving zinc flotation.  The following tests were completed:

·F18:  Na2S was added following zinc rougher flotation

·F19:  6493 was added following zinc rougher flotation

·F23-F26:  Na2S was added in primary grind mill

However, the oxide flotation was unsuccessful and there was no improvement of zinc recovery.

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13.4.8  Locked Cycle Flotation Testing

Five locked cycle tests were completed to confirm the batch cleaner test conditions and project the metallurgy.  The flotation conditions are summarized in Table 13-4.

LCT1, LCT2 and LCT3 were conducted to test the effect of primary grind size on flotation.  The grind sizes were conducted at k80’s of 46, 76 and 102 microns, respectively with all other conditions unchanged.  Three stages of lead cleaning and five staged of zinc cleaning were employed in those three tests.

LCT10 was conducted at a 45 microns grind size with 4 stages of lead and zinc cleaning and LCT11 was carried out at an 80 microns grind size with extended lead rougher flotation time and higher collector addition level.  The results are summarized in Table 13-5.

Table 13-4:  Summary Conditions of LCTs

Stage	Flotation Conditions	LCT-1	LCT-2	LCT-3	LCT-10	LCT-11
Primary Grind	Size, µm (K80)	46	76	102	45	80
	NaCN/ZnSO4, g/t	125/375	125/375	125/375	125/375	125/375
	pH	7.7	7.5	7.2	7.6	7.5
Pb Roughing	Stage	3	3	3	3	4
	3418A, g/t	30	30	30	30	80
	pH	7.6-7.7	7.5-7.6	7.2-7.4	7.6-7.7	7.5-7.6
	Float Time (min.)	7	7	7	7	11
Pb Regrind	Size, µm (K80)	17	17	19	17	19
	NaCN/ZnSO4, g/t	25/75	25/75	25/75	25/75	25/75
	pH	7	6.8	6.9	7	7.2
Pb Cleaning	Stage	3	3	3	4	4
	NaCN/ZnSO4, g/t	32.5/97.5	32.5/97.5	32.5/97.5	32.5/97.5	32.5/97.5
	3418A, g/t	12.5	12.5	12.5	13.5	15
	pH	6.8-7.1	6.7-6.9	6.7-6.9	6.7-7	6.9-7.3
	Float Time (min.)	11	11	11	14	14
Zn Roughing	Stage	2	2	2	2	2
	CuSO4	250	250	250	250	250
	SIPX	180	180	180	180	180
	pH	7-7.6	6.8-7.4	6.7-7.3	7-7.5	6.8-7.4
	Float Time (min.)	16	16	16	16	16
Zn Regrind	Size, µm (K80)	21	22	22	20	15
	CuSO4	100	100	100	100	100
	pH	6.9	6.8	6.8	7	6.7
Zn Cleaning	Stage	5	5	5	4	5
	SIPX	37	37	37	35	37
	pH	6.9-7.9	6.9-7.6	6.9-7.4	7.1-7.6	7-7.3
	Float Time (min.)	27	27	27	22	27

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Table 13-5:  Summary Results of Locked Cycle Test

Grade		LCT1,46 µm	LCT2,76 µm	LCT3,102µm	LCT10,45 µm	LCT11,80 µm
Feed (calc.)	Au, g/t	0.36	0.35	0.35	0.35	0.39
	Ag, g/t	281	276	291	281	289
	Cu, %	0.12	0.11	0.12	0.12	0.12
	Pb, %	2.35	2.39	2.43	2.42	2.52
	Zn, %	5.23	5.25	5.27	5.29	5.05
	Fe, %	3.55	3.44	3.40	3.60	3.27
	S, %	3.38	3.40	3.35	3.45	2.97
	F, %	3.64	3.56	3.60	3.22	3.23
Pb Cln Conc	Au, g/t	5.9	4.7	4.7	6.9	5.8
	Ag, g/t	5884	4816	4995	6054	5271
	Cu, %	1.62	1.35	1.33	1.64	1.57
	Pb, %	60.7	61.8	61.4	64.1	56.3
	Zn, %	8.98	8.60	8.73	8.01	11.2
	Fe, %	3.15	3.31	3.29	2.80	3.79
	S, %	15.7	16.0	15.9	14.8	16.6
	F, %	0.21	0.21	0.18	0.15	0.24
Zn Cln Conc	Au, g/t	0.52	0.70	0.69	0.65	0.8
	Ag, g/t	531	866	874	631	552
	Cu, %	0.59	0.69	0.69	0.58	0.56
	Pb, %	1.19	1.30	1.36	2.14	0.95
	Zn, %	56.9	55.2	54.7	52.5	51.8
	Fe, %	6.39	7.37	7.37	7.55	9.20
	S, %	33.0	31.7	30.0	32.9	30.3
	F, %	0.05	0.08	0.09	0.13	0.12

Distribution		LCT1,46 µm	LCT2,76 µm	LCT3,102µm	LCT10,45 µm	LCT11,80 µm
Pb Cln Conc	Au,%	60.6	50.3	49.9	62.2	60.2
	Ag, %	73.3	62.7	62.1	70.5	74.8
	Cu, %	48.3	41.4	39.4	48.4	53.3
	Pb, %	89.7	89.4	88.4	86.4	91.5
	Zn, %	5.98	5.73	5.73	4.89	9.05
	Fe, %	3.15	3.51	3.54	2.56	4.74
	S, %	16.7	17.7	18.4	14.2	22.9
	F, %	0.20	0.20	0.18	0.15	0.31

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Distribution		LCT1,46 µm	LCT2,76 µm	LCT3,102µm	LCT10,45 µm	LCT11,80 µm
Zn Cln Conc	Au,%	9.29	13.4	13.7	12.3	12.1
	Ag, %	11.6	20.4	20.3	15.5	11.6
	Cu, %	30.9	38.3	38.1	36.1	28.2
	Pb, %	3.08	3.40	3.66	6.10	2.28
	Zn, %	66.5	66.8	67.1	67.7	62.1
	Fe, %	11.2	14.2	14.8	14.5	17.0
	S, %	61.7	63.7	64.8	66.8	61.7
	F, %	0.09	0.14	0.15	0.27	0.22

Very good and consistent results were obtained.  The results further confirmed the findings in batch tests:

·Finer primary grinding improved gold and silver recovery in lead cleaner concentrate.  The gold and silver recoveries were 60.6% and 73.3% respectively in LCT1 which was carried out at 46 microns of grind size, while the gold and silver recoveries decreased to approximately 50% and 62% in LCT2 and LCT3 which were completed at 76 and 102 microns of grind size.

·The lead and silver grades improved with the addition of one more lead cleaner stage, but both the lead and silver recoveries decreased by approximately 3% when comparing LCT10 and LCT1.  The zinc grade decreased to 52.5% from 56.9% with the four stages of cleaning instead of five while the zinc recovery improved by 1% (67.7% in LCT10 vs. 66.5% in LCT1).  The four stages of zinc cleaning also resulted in higher fluorine grade in final zinc cleaner concentrate.  The fluorine grade increased to 0.129% F in LCT10 from 0.054% F in LCT1.

·Extended flotation time and higher collector dosage in the lead rougher improved lead, silver and gold recoveries even at a coarser primary grind size.  LCT11 was completed at 80 microns of grind with extended lead rougher flotation time and increased collector dosage.  The lead, silver and gold recoveries increased to 91.5%, 74.8% and 60.2% respectively, compared to 89.4%, 62.7% and 50.3% in LCT2 which was carried out at a grind of 76 microns.  However, the longer flotation time and increased collector addition in lead rougher resulted in lower final zinc recovery due to more zinc appearing in the lead cleaner concentrate.  The final zinc recovery in LCT11 was 62.1% compared to 66.8% in LCT2.  The zinc recovery could improve by optimizing conditions to bring down the zinc misplacement in lead cleaner concentrate, 9.1% in LCT11 vs. 5.7% in LCT2.

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13.4.9  Product Characterization

ICPSCAN and whole rock analysis were conducted on selected lead and zinc cleaner concentrates.  Hg, Cl and Se were also assayed.  The results are summarized in Table 13-6.

Table 13-6:  ICPSCAN and WRA Analysis

Analyte	Unit	LCT-1-5 PbCln3 Conc	LCT-1-5 ZnCln5 Conc	LCT-2-5 PbCln3 Conc	LCT-2-5 ZnCln5 Conc	LCT-3-5 PbCln3 Conc	LCT-3-5 ZnCln5 Conc
Al	%	0.14	0.04	0.15	0.06	0.14	0.06
Ba	ppm	220	40	210	50	180	60
Be	ppm	<5	<5	<5	<5	<5	<5
Ca	%	0.2	<0.1	0.1	<0.1	0.1	<0.1
Cr	ppm	40	50	40	80	70	90
Fe	%	3.22	6.26	3.06	7.34	3.22	7.66
K	%	<0.1	<0.1	<0.1	<0.1	<0.1	<0.1
Li	ppm	<10	<10	<10	<10	<10	<10
Mg	%	0.02	<0.01	0.02	<0.01	0.02	<0.01
Mn	ppm	180	1370	160	1340	170	1300
Ni	ppm	6	8	6	12	15	11
P	%	<0.01	<0.01	<0.01	<0.01	<0.01	<0.01
Sc	ppm	<5	<5	<5	<5	<5	<5
Sr	ppm	10	<10	10	<10	10	<10
Ti	%	0.01	<0.01	0.01	<0.01	0.01	<0.01
V	ppm	62	21	56	24	54	24
As	ppm	1680	1160	1500	1900	1660	2150
Bi	ppm	83.5	2.2	87.4	2.7	81.6	3.2
Cd	ppm	4520	5560	3530	5710	4130	5710
Ce	ppm	1.1	0.4	1.1	0.4	1.1	0.4
Co	ppm	17.3	97	15.6	96.5	17.3	98.1
Cs	ppm	0.7	0.4	0.7	0.6	0.6	0.7
Dy	ppm	0.17	0.08	0.15	0.08	0.14	0.08
Er	ppm	0.11	<0.05	0.07	<0.05	0.09	0.06
Eu	ppm	0.09	<0.05	0.09	<0.05	0.08	<0.05
Ga	ppm	3	10	3	9	3	10
Gd	ppm	0.14	0.07	0.12	0.05	0.14	0.06
Ge	ppm	5	1	4	1	4	1
Hf	ppm	<1	<1	<1	<1	<1	<1
Ho	ppm	<0.05	<0.05	<0.05	<0.05	<0.05	<0.05
In	ppm	<0.2	<0.2	<0.2	<0.2	<0.2	0.2
La	ppm	0.7	0.2	0.6	0.2	0.6	0.2
Lu	ppm	<0.05	<0.05	<0.05	<0.05	<0.05	<0.05
Mo	ppm	11	4	8	8	9	8
Nb	ppm	<1	<1	<1	<1	<1	<1
Nd	ppm	0.6	0.2	0.6	0.3	0.6	0.2
Pr	ppm	0.15	<0.05	0.15	0.06	0.13	0.06
Rb	ppm	3.1	0.8	3.1	1.1	2.6	1.2
Sb	ppm	1430	167	1540	172	1490	182
Sm	ppm	1	0.1	0.9	0.1	0.8	0.2
Sn	ppm	<1	<1	<1	<1	<1	<1
Ta	ppm	<0.5	<0.5	<0.5	<0.5	<0.5	<0.5
Tb	ppm	<0.05	<0.05	<0.05	<0.05	<0.05	<0.05
Th	ppm	0.2	<0.1	0.2	<0.1	0.2	0.1
Tl	ppm	1.4	<0.5	1.5	<0.5	1.5	<0.5
Tm	ppm	<0.05	<0.05	<0.05	<0.05	<0.05	<0.05
U	ppm	1.89	0.61	1.82	0.73	1.56	0.78
W	ppm	18	5	17	6	16	6
Y	ppm	0.9	<0.5	0.8	<0.5	0.7	<0.5
Yb	ppm	<0.1	<0.1	<0.1	<0.1	<0.1	<0.1
Zr	ppm	6.2	2.3	5.3	3.4	5.5	5

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Analyte	Unit	LCT-1-5 PbCln3 Conc	LCT-1-5 ZnCln5 Conc	LCT-2-5 PbCln3 Conc	LCT-2-5 ZnCln5 Conc	LCT-3-5 PbCln3 Conc	LCT-3-5 ZnCln5 Conc
Al2O3	%	0.27	0.08	0.3	0.1	0.26	0.11
Ba	%	0.021	0.003	0.02	0.004	0.016	0.005
CaO	%	0.35	0.12	0.33	0.13	0.29	0.14
Cr2O3	%	<0.01	<0.01	<0.01	0.01	<0.01	0.01
K2O	%	0.07	0.02	0.08	0.03	0.06	0.03
MgO	%	0.03	<0.01	0.04	0.01	0.03	0.01
MnO	%	0.02	0.16	0.02	0.16	0.02	0.16
Na2O	%	<0.01	0.01	0.02	0.01	0.03	0.02
Nb	%	<0.001	<0.001	<0.001	<0.001	<0.001	<0.001
P2O5	%	<0.01	0.03	<0.01	0.06	0.02	0.03
SiO2	%	5.19	1.29	4.74	1.53	4.46	1.68
Sr	%	<0.001	<0.001	<0.001	<0.001	<0.001	<0.001
TiO2	%	0.02	<0.01	0.02	<0.01	0.02	<0.01
Y	%	<0.001	<0.001	<0.001	<0.001	<0.001	<0.001
Zr	%	<0.001	<0.001	<0.001	<0.001	<0.001	<0.001
LOI	%	7.2	15.8	7.03	16.2	7.15	16
Hg	ppm	59.2	29.5	52.1	29.9	49.2	29.1
Se	ppm	316	65	311	67	305	67
Cl	ppm	<50	<50	<50	50	<50	<50

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13.5  Variability Composite Flotation

One cleaner flotation test was conducted on each of the 21 variability composites by employing the standard flotation conditions (LCT2 conditions).  The results are compared to master composite (LCT2) and plotted in Figure 13-14, Figure 13-15 and Figure 13-16.

Figure 13-14:  Silver Flotation Results of Variability Composites

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Figure 13-15:  Lead Flotation Results of Variability Composites

Figure 13-16:  Zn Flotation Results of Variability Composites

13.5.1  Silver Flotation

Silver recoveries in final lead cleaner concentrates ranged from 34.5 to 81.4% at 1280 to 24,017 g/t silver grade.  The average recovery of the 21 composites was 58.5% at a grade of 4,872 g/t silver, compared to 62.2% of silver recovery and 4,818 g/t of silver grade obtained from master composite (LCT2).  The variability tests were conducted in open circuit and the average silver recovery will be higher in locked cycle circuit.

Low silver recovery of 34.5% was obtained from Var 20.  The sample was from Southeast zone, which had a higher lead deportment percentage to Pb-Mn (Ba) Oxides.  Generally, the four composites (Var18-Var21) from Southeast zone had higher lead oxide minerals when compared to composites (Var14-Var17) from Northwest zone per QEMSCAN analyses.

13.5.2  Lead Flotation

Lead recoveries in final cleaner concentrates ranged from 34.5 to 96.7% at 44 to 80.8% of lead grade.  The average lead recovery of the 21 composites was 81.7% at a grade of 65.1% Pb, compared to 89.4% of lead recovery and 61.8% of lead grade obtained from master composite (LCT2).  The lead recovery will be higher in locked cycle circuit.

It should be noted that LCT1 data (lead recovery of 89.7% in lead concentrate and lead recovery of 3.08% in zinc concentrate were used in the economic analysis.

Lead recoveries of the composites (Var18-Var21) from Southeast zone were also impacted by occurrence of the lead oxide minerals.

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13.5.3  Zinc Flotation

Zinc recoveries in final cleaner concentrates ranged from 36.6 to 79.9% at 38.3 to 60.2% of zinc grade.  The average zinc recovery of the 21 composites was 61.9% at a grade of 55.2% Zn, compared to 66.8% of zinc recovery and 55.2% of zinc grade obtained from master composite (LCT2).  The zinc final recovery will be higher in locked cycle circuit.

It should be noted that LCT1 data (zinc recovery of 5.98% in lead concentrate and zinc recovery of 66.5% in zinc concentrate were used in the economic analysis.

Lower zinc recoveries were also obtained from two Southwest composites (Var20 and Var21), which had a high willemite content of 51.6% and 35.5% respectively.  Zinc misplacement of Var 6 and Var 17 in lead circuit was high and optimization is required to improve zinc recovery.  There was a good correlation between zinc recovery and sphalerite content.

13.6  Cyanide Destruction

The slurries used for the cyanide destruction (CND) testwork were the zinc tailings from LCT12 and LCT13 (zinc first cleaner tail and zinc rougher tail combined).  The target primary grind size of the two tests was 74 and 46 microns respectively.  Cyanide was added during flotation and the cyanide concentration (CNWAD) was 10.4 mg/L before destruction.  The objective of the testwork was to investigate the amenability of the samples to detoxification using SO2/air and to produce a treated product containing <1 mg/L residual CNWAD for filtration testwork.

One batch test and one continuous test were completed on each of the two tailings pulp generated from LCT12 and LCT13.  The test conditions and results are summarized in Table 13-7 and Table 13-8.  All reagent consumption was calculated on pure basis.

13.7  Fluorine Control

During operations, it has become evident that additional removal of fluorine is necessary.  The key reason for the high fluorine content in the lead and zinc concentrates is the high fluorite (CaF2) content in the orebody, head grade ranges in two categories from 25,000 to 40,000 F ppm to 40,000 to 60,000 F ppm.  The fluorine content in the fluorite is 48.7% by weight.  In general, the higher the fluorite in the plant feed, the higher the fluorine content in the concentrates.  The fluorine content at the mine is approximately 40,000 ppm, this means to achieve the target fluorine content of less than 1,500 ppm in the lead concentrate and less than 500 ppm in the zinc concentrate, a fluorite rejection of 99.91 and 99.95 is needed, which is a challenging task.  Due to the challenging nature, there is a need to study the liberation status of the fluorine carrier minerals on a size-by-size basis to know if there are liberation restrictions to achieve the required targets.

The Cerro Los Gatos Mine has reduced the fluorite content of the concentrate by providing additional cleaning stages, four and six stages in the lead and zinc floatation sections (from three and five).  This is done to minimize the recovery of the fluorite by entrainment/achieve the targeted fluorite rejection.  The problem is exacerbated in the lead concentrate due to the smaller particle size P80 10 µm vs 20 µm in the zinc concentrate.  The finer the particle size the higher the unselective mineral recovery by entrainment.  At this stage, the processing plant metallurgy team has focused on implementing best practice tools to minimize the entrainment, such as:  running a stable operation, increasing the number of cleaning stages,

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adding dilution water and froth washing in the cleaning circuit, good air control, deep frothers (60 cm), and having a less tenacious frother by using a MIBC (methyl isobutyl carbinol) based frother.  There are further opportunities to reduce fluorine by reducing the process variability.

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Table 13-7:  Summary Conditions of CND Test

		Pulp	Reactor	Test	Retention		g/g CN WAD			g/L Feed Pulp
		Density	Vol	Duration	Time		SO2			SO2
Test No\Feed	Process	%	L	Min	Min	pH	Equiv.	Lime	Cu	Equiv.	Lime	Cu
CND-1/LCT12	SO2/Air	34	6	385	19	8.5	17.67	11.59	0.70	0.15	0.10	0.0059
CND-2/LCT13	SO2/Air	23	6	280	14	8.5	17.10	12.03	0.81	0.15	0.11	0.0073

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Table 13-8:  Summary Results of CND Test

Analyte	Unit	CND1Feed	CND2Feed	CND1Final	CND2Final
Total NaCN	mg/L	10.6	10.4	0.02	0.02
Free NaCN	mg/L	<3.5	<3.5	<0.01	<0.01
Cyanide WAD	mg/L	10.4	10.4	<0.01	<0.01
Thiocyanate	mg/L	4	4.5	4.1	4.3
Ag	mg/L	<0.08	<0.08	<0.08	<0.08
Al	mg/L	<0.2	<0.2	<0.2	<0.2
As	mg/L	<3	<3	<3	<3
Ba	mg/L	0.054	0.065	0.053	0.068
Be	mg/L	<0.002	<0.002	<0.002	<0.002
Bi	mg/L	<1	<1	<1	<1
Ca	mg/L	38.3	34.6	122	138
Cd	mg/L	<0.09	<0.09	<0.09	<0.09
Co	mg/L	<0.3	<0.3	<0.3	<0.3
Cr	mg/L	<0.1	<0.1	<0.1	<0.1
Cu	mg/L	14.3	13.9	0.54	0.74
Fe	mg/L	0.60	0.70	0.60	0.80
K	mg/L	17	19	17	22
Li	mg/L	<2	<2	<2	<2
Mg	mg/L	1.8	3.09	4.38	5.51
Mn	mg/L	0.75	0.99	0.68	0.62
Mo	mg/L	<0.6	<0.6	<0.6	<0.6
Na	mg/L	834	588	227	186
Ni	mg/L	<0.6	<0.6	<0.6	<0.6
P	mg/L	10	11	10	10
Pb	mg/L	<2	<2	<2	<2
Sb	mg/L	<1	<1	<1	<1
Se	mg/L	<3	<3	<3	<3
Sn	mg/L	<2	<2	<2	<2
Sr	mg/L	0.60	0.55	1.02	1.05
Ti	mg/L	<0.02	<0.02	<0.02	<0.02
Tl	mg/L	<3	<3	<3	<3
V	mg/L	<0.2	<0.2	<0.2	<0.2
Y	mg/L	<0.02	<0.02	<0.02	<0.02
Zn	mg/L	<0.7	1	<0.7	<0.7

The cyanide destruction tests turned out to be very successful.  The cyanide WAD concentration was reduced to 0.2 mg/L from 10.4 mg/L after 14-19 minutes of retention time during the continuous testing.

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14.0  MINERAL RESOURCE ESTIMATES

This Resource estimate is an update and includes drilling that has been completed since the 2017 feasibility study.  This includes infill drilling from both surface and from underground to define mineralization and to upgrade the mineral classification where possible.  Drilling has been performed in the Central Zone, but most of the program focused on the North West zone and its possible extension to the north (Drilling 2019).  Mineral Resources have been estimated for the epi-thermal veins of the Cerro Los Gatos deposit by multi-pass Ordinary Kriging (OK) of capped and composited drill hole samples.

Resources were estimated for the Amapola and Esther zones in August 2012.  No additional information has been produced for these two areas since the 2012 estimate, and therefore, the Resources have been brought forward and included in this report for completeness of the Resource of the project.

14.1  Cerro Los Gatos Deposit

Estimated Measured, Indicated and Inferred Mineral Resources for the Cerro Los Gatos deposit, effective September 6, 2019, are shown in Table 14-1 at a 150 AgEq g/t cutoff grade.

Table 14-1:  Mineral Resource Estimate

Classification	Tonnes	AgEq  g/t	Ag  g/t	Pb %	Zn  %	Au  g/t	Cu  %	AgEq toz M	Ag toz M	Pb lbs M	Zn lbs M	Au toz K	Cu lbs M
Measured	5,774,314	652	324	2.9	5.8	0.39	0.11	121	60	375	744	72	13
Indicated	4,586,507	489	202	2.5	5.2	0.28	0.11	72	30	251	528	42	12
Measured & Indicated	10,360,822	576	269	2.7	5.5	0.34	0.11	193	90	626	1,272	114	25
Inferred	3,717,063	361	107	2.8	4.0	0.28	0.14	43	13	231	330	34	12

NOTES:

1) 150 AgEq g/t cutoff grade has been calculated using $18/toz Ag, $0.92/lbs Pb, and $1.01/lbs Zn,

2) Columns may not total due to rounding,

3) Mineral Resources are stated as undiluted, inclusive of Mineral Reserves.

4) One troy ounce (toz) is equal to 31.1035 grams (g) and one Tonne is equal to 2,204.62 lbs.

Mineral Resources were estimated from 2,356 samples intersecting modeled veins, sourced from 426 diamond drill holes.  Capping was analyzed for each metal estimated using histograms and probability plots to determine where high-grade distribution tails deviated from lognormal.  Sampled intervals were composited to 2 m.  Composite intervals initiated a terminated at the top and bottom of the vein contacts.

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Vein model solids were constructed in MicroMine™ modeling software; the resulting solids are shown in Figure 14-1, below.  Grade-shells were used to further isolate +150 AgEq g/t grade population for estimation.

Figure 14-1:  Vein Solids 3D View

Blocks and composites from each vein and post mineral fault block domain were independently transformed, realigned and made relative to the footwall and hanging-wall for estimation.  Realignment allowed for estimation to occur across post-mineral fault blocks approximating pre-fault orientation of the veins.  Estimations relative to footwall and hanging-wall position allowed for better data honoring across the dip of the vein.

Only composites within the same vein were permitted to estimate blocks of a given vein domain; because of the transformation and realignment estimation was permitted across post mineral fault block areas with the same vein code.  Resulting block AgEq g/t grades are shown in Figure 14-2, and the Mineral Resource classifications are shown in Figure 14-3.

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Figure 14-2:  Estimated AgEq Block Grades within Grade-Shell 3D View

Figure 14-3:  Block Resource Classification 3D View

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14.1.1  Input Data

Drill hole data was provided by MPR in Microsoft Access format finalized as of August 5, 2019.  Table 14-2 shows the drill hole and sample counts for the project and the samples used to support the vein model.  Figure 14-4 shows the cross-section lines in red and drill holes in green.

Table 14-2:  Drill Hole and Sample Count

	Drill Hole Count	Sample Count
Cerro Los Gatos Project Area	484	15,486
Resource Area	426	6,483
Vein Domain (100-20000)	368	2,356
High-Grade Domain	296	1,896

Figure 14-4:  Drill Hole Plan and Cross-Section Index

14.1.2  Grade Capping

Grade populations for Ag, Pb, Zn, Au, and Cu within the vein interpretation and the +150 AgEq g/t grade-shell were analyzed using histograms and probability plots to determine where high-grade distribution tails became unsupported or deviate from lognormal.  Upper limits were applied to intervals before compositing.  Table 14-3 details the upper thresholds chosen as well as the number of samples capped and the mean before and after capping.  Figure 14-5 to Figure 14-9 are histograms showing the uncapped grade populations as well as the upper limits chosen.

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Table 14-3:  Grade Capping Statistics

Element	Unit	Uncapped Mean	Cap	Number Capped	Capped Mean
Ag	g/t	304	2,700	36	283
Pb	%	2.6	15	38	2.5
Zn	%	5.4	23	34	5.2
Au	g/t	0.37	4	13	0.34
Cu	%	0.11	0.7	11	0.10

Figure 14-5:  Histogram for Capping Analysis Ag

Figure 14-6:  Histogram for Capping Analysis Pb

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Figure 14-7:  Histogram for Capping Analysis Zn

Figure 14-8:  Histogram for Capping Analysis Au

Figure 14-9:  Histogram for Capping Analysis Cu

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14.1.3  Compositing

Sample intervals were composited to 2 m, which is the mode sample length.  Compositing initiated and terminated at the top and bottom of continuous selected vein samples, resulting composites were permitted to be 1-2 m in length, intervals less than 1 m were rejected.  As part of the Kriging process, composite influence was additionally weighted by interval length to provide further normalization.  Compositing greater than 2 m is not appropriate because 3 m composites would cause samples to be split and 4 m composites are too large to represent the vein across dip in some areas.

14.1.4  Geologic Modeling

A wireframe solid geologic model was generated for each of the fault block zones.  Data was integrated from: surface mapping completed by MPR, 2D cross-section completed by MPR, and the downhole geologic logging, to create the model.

The model was divided into four geological domains:

·Epiclastic erosional volcanic sediments above the rhyolite and andesite;

·Rhyolite intruding the epiclastic and andesite from the southeast;

·Undifferentiated andesitic volcanic, the primary vein host, situated in the footwall of the Los Gatos fault; and,

·Dacite/volcanic tuff within the andesite, which comprises the immediate footwall of the mineralization and occasional host.

Figure 14-10 shows the geologic model in 3d looking southwest from above.  Figure 14-11 is a cross-section of the solids in the Central zone showing the relationship of the veins as well as the downhole logging.

Figure 14-10:  Geologic Model Solids 3D View

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Figure 14-11:  Geologic Model Cross-Section 27 Looking NW

14.1.5  Vein Modeling

Drill hole intervals were sub-divided into three veins and five fault blocks.  Each interval interpreted to be with the vein model was coded with a vein and fault block code.  Fault blocks include from the northwest to the southeast include: Northwest (NW or 1), Central (CN or 2), Southeast (SE or 3), Southeast 2 (S2 or 4), and Southeast 3 (S3 or 5).  The use of numeric codes for fault blocks is exclusively for the block model export necessitated by software datatype restrictions.

Vein horizons include 100, 1000, 10000, 200, 2000, 20000 and 300, vein 300 is only recognized and used in the Central fault block and veins greater than or equal to 1000 are only used in the NW fault block.  Vein intervals were selected in cross-section as well as 3D and reviewed in 3D, level-plan, inclined level-plan, and vein plane space.

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Figure 14-12:  Vein Solids 3D View

Following vein assignments, intervals belonging to each vein and fault block were composited across the entire coded vein thickness.  For instance, in the Central block, three single interval composites were generated, one for 100, 200, and 300.

The single thickness intervals from each vein and fault block were included into two domains, low-grade and high-grade.  The high-grade domain represents the area or grade-shell of +150 AgEq g/t.  The grade-shell boundary was constructed at half the distance between single-interval composites less than 150 AgEq g/t and greater than or equal to 150 AgEq g/t.

Following the construction of the boundary, the initial vein solids were clipped by the boundary.  Extrapolation is most significantly limited by this process as shown below.  The boundaries were constructed and updated during each drill campaign.

Figure 14-13 to Figure 14-20 are 3D longitudinal views showing vein assignments, block assignments, single interval composite grades (AgEq, Ag, Pb, Zn, Au, Cu), and interval apparent thickness.

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Figure 14-13:  Long-Section Vein Intervals Assignments

Figure 14-14:  Long-Section AgEq Grade of Vein Intervals

Figure 14-15:  Long-Section Ag Grade of Vein Intervals within Grade-Shell

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Figure 14-16:  Long-Section Pb Grade of Vein Intervals within Grade-Shell

Figure 14-17:  Long-Section Zn Grade of Vein Intervals within Grade-Shell

Figure 14-18:  Long-Section Au Grade of Vein Intervals within Grade-Shell

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Figure 14-19:  Long-Section Cu Grade of Vein Intervals within Grade-Shell

Figure 14-20:  Long-Section Thickness of Vein Intervals within Grade-Shell

Following the construction of vein and grade-shell solids, dilution halo solids were constructed for each fault block.  Dilution solids were constructed to encompass the entire vein system and initiated and terminated along dip were any grade of AgEq greater than 0 was observed.

To correct for post mineral faulting as well as down-drop faulting in the NW fault block, each vein was first tilted and then rotated to a flat best fit vein space and then realigned and rotated to the interpreted orientation prior to faulting.

As discussed above, single interval composites from each vein and block were rotated and tilted around a fixed point to flattened best fit vein space.  Once in flat vein space single interval composites from each vein and block were rotated and shifted into estimation space Table 14-4 details the transformation for each vein and block.

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Table 14-4:  Vein Transformations

Vein	Block	Vein Strike	Vein Dip	Vein Space Rotation	Vein Space Tilt	Estimate Space Rotation Point X	Estimate Space Rotation Point Y	Estimate Space Rotation	Estimate Space Shift X	Estimate Space Shift Y	Estimate Space Separation Y
100	NW	310	75	-40	-75	368,426	3,048,249	-6	0	0	0
200	NW	310	75	-40	-75			-6	0	0	1100
1000	NW	310	35	-40	-35			3	-10	-50	0
2000	NW	310	35	-40	-35			1	-95	-60	1100
10000	NW	310	35	-40	-35			0	15	70	0
20000	NW	310	35	-40	-35			-4.5	25	125	1100
100	CN	300	40	-30	-40	368,417	3,048,211	10	40	6	0
200	CN	300	40	-30	-40			10	40	6	1100
300	CN	300	40	-30	-40			10	40	6	2200
100	SE	300	40	-30	-40	369,074	3,047,955	7	38	47	0
200	SE	300	40	-30	-40			7	66	47	1100
100	S2	300	40	-30	-40	369,162	3,047,952	7	30	-62	0
200	S2	300	40	-30	-40			7	60	-62	1100
100	S3	305	75	-35	-75	369,275	3,048,064	7	110	-87	0
200	S3	314	77	-44	-77	369,332	3,047,168	8	160	-147	1100

In the NW fault block veins in the hanging-wall are interpreted to be stacked as a result of additional down-drop faulting oriented in the same direction as the vein strike.  Conceptually, the NW fault block faulted down along strike in two events of detachment, where the up-dip portions of the vein continued to drop down when the down-dip portions of the vein had ceased to drop down.  The 10000 vein is considered the down-dip portion of the 1000 vein but is above the 1000 vein in real space.  Vein 1000 is considered the down-dip portion of the 100 vein but it is above the 100 vein in real space.

Veins 100, 1000, and 10000 have been modeled as a continuous vein horizon.  The reorganization of the NW veins to account for the faulting as well as alignment of all five fault blocks along strike.

The location of composites and blocks across dip, or z in estimation space, were also made relative to the hanging-wall and footwall of each vein.  This allowed for tracking of the vein shape for estimation.  Other techniques, such as fixed search ellipse orientation or block by block orientation, fell short where hanging-wall and footwall relative z coordinates were able to best capture hanging-wall and footwall grade biases that can be frequently observed.

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14.1.5.1  Specific Gravity Determination

Specific gravity (SG) measurements in g/cm3 were made on core by MPR geologist using epoxy coating and the water immersion technique.  Measurements have been divided into three categories; inside the vein interpretations, outside of interpretation but within the dilution halo, and outside of the grade estimation area.  The grade of Pb plus Zn and the Measured SG were used to develop regressions for inside the vein interpretation and outside but within the dilution halo.  The regressions were manually fit to observations and compared to the theoretical SG at the deposit’s average grade assuming constituent minerals quartz, galena, and sphalerite.  SG values for rock units outside of the estimation area were derived from the average SG of measurements within the modeled lithologic domains and outside of the dilution halo and vein domains. Table 14-5 details the measurements and regressions within the various areas.

Table 14-5:  Specific Gravity Measurements and Regressions

Area	Outlier Definition	SG Average w/o Outliers	Regression Equation
Within Vein	SG>3.9 and Pb+Zn>300,000 or SG>4.2	2.86	2.63+(Pb+Zn)*0.0000023
Outside Vein and Within Dilution Halo	SG>2.9 or Pb+Zn>20,000	2.58	2.55+(Pb+Zn)*0.0000043
Andesite	SG>2.83	2.52	NA
Dacite	SG>2.67	2.53
Rhyolite	NA	2.44
Epiclastic	SG>2.5	1.91

Figure 14-21 shows the regression equation compared to the measurements within the interpreted veins.

Figure 14-21:  SG Regression within Vein Interpretation

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14.1.6  Estimation Methods and Parameters

Resources have been estimated for five fault block areas and three vein horizons using hanging-wall and footwall relative multi-pass ordinary kriging of 2 m composites.

14.1.6.1  Variography and Search

Log-normal experimental variograms were generated for Ag, Pb, Zn, Au, and Cu in vein-relative estimation space for 2 m composites within the +150 AgEq g/t grade-shell.  Primary plunge of the central zone has an azimuth of 300° and a dip of 10° as measured in true space, which equate to 100° in transformed estimation space.

Table 14-6 and Table 14-7 detail the components of the modeled variograms used for estimation  In Figure 14-22 to Figure 14-26 the primary direction is shown in red, secondary in green, and tertiary in blue.

Table 14-6:  Modeled Variograms for Ag

Direction	Orientation	Azimuth (Estimation Space)	Geologic Basis	Nugget	Nugget % of Total Sill	C1 Partial	C1 Range m	C2 Partial Sill	C2 Range m	Total Sill
Along Plunge	Primary	100	Observed plunge, 10° from strike	0.5	36	0.6	40	0.3	100	1.4
Perpendicular to Plunge	Secondary	190	Perpendicular to plunge, 10° from dip	0.5	36	0.6	15	0.3	50	1.4
Across-vein	Tertiary	190	Aligned to vein thickness	0.5	36	0.6	5	0.3	10	1.4

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Table 14-7:  Modeled Variograms for Pb, Zn, Au, Cu

Element	Orientation	Azimuth (Estimation Space)	Nugget	Nugget % of Total Sill	C1 Partial	C1 Range m	C2 Partial Sill	C2 Range m	Total Sill
Pb	Primary	90	0.55	42	0.65	30	0.1	75	1.3
Pb	Secondary	180	0.55	42	0.65	15	0.1	40	1.3
Pb	Tertiary	180	0.55	42	0.65	7	0.1	10	1.3
Zn	Primary	100	0.4	38	0.2	30	0.45	85	1.05
Zn	Secondary	190	0.4	38	0.2	20	0.45	60	1.05
Zn	Tertiary	190	0.4	38	0.2	4	0.45	8	1.05
Au	Primary	100	0.4	44	0.5	55	—	—	0.9
Au	Secondary	190	0.4	44	0.5	40	—	—	0.9
Au	Tertiary	190	0.4	44	0.5	8	—	—	0.9
Cu	Primary	100	0.4	49	0.2	40	0.22	90	0.82
Cu	Secondary	190	0.4	49	0.2	25	0.22	40	0.82
Cu	Tertiary	190	0.4	49	0.2	4	0.22	10	0.82

Figure 14-22:  Experimental and Modeled Variography Ag

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Figure 14-23:  Experimental and Modeled Variography Pb

Figure 14-24:  Experimental and Modeled Variography Zn

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Figure 14-25:  Experimental and Modeled Variography Au

Figure 14-26:  Experimental and Modeled Variography Cu

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A sub-blocked model was fit to the extents of the dilution halo and sub-blocked to the modeled vein solids with the parameters shown in Table 14-8.

Table 14-8:  Block Model Setup Parameters

Direction	Origin (Corner)	Parent Block Size m	Parent Blocks	Length m	Block Divisions	Smallest Child Block Size m	Rotation About Axis (Clockwise)
X	367,000	10	274	3,200	2	5	0
Y	3,047,700	5	144	905	2	2.5	0
Z	900	5	150	750	2	2.5	30

Blocks and 2 m composites were transformed to estimation space as described above.  Table 14-9 details the search ellipse sizes, and orientations along with sample selection criteria for each pass.  The search ellipse was rotated to align with the principal plunge of mineralization, but tilting was not required because the composites and blocks were transformed and made relative to the hanging wall and footwall before estimation.

Table 14-9:  Pass Parameters

Pass	Method	Max Search m	Primary Search Azi	Ratio to Max 1st:2nd:3rd	Sectors	Comp Per Sector Max	DH Min	DH Max	Comp per DH Max	Comp Min	Comp Max
1	Ordinary Kriging	40	100	1:0.5:0.25	4	4	3	8	2	3	16
2	Ordinary Kriging	70	100	1:0.7:0.25	1	12	2	6	2	1	12
3	Ordinary Kriging	90	100	1:0.7:0.25	1	12	1	6	2	1	12
4	Ordinary Kriging	125	100	1:0.7:0.25	1	8	1	4	2	1	8
5	Ordinary Kriging	190	100	1:0.7:0.25	1	6	1	3	2	1	6

Figure 14-27 shows the resulting estimated block AgEq g/t grades in 3D view.  Figure 14-28 to Figure 14-32 show AgEq, Ag, Pb, Zn, Au, and Cu in 3D view.

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Figure 14-27:  Estimated AgEq Block Grades within Grade-Shell 3D View

Figure 14-28:  Estimated Ag Block Grades within Grade-Shell 3D View

Figure 14-29:  Estimated Pb Block Grades within Grade-Shell 3D View

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Figure 14-30:  Estimated Zn Block Grades within Grade-Shell 3D View

Figure 14-31:  Estimated Au Block Grades within Grade-Shell 3D View

Figure 14-32:  Estimated Cu Block Grades within Grade-Shell 3D View

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14.1.7  Mineral Resource Classification

Resource classification was assessed primarily by 3D drill hole (sample) spacing followed by manual evaluation and polishing using regions but also considered:  pass (including maximum search, sectors, drill hole and sample requirements), and nearest sample.

Block classification criteria related to the estimation pass are shown in Table 14-10 below, block classification is shown in 3D in Figure 14-33, looking northwest from above.

Table 14-10:  Mineral Resource Classification

Classification	Pass	Search Max (m)	Drill Hole Min	Drill Hole Spacing
Measured	1, 2	40, 70	3, 2	<25
Indicated	1, 2, 3	40, 70, 90	2, 1	<50
Inferred	2, 3, 4, 5	70, 90, 125, 190	1	All

Figure 14-33:  Mineral Resource Classification 3D View

Figure 14-34 is a stacked histogram of the resulting block classifications and their nearest composite sample, demonstrating that blocks classified as measured are on average about 15 m from the nearest composite and the majority are less than 25 m.  Blocks classified as Indicated are on average about 25 m from the nearest composite and the majority are less than 50 m.

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Figure 14-34:  Stacked Histogram of Nearest Composite Sample to Classified Blocks

14.1.8  Dilution

No dilution has been accounted for in Mineral Resource estimation or in the statement of Mineral Resources.

14.1.9  Cutoff Grade and Reasonable Prospects for Economic Extraction

Cutoff grade has been estimated using generalized parameters prior to detailed analysis by mining study (see Table 14-11).  The base case cutoff grade of 150 AgEq g/t accounts for typical costs to mine and process the Measured and Indicated Mineral Resources for 8-10 years and is adequate to approximate reasonable prospects for economic extraction.  AgEq is calculated for each block in the model using $18/toz Ag, $0.92/lb Pb, and $1.01/lb Zn.

Table 14-11:  Cutoff Grade Parameters

	Value	Unit
Mining Cost	40	$/Tonne
Processing Cost	20	$/Tonne
G&A	4	$/Tonne
Ag Recovery	83	%
Ag Price	18	$/toz
Pb Recovery	92	%
Pb Price	0.92	$/lbs
Zn Recovery	77	%
Zn Price	1.01	$/lbs
Cutoff Grade	150	AgEq g/T

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Gold and copper, being part of the epithermal system, are included in the Mineral Resource estimate as comparisons, however they have not been considered for purposes of determining the Ag/Eq cut-off grade since copper in the final concentrates sold would not receive any payment and gold revenue is estimated at only 2.78% of the payable metal.  Preparing a mine plan that excluded gold from the AgEq calculations had little or no bearing on estimates of ore that will be mined.  The cash flow estimates include the estimated gold revenue from the mine plan.  The incremental ore that could be processed from including the approximately 0.8 ounces AgEq from the calculation is immaterial and conservative.

Subsequent studies could materially alter any or all the parameters used to approximate a reasonably informed cutoff grade.

14.1.10  Statement of Mineral Resources

Estimated Measured, Indicated, and Inferred Mineral Resources, effective September 6, 2019, for the Cerro Los Gatos deposit at a 150 AgEq g/t cutoff grade are shown in Table 14-12, and Mineral Resources by fault block are shown in Table 14-13.  Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Table 14-12:  Mineral Resource Estimate

Classification	Tonnes	AgEq g/t	Ag g/t	Pb %	Zn %	Au g/t	Cu %	AgEq  toz M	Ag toz M	Pb lbs M	Zn lbs M	Au toz K	Cu lbs M
Measured	5,774,314	652	324	2.9	5.8	0.39	0.11	121	60	375	744	72	13
Indicated	4,586,507	489	202	2.5	5.2	0.28	0.11	72	30	251	528	42	12
Measured and Indicated	10,360,822	576	269	2.7	5.5	0.34	0.11	193	90	626	1,272	114	25
Inferred	3,717,063	361	107	2.8	4.0	0.28	0.14	43	13	231	330	34	12

NOTES:

1) 150 AgEq g/t cutoff grade has been calculated using $18/toz Ag, $0.92/lbs Pb, and $1.01/lbs Zn,

2) Columns may not total due to rounding,

3) Mineral Resources are stated as undiluted, and are inclusive of Mineral Reserves.

4) One troy ounce (toz) is equal to 31.1035 grams (g) and one Tonne is equal to 2,204.62 lbs. 3,717,063

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Table 14-13:  Mineral Resource Estimate by Fault Block

Fault Block	Classification	Tonnes M	AgEq g/t	Ag g/t	Pb %	Zn %	Au g/t	Cu %
NW	Measured	2.4	710	345	3.4	6.4	0.47	0.11
NW	Indicated	0.9	528	223	2.5	5.6	0.27	0.10
NW	Measured and Indicated	3.3	660	312	3.2	6.2	0.42	0.11
NW	Inferred	1.6	299	95	2.1	3.4	0.25	0.13
CN	Measured	3.4	613	311	2.6	5.5	0.33	0.10
CN	Indicated	2.2	525	251	2.1	5.2	0.37	0.11
CN	Measured and Indicated	5.6	578	287	2.4	5.4	0.35	0.10
CN	Inferred	0.2	351	199	1.3	2.8	0.36	0.05
SE	Indicated	0.2	383	113	2.3	5.0	0.18	0.11
SE	Inferred	—	—	—	—	—	—	—
S2	Indicated	0.6	534	136	3.7	7.0	0.15	0.17
S2	Inferred	0.1	377	109	3.2	4.0	0.16	0.21
S3	Indicated	0.7	328	102	2.7	3.4	0.16	0.10
S3	Inferred	1.7	402	106	3.5	4.5	0.30	0.15

NOTES:

1) 150 AgEq g/t cutoff grade has been calculated using $18/toz Ag, $0.92/lb Pb, and $1.01/lb Zn,
2) Columns may not total due to rounding, and
3) Mineral Resources are stated as undiluted and are inclusive of Mineral Reserves,

Grade tonnage curves for Measured and Indicated Mineral Resources are shown in Figure 14-35, and Inferred Mineral Resources are shown in Figure 14-36.

Figure 14-35:  Grade Tonnage Curve Measured and Indicated Resources

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Figure 14-36:  Grade Tonnage Curve Inferred Resources

14.1.11  Model Verification

Mineral Resource estimations have been verified by visual review, internal peer review, population analysis, alternative estimate methodology, and progressive testing by way of interim modeling throughout the 2018-2019 Resource definition drilling campaign.

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Figure 14-37 to Figure 14-40 compare drill holes and resulting blocks in cross-section for AgEq, Ag, Pb, and Zn.

Figure 14-37:  Cross-Section 13 AgEq

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Figure 14-38:  Cross-Section 13 Ag

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Figure 14-39:  Cross-Section 13 Pb

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Figure 14-40:  Cross-Section 13 Zn

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Alternative estimates were used to validate the OK estimation method ultimately selected for Mineral Resource reporting.  Using the same search and selection parameters, blocks were estimated for Ag, Pb, and Zn using inverse distance to the power of two (ID2), multiple indicator Kriging (MIK), and nearest neighbor (NN).  Table 14-14 shows the tonnage and grade percent differences compared to the OK estimate at a 150 AgEq g/t cutoff grade for Measured and Indicated blocks.  Difference percent was calculated by dividing the alternative estimate by the OK estimate, multiplying by 100 and subtracting 100.  The results show the tonnage estimated by ID2 and MIK are very similar to OK.  MIK produced differences in Pb and Zn grade, but the differences are within a reasonable range and do not affect the estimated tonnage above cutoff.  Differences between OK and NN are as expected and indicate 7% less AgEq contained metal is estimated by NN than OK.

Table 14-14:  Alternative Estimate Difference Percent, Measured and Indicated Blocks

Method	Tonnes	AgEq g/t	Ag g/t	Pb %	Zn %	AgEq toz	Ag toz	Pb lbs	Zn lbs
OK	0	0	0	0	0	0	0	0	0
ID2	-4	-1	2	-3	-2	-4	-3	-6	-6
NN	-15	14	16	10	9	-7	-2	-10	-11
MIK	-3	-7	-1	-11	-12	-9	-4	-13	-14

Figure 14-41 to Figure 14-45 compare the tonnage and grade of OK to the alternative estimates at a range of cutoff grades for Measured and Indicated Mineral Resources.

Figure 14-41:  Alternative Estimate Comparison Tonnage Curve

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Figure 14-42:  Alternative Estimate Comparison AgEq Grade Curve

Figure 14-43:  Alternative Estimate Comparison Ag Grade Curve

Figure 14-44:  Alternative Estimate Comparison Pb Grade Curve

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Figure 14-45:  Alternative Estimate Comparison Zn Grade Curve

In addition to alternative estimations, the OK estimate was compared to the capped Ag, Pb, and Zn assay intervals and volume within the +150 AgEq g/t grade-shell.  Tonnage was derived from wireframe volume multiplied by the density measurements within the +150 AgEq g/t grade-shell.

14.1.12  Relevant Factors

Estimation domains have been constructed using AgEq grades that include Ag, Pb, and Zn.  Significant changes in any or all the three metal values could materially affect the boundary of a +150 AgEq g/t grade-shell across vein thickness and in the plane of the vein.  The grade tonnage curves of Measured and Indicated Mineral Resources shown in Figure 14-35 shows that blocks within the grade-shell are relatively insensitive to a cutoff increases up to 200 AgEq g/t; however, sensitivity analysis has not been performed with grade-shells constructed at various cutoffs.

Recent startup of mining and is showing that the 150 ppm cutoff grade is a reasonable cutoff.  Recent startup of production should provide additional data to refine the cutoff grade parameters in the future, as needed.

There are no additional environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that the author of this report is aware of that could materially affect the Mineral Resource estimate.

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14.2  Amapola Zone

The Amapola Zone was included in the Resource report produced by Tetra Tech in August 2012.  No additional exploration has been completed in this area.  For completeness of the Resource estimate in this update, the August 2012 estimate has been brought forward to be included in this report.

14.2.1  Amapola Drill Hole Database

Tetra Tech queried the database only for data identified in the drill hole collar file as belonging to the Amapola Zone and created a subset used for this Resource estimation.  At the time of the current estimate, there were 79 drill holes that had the prefix of “AM” all of which contain assay data.

The Amapola assay database contains 6,655 assays for 12,704.6 m of sampling with an average length of 1.9 m with the majority of sampling at 2 m.  All samples have been assayed by multi-element ICP analysis for 36 elements by ALS Chemex of Vancouver, as described in Section 11.0 of the 2011 Behre Dolbear Report.  Tetra Tech concludes the database provided by MPR is free of recognized errors and of excellent quality and organization.

14.2.2  Amapola Geologic Modeling

The Amapola Zone contains several vein systems at varying degrees of strike and dip that are the target of MPR exploration.  Currently four of these veins comprise adequate Ag grade and thickness to be considered as mineralized material and geologically modeled.  The four veins include the Albita, Elizabeth, Cascajal, and Julia.

The Resource is principally focused on the Albita and Elizabeth veins which together comprise a “corridor” of mineralization up to ~50 m thick.  The Albita and Elizabeth Mineral corridor is a general zone of alteration, where distinctions between the Albita and Elizabeth veins are difficult to make without the assistance of chemical assay and 3D review in advanced mining software.  The remaining veins were, for the purposes of this interpretation, considered with a lower level of geologic confidence.  The interpretation of the Cascajal and Julia veins are based on fewer data points and are in general thinner and of lower average grade.

Tetra Tech’s geologic understanding of the Amapola veins are based on review of drill hole data, drill core and most importantly discussions with MPR exploration staff regarding their interpretation of the mineralizing systems and controls at the Amapola Zone.  Geologic interpretation began with MPR’s preparation of surface geologic maps showing structural trends and alteration patterns which guided the drill hole program.  General trends established by the surface mapping are supported at depth by data collected from the 79 drill holes used in Tetra Tech’s geologic interpretation.

Planar relationships between drill hole intercepts of anomalous values of Ag, Pb, Zn were used as the main indicators defining the various veins.  Ag is the principal element for defining mineralized downhole intercepts at Amapola.  Despite a positive correlation, unlike the Cerro Los Gatos Zone, the Amapola Zone has disproportionately low base metal content.  Lack of base metal grade negated the use of metal price equivalency at the geologic interpretation stage, but is considered in the final calculation of Resources.

Ca and Ba have also been identified by MPR staff to be indicators of high-grade mineralization and have a correlation with Ag relative to elevation at the highest values of Ag; these variables assisted in defining

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the separation between the Albita and Elizabeth veins.  Unique to Amapola, Ca concentrations trend inversely and Ba trends directly with change in elevation.

Following identification of mineralized downhole intervals of Ag and assigning those intervals to corresponding planar veins, hanging wall and foot wall coordinates for each downhole interval were determined.  Hanging wall and foot wall distinctions for the Albita and Elizabeth veins were based on the general assumption that the Albita and Elizabeth mineral corridor has a strike of 351 degrees and a west dip of 88 degrees.  In places the dip is vertical or “overturned”, however hanging wall and foot wall distinctions are kept consistent with the foot wall as the east boundary and the hanging wall the west boundary.

Using the hanging wall and foot wall points, interpolated point grids were produced using MapInfo® GIS software.  One hanging wall and one-foot wall point grid was produced for each vein using a minimum curvature algorithm, which is analogous to a 3D spline fit.  The algorithm is a good estimation of the hanging wall and foot wall surface; however, the original hanging wall and foot wall points are not honored exactly.  The interpolated grid points are removed where true data points are present.  The corrected interpolated point grid for each hanging wall and foot wall is then used to create a triangulated surface using MicroMine®.

The resultant hanging wall and foot wall surface for each of the four veins were then clipped by a series of shapes.  The clipping shapes are long-section perimeters based on criteria including thickness, distance from nearest sample.

The thickness criteria are satisfied only where the difference between the hanging wall and foot wall surfaces is greater than or equal to 1 m.  Thickness was considered simultaneously with vein interval selections so internally the veins are generally greater than 1 m and are only clipped on the outside edge of the vein limiting the extent of the vein away from observed data points.

The distance from the sample clipping perimeter is determined by a perimeter extending approximately 100 m from the furthest sample on the vein plane up and down dip and along strike.  Internal distances between samples were permitted to be greater than 100 m.

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Figure 14-46:  Plan view of the Amapola Veins

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Figure 14-47:  Amapola Vein Wireframe Solids Looking Northwest

14.2.3  Amapola Assays and Composting

A total of 235 assays were identified within vein intervals for the Cerro Los Gatos zone, Table 14-15 details the assay statistics.

Table 14-15:  Assay Statistics In Modeled Veins — Amapola

Metal	Number of Samples	Mean	Minimum	Maximum
Ag	235	91.8 g/t	0.5 g/t	1450 g/t
Au	235	0.08 g/t	0.005 g/t	0.61 g/t
Pb	235	0.20%	0.001%	13.7%
Zn	235	0.38%	0.001%	19.4%
Cu	235	0.025%	0.0007%	0.8%

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Assays intervals with the vein zones were reviewed using probability plots and log normal histograms and capping values were determined, Table 14-16 details capping values and samples capped.

Table 14-16:  Assay Capping — Amapola

Metal	Capping Value	Number of Samples Capped
Ag	743 g/t	2
Au	0.51 g/t	2
Pb	0.88%	5
Zn	2.4%	6
Cu	0.24%	2

Following capping, raw assays were composited to a length of 2 m across interpreted vein intervals only.  Samples outside the interpreted veins or intervals less than 1 m after compositing were not used for estimation.

14.2.4  Amapola Variography and Search Orientation

Geostatistical modeling of spatial relationships of the Ag assay data was completed by indicator variogram analysis.  Indicator variograms were calculated on assays from within the Albita and Elizabeth mineralized corridor, encompassing the high grade within the modeled veins but also with lower grade assays present within the corridor.  Incorporating the lower grades was essential for the reliability and usefulness of the variograms.  Individual variograms were generated along strike, down dip, and down hole of the Albita and Elizabeth mineral corridor and were applied to the subordinate veins which contain limited data.

The down hole variogram in conjunction with along strike and down dip indicator variograms were used to establish a relativized nugget at 45% of the sill, using a spherical model.  The indicator variogram Indicated a sill at 130 m.  The range was limited to 100 m for the Cascajal and Julia veins.

Indicator variograms calculated for Ag were applied to the additional metals estimated in this Resource; separate variograms were not calculated for Au, Cu, Pb and Zn.

Search orientation and anisotropy ratios were determined by the author using the physical properties of the mineralized body.  Anisotropy of the search ellipse was 130:78, the third axis was the same dimension as the primary access to correct for undulations in the mostly planar veins.  The search ellipse was orientated along an azimuth of 350, with a plunge of -16 along the main axis, and a plunge of -74 along the secondary axis.  A large third axis was permissible because the estimate was contained by a vein solid in the third axis direction and the number of samples per drill hole was limited to two.

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14.2.5  Amapola Resources Estimation and Categorization

Grade estimation of the Amapola Zone Resource was completed with MicroMine® using ordinary Kriging.  A single Kriging pass was made on each of the four veins and secondary nearest neighbor Inferred pass was made only for the Albita and Elizabeth veins.

A total of four individual block models were generated from a parent model of the Amapola zone with a block size of 5x10x10.  Each block in the parent model was queried for its portion that resided within each vein solid.  Each block was indexed by column, row, level, and vein.

Each of the four models was estimated separately.  Samples only with the same vein identifier as the block model being estimated were used in the estimate for that model.

A maximum of two composite values per drill hole per block estimate were permitted, along with an absolute maximum of ten points for each block estimate.

Within the first pass all samples estimated were initially assigned to the Inferred class.  Then any block that was estimated from the first pass from samples derived from at least two drill holes and had a relative Kriging error of less than 1.04 was then assigned to the Indicated class.  Only blocks within the Albita and Elizabeth solids were eligible to be classified as Indicated.

A second nearest neighbor pass with a spherical range of 50 m used the block values estimated in the first pass cut to the 90th percentile as its source data.  Capping values from the blocks were used to prevent high-grade influencing from within the higher-grade core of the deposit and provide a more moderated estimate on the extents of the modeled veins.  Both passes were constrained inside of the 3D mineralized solids, therefor the nearest neighbor pass was subject to the 100 m down-dip and along strike extension of the solids.

The first and second pass routines were then preformed on the remaining metals (Au, Cu, Pb and Zn) in the same manner as with Ag.

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Figure 14-48 shows all vein models combined, green blocks are classified as Indicated and blue blocks are classified as Inferred.

Figure 14-48:  Amapola Vein Block Classification Looking West

14.2.6  Amapola Specific Gravity

Specific gravity measurements, which were collected on site by MPR staffing using an epoxy dip air weight/water weight specific gravity method, were provided by MPR with the project database.  Samples are taken by MPR with consideration of mineralized zones and provide good coverage to estimate the density of the veins of the Amapola Zone.

The samples that fall inside the mineralized zone were then flagged.  The flagged samples were then used for multi element statistical evaluation.  A linear fit regression was performed to determine which of the elements contributed to the specific gravity of the sample.  For Amapola, the contributing metals for specific gravity were Silver and Zinc.  The other metals did not show a statistically significant contribution to the specific gravity for this zone.  For Amapola, the regression factor for Silver is 0.0003209 and 0.0000050 for Zinc.  The base specific gravity was 2.4870111.  These values were then used to determine a specific gravity value for each composite.  The specific gravity values were then Kriged along with the other elements in the model.

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14.2.7  Cutoff Grade and Reasonable Prospects for Economic Extraction

Cutoff grade has been estimated using generalized parameters prior to detailed analysis by mining study (see Table 14-17).  The base case cutoff grade of 100 AgEq g/t accounts for typical costs to mine and process Mineral Resources and is adequate to approximate reasonable prospects for economic extraction.  Metallurgical recoveries were not changed from the 2012 estimate, when they were estimated from preliminary testwork, as additional testwork has not been completed for this deposit.  AgEq is calculated for each block in the model using $22.30/toz Ag, $0.97/lb Pb, and $0.91/lb Zn.

Table 14-17:  Cutoff Grade Parameters

	Value	Unit
Mining Cost	40	$/Tonne
Processing Cost	20	$/Tonne
G&A	3	$/Tonne
Ag Recovery	100	%
Ag Price	22.30	$/toz
Pb Recovery	95	%
Pb Price	0.97	$/lbs
Zn Recovery	88	%
Zn Price	0.91	$/lbs
Cutoff Grade	100	AgEq g/T

Gold and copper, being part of the epithermal system, are included in the Mineral Resource estimate as comparisons, however they have not been considered for purposes of determining the Ag/Eq cut-off grade since copper in the final concentrates sold would not receive any payment and gold revenue is estimated at only 2.78% of the payable metal.  Preparing a mine plan that excluded gold from the AgEq calculations had little or no bearing on estimates of ore that will be mined.  The cash flow estimates include the estimated gold revenue from the mine plan.  The incremental ore that could be processed from including the approximately 0.8 ounces AgEq from the calculation is immaterial and conservative.

Subsequent studies could materially alter any or all the parameters used to approximate a reasonably informed cutoff grade.  It is recommended that the cutoff grade is refined following the completion of a detailed project study for this deposit, including costs, metal prices, and metallurgical recovery.

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14.2.8  Amapola Deposit Resource Estimate

The Amapola Deposit Indicated Resources, effective December 21, 2012, at a base case cutoff grade of 50 g/t EqAg include 480,000 tonnes at 101 g/t Ag for 1,600,000 ounces of Ag.  Table 14-18 lists the Amapola Deposit Indicated Resources from 50 to 100 g/t EqAg cutoffs.  The Amapola Deposit Indicated Resources are shown in Figure 14-49 as grade and tonnage curves at various cutoff grades.  Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Table 14-18:  Estimated Indicated Resources — Amapola

Cutoff EqAg g/t	Tonnes	EqAg g/t	Ag g/t	Ag Ounces	Au g/t	Pb %	Zn %	Cu %
50	480,000	116	101	1,600,000	0.08	0.1	0.2	0.02
60	440,000	121	106	1,500,000	0.08	0.1	0.2	0.02
75	370,000	132	116	1,400,000	0.09	0.1	0.2	0.02
100	250,000	154	135	1,100,000	0.1	0.1	0.3	0.02

Note 1:  Based on a cut-off grade of 100 g/t silver equivalent using metal prices of $22.30/toz silver, $0.97/lb lead, and $0.91/lb zinc, metallurgical recoveries were not considered.

Note 2:  Figures may not total due to rounding of significant figures.

Note 3:  Indicated Resources are equivalent to US SEC Industry Guide 7 “Mineralized Material”.

Figure 14-49:  Grade Tonnage Curve Indicated Amapola Deposit

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The Amapola Deposit Inferred Resources, effective December 21, 2012, at a base case cutoff grade of 50 g/t EqAg include 5,980,000 Tonnes at 106 g/t Ag for 20,300,000 ounces of Ag.  Table 14-19 lists the Amapola Deposit Inferred Resources from 50 to 100 g/t EqAg cutoffs.  The Amapola Deposit Inferred Resources are shown in Figure 14-50 as grade and tonnage curves at various cutoff grades.

Table 14-19:  Estimated Inferred Resources — Amapola

Cutoff EqAg g/t	Tonnes	EqAg g/t	Ag g/t	Ag Ounces	Au g/t	Pb %	Zn %	Cu %
50	5,980,000	125	106	20,300,000	0.09	0.1	0.3	0.02
60	5,410,000	133	112	19,500,000	0.09	0.2	0.3	0.03
75	4,450,000	147	125	17,900,000	0.1	0.2	0.3	0.03
100	3,440,000	164	140	15,500,000	0.1	0.2	0.3	0.03

Note 1:  Figures may not total due to rounding of significant figures.

Figure 14-50:  Grade Tonnage Curve Inferred Amapola Deposit

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14.3  Esther Zone

14.3.1  Esther Drill Hole Database

Tetra Tech queried the database only for data identified in the drill hole collar file as belonging to the Esther Zone and this data was used for the Resource estimation.  At the time of this estimate there were 42 drill holes with the prefix of “ES” all of which contain assay data.  Drill hole ES-42 is the cut off for this report.

The Esther assay database contains 4,959 assays for 9,715.07 m of sampling with an average length of 1.9 m with the majority of sampling at 2 m.  All samples have been assayed by multi-element ICP analysis for 36 elements by ALS Chemex of Vancouver, as described in Section 11.0 of the report.  Tetra Tech concludes the database provided by MPR is free of recognized errors and of excellent quality and organization.

14.3.2  Esther Geologic Modeling

The Esther Zone contains several vein systems at varying degrees of strike and dip that are the target of MPR exploration.  Tetra Tech modeled one main vein, as well as two secondary veins that appear to be associated with the main vein.  The main vein was modeled with higher confidence than the two secondary veins.  The main vein is approximately 1 to 10 m thick.  Wireframes are shown below in Figure 14-51 and Figure 14-52.

Tetra Tech’s geologic understanding of the Esther vein is based on review of drill hole data, drill core and most importantly discussions with MPR exploration staff regarding their interpretation of the mineralizing systems and controls at the Esther Zone.  Geologic interpretation began with MPR’s preparation of surface geologic maps showing structural trends and alteration patterns which guided the drill hole program.  General trends established by the surface mapping are supported at depth by the drilling data collected and used for Tetra Tech’s geologic interpretation.

Following identification of mineralized downhole intervals of Ag and assigning those intervals to corresponding planar veins, hanging wall and foot wall coordinates for each downhole interval were determined.  Hanging wall and foot wall distinctions for the Esther zone were based on the general assumption that the Esther mineral corridor has a strike of 110 degrees and a southwest dip of -57 degrees.

Using the hanging wall and foot wall points, interpolated point grids were produced using MapInfo® GIS software.  One hanging wall and one-foot wall point grid was produced for each vein using a minimum curvature algorithm.  The algorithm is a good estimation of the hanging wall and foot wall surface; however, the original hanging wall and foot wall points are not honored exactly.  The interpolated grid points are removed where true data points are present.  The corrected interpolated point grid for each hanging wall and foot wall is then used to create a triangulated surface using MicroMine®.

The resultant hanging wall and foot wall surface for each of the three veins were then clipped by a series of shapes.  The clipping shapes are long-section perimeters based on criteria including thickness and distance from nearest sample.

The thickness criteria is satisfied only where the difference between the hanging wall and foot wall surfaces is greater than or equal to 1 m.  Thickness was considered simultaneously with vein interval

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selections so internally the veins are generally greater than 1 m and are only clipped on the outside edge of the vein limiting the extent of the vein away from observed data points.  The thickness criteria was relaxed in areas of the Esther Offset and Esther Footwall veins, these vein were permitted to be a minimum of 0.5 m thick in some areas as a result all blocks from these veins were restricted to Inferred.

The distance from the sample clipping perimeter is determined by a perimeter extending approximately 100 m from the furthest sample on the vein plane up and down dip and along strike.  Internal distances between samples were permitted to be greater than 100 m.

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Figure 14-51:  Esther Zone Wireframes Looking Northwest

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Figure 14-52:  Esther Zone Wireframes Looking North

14.3.3  Esther Assays and Composting

A total of 88 assays were identified within vein intervals for the Esther Deposit, Table 14-20 details the assay statistics.

Table 14-20:  Assay Statistics in Modeled Veins — Esther

Metal	Number of Samples	Mean	Minimum	Maximum
Ag	88	94.8 g/t	1 g/t	857 g/t
Au	74	0.10 g/t	0.005 g/t	0.93 g/t
Pb	88	1.0%	0.031%	7.7%
Zn	88	2.2%	0.031%	17.4%
Cu	88	0.037%	0.0006%	0.63%

Assays intervals with the vein zones were reviewed using probability plots and log normal histograms and capping values were determined, Table 14-21 details capping values and samples capped.

Table 14-21:  Assay Capping — Esther

Metal	Capping Value	Number of Samples Capped
Ag	480 g/t	2
Au	1.0 g/t	0
Pb	5.0%	5
Zn	7.8%	4
Cu	0.22%	1

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Following capping, raw assays were composited to a length of 2 m across interpreted vein intervals only.  Samples outside the interpreted veins or intervals less than 1 m after compositing were not used for estimation.

14.3.4  Esther Variography and Search Orientation

Geostatistical modeling of spatial relationships of the Ag assay data was completed by indicator variogram analysis.  Indicator variograms were calculated on assays from within the three main veins and the mineralized corridor, encompassing the high grade within the modeled veins but also with lower grade assays present within the corridor.  Individual variograms were generated along strike, down dip, and down hole of the mineral corridor.

The down hole variogram in conjunction with along strike and down dip indicator variograms were used to establish a relativized nugget at 37% of the sill, using a spherical model.  The indicator variogram Indicated a sill at 130 m.

Indicator variograms calculated for Ag were applied to the additional metals estimated in this Resource; separate variograms were not calculated for Au, Cu, Pb and Zn.

Search orientation and anisotropy ratios were determined by the author using the physical properties of the mineralized body.  Anisotropy of the search ellipse was 130:65:26.  The search ellipse was orientated along an azimuth of 110, with a plunge of 8 along the main axis, and a plunge of -64 along the secondary axis.

14.3.5  Esther Resources Estimation and Categorization

Grade estimation of the Esther Zone Resource was completed with MicroMine® using ordinary Kriging.  An initial Kriging pass was made on each of the three veins and followed by secondary nearest neighbor Inferred pass.

A total of three individual block models were generated from a parent model of the Esther zone with a block size of 10x5x10.  Each block in the parent model was queried for its portion that resided within each vein solid.  Each block was indexed by column, row, level, and vein.

Each of the three models was estimated separately.  Samples only with the same vein identifier as the block model being estimated were used in the estimate for that model.

A maximum of two composite values per drill hole per block estimate were permitted, along with an absolute maximum of ten points for each block estimate.

Within the first pass all samples estimated were initially assigned to the Inferred class.  Then any block that was estimated from the first pass from samples derived from at least two drill holes and had a relative Kriging error of less than 1.25 was then assigned to the Indicated class.  Only blocks within the main vein solid were eligible to be classified as Indicated.

A second nearest neighbor pass with a spherical range of 50 m used the block values estimated in the first pass cut to the 90th percentile as its source data.  Cut values from the blocks were used to prevent high-grade influence from within the higher-grade core of the deposit and provide a more moderated estimate on the extents of the modeled veins.  Both passes were constrained inside of the 3D mineralized solids, therefor the nearest neighbor pass was subject to the 100 m down-dip and along strike extension of the solids.

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The first and second pass routines were then preformed on the remaining metals (Au, Cu, Pb and Zn) in the same manner as with Ag.

Figure 14-53 shows all vein models combined, green blocks are classified as Indicated and blue blocks are classified as Inferred.

Figure 14-53:  Esther Zone Block Classification Looking North

14.3.6  Esther Specific Gravity

Specific gravity measurements, which were collected on site by MPR staffing using an epoxy dip air weight/water weight specific gravity method, were provided by MPR with the project database.  Samples are taken by MPR with consideration of mineralized zones and provide good coverage to estimate the density of the veins of the Esther Zone.

The samples that fall inside the mineralized zone were then flagged.  The flagged samples were then used for multi element statistical evaluation.  A linear fit regression was performed to determine which of the elements contributed to the specific gravity of the sample.  For Esther, the contributing metal for specific gravity was and Zinc.  The other metals did not show a statistically significant contribution to the specific gravity for this zone.  For Esther, the regression factor 0.00000244 for Zinc.  The base specific gravity was 2.53515226.  These values were then used to determine a specific gravity value for each composite.  The specific gravity values were then Kriged along with the other elements in the model.

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14.3.7  Cutoff Grade and Reasonable Prospects for Economic Extraction

Cutoff grade has been estimated using generalized parameters prior to detailed analysis by mining study (see Table 14-22).  The base case cutoff grade of 100 AgEq g/t accounts for typical costs to mine and process Mineral Resources and is adequate to approximate reasonable prospects for economic extraction.  Metallurgical recoveries were not changed from the 2012 estimate, when they were estimated from preliminary testwork, as additional testwork has not been completed for this deposit.  AgEq is calculated for each block in the model using $22.30/toz Ag, $0.97/lb Pb, and $0.91/lb Zn.

Table 14-22:  Cutoff Grade Parameters

	Value	Unit
Mining Cost	40	$/Tonne
Processing Cost	20	$/Tonne
G&A	3	$/Tonne
Ag Recovery	100	%
Ag Price	22.30	$/toz
Pb Recovery	95	%
Pb Price	0.97	$/lbs
Zn Recovery	88	%
Zn Price	0.91	$/lbs
Cutoff Grade	100	AgEq g/T

Gold and copper, being part of the epithermal system, are included in the Mineral Resource estimate as comparisons, however they have not been considered for purposes of determining the Ag/Eq cut-off grade since copper in the final concentrates sold would not receive any payment and gold revenue is estimated at only 2.78% of the payable metal.  Preparing a mine plan that excluded gold from the AgEq calculations had little or no bearing on estimates of ore that will be mined.  The cash flow estimates include the estimated gold revenue from the mine plan.  The incremental ore that could be processed from including the approximately 0.8 ounces AgEq from the calculation is immaterial and conservative.

Subsequent studies could materially alter any or all the parameters used to approximate a reasonably informed cutoff grade.  It is recommended that the cutoff grade is refined following the completion of a detailed project study for this deposit, including costs, metal prices, and metallurgical recovery.

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14.3.8  Esther Deposit Resource Estimate

The Esther Deposit Indicated Resources, effective December 21, 2012, at a base case cutoff grade of 50 g/t EqAg include 620,000 tonnes at 113 g/t Ag for 2,300,000 ounces of Ag.  Table 14-23 lists the Esther Deposit Indicated Resources from 50 to 100 g/t EqAg cutoffs.  The Esther Deposit Indicated Resources are shown in Figure 14-54 as grade and tonnage curves at various cutoff grades.  Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Table 14-23:  Estimated Indicated Resources — Esther

Cutoff EqAg g/t	Tonnes	EqAg g/t	Ag g/t	Ag Ounces	Au g/t	Pb %	Zn %	Cu %
50	620,000	181	113	2,300,000	0.04	0.6	1.7	0.02
60	600,000	185	116	2,200,000	0.04	0.6	1.7	0.02
75	580,000	191	119	2,200,000	0.04	0.6	1.8	0.02
100	460,000	217	133	2,000,000	0.04	0.7	2.1	0.02

Note 1:  Based on a cut-off grade of 100 g/t silver equivalent using metal prices of $22.30/toz silver, $0.97/lb lead, and $0.91/lb zinc.  Metallurgical recoveries were not considered.

Note 2:  Figures may not total due to rounding of significant figures.

Note 3:  Indicated Resources are equivalent to US SEC Industry Guide 7 “Mineralized Material”.

Figure 14-54:  Grade Tonnage Curve Indicated Esther Deposit

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The Esther Deposit Inferred Resources, effective December 21, 2012, at a base case cutoff grade of 50 g/t EqAg include 2,940,000 Tonnes at 87 g/t Ag for 8,200,000 ounces of Ag.  Table 14-24 lists the Esther Deposit Inferred Resources from 50 to 100 g/t EqAg cutoffs.  The Esther Deposit Inferred Resources are shown in Figure 14-55 as grade and tonnage curves at various cutoff grades.

Table 14-24:  Estimated Inferred Resources — Esther

Cutoff EqAg g/t	Tonnes	EqAg g/t	Ag g/t	Ag Ounces	Au g/t	Pb %	Zn %	Cu %
50	2,940,000	205	87	8,200,000	0.1	1.3	2.5	0.04
60	2,720,000	218	93	8,100,000	0.11	1.4	2.6	0.05
75	2,600,000	225	95	7,900,000	0.11	1.4	2.7	0.05
100	2,290,000	243	98	7,200,000	0.12	1.6	3	0.05

Note 1:  Figures may not total due to rounding of significant figures.

Note 2:  Inferred Resources are not defined or recognized by US SEC Industry Guide 7.  Acceptable in proposed New Guidelines.

Figure 14-55:  Grade Tonnage Curve Inferred Esther Deposit

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14.4  Relevant Factors

For Los Gatos deposits estimation domains have been constructed using AgEq grades that include Ag, Pb, and Zn.  Significant changes in any or all the three metals values could materially affect the boundary of a +150 AgEq g/t grade-shell across vein thickness and in the plane of the vein.  The grade tonnage curves of Measured and Indicated Mineral Resources indicate that blocks within the grade-shell are relatively insensitive to cutoff increases up to 200 AgEq g/t, however sensitivity analysis has not been performed with grade-shells constructed at various cutoffs.

Additional factors that could materially affect the Mineral Resources are cutoff grade parameters that have not been developed in detail with an optimized mining plan.  Subsequent mining studies could materially alter any or all of the parameters used to estimate a reasonably informed cutoff grade.

There are no additional environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that the author of this report is aware of that could materially affect the Mineral Resource estimate.

14.5  Conclusions and Recommendations

14.5.1  Geology and Resources

Project geologic and drill hole data has been collected and analyzed by MPR using industry standard methods and practices and is sufficient to characterize grade and thicknesses of the deposit and to support the estimation of Measured, Indicated and Inferred Mineral Resources.  Although the deposit has been densely drilled, Resource expansion potential and project upside exist in the immediate deposit area as well as at other identified prospects such as the Amapola and Esther, which have been preliminarily investigated with drilling showing Indicated and Inferred Mineral Resources, and other prospects throughout the land package.

Drilling at the site has proven to be productive.  Additional Resources have been identified in each drilling campaign.  The drilling for each of the latest Resource Estimates vs the AgEq Oz are shown in Figure 14-56, below.  Based on these positive results, Tetra Tech recommends exploration at the site continues.

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Figure 14-56:  Additional drilling and total AgEq oz per estimate

14.5.1.1  Data Verification

The current QA/QC program is adequate and supports the conclusion that data collected, and the monitoring of QC data is reliable for the purposes of estimating Resources; however, additional improvements are required for the QA/QC program to align with industry best practice and facilitate more meaningful QC.

Clerical corrections and optimization of standard reference material is necessary to assess laboratory analytical performance.  Standard performance for Ag, Pb, Zn, Au, and Cu are good, with few results outside of +/-2 standard deviations of the certified value.  To leverage the use of standards, material should be sourced closer to the range of the deposits average grade for Ag, Pb and Zn.  Most of the standards used are too low-grade Ag, Pb, and Zn.

Field duplicate testing has shown good reproducibility; however, current protocols do not adequately test the variability of the deposit within the likely mining areas.  Field duplicates contain too few ore-grade samples chosen from within the vein.  The field duplicates that have been analyzed and are above 100 g/t Ag show a similar range of variability as the sample pairs below 100 g/t Ag.  Collecting more field duplicates from within the expected mining area will help to evaluate the variability that could be encountered.

In-stream blank material analysis for Ag has demonstrated acceptable sample preparation and laboratory performance for Ag; the performance for Pb, Zn, and Cu show many samples with values several times the detection limit, and exceedances are not significant in relation to the average deposit grade.

Exceedances could be a result of background levels of Pb, Zn, and Cu in the blank material or a result of contamination from sample preparation or analysis.  Because the Ag blank testing has shown few failures, it is possible that the blank material contains base amounts of Pb, Zn, and Cu; however, the blank failures often visually correlate with preceding samples of higher-grade.

Umpire (third party) sampling should be conducted to meet industry standard practices to confirm the analyses performed at ALS Chemex.

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14.6  Recommendations

Tetra Tech recommends the following to continue exploration and possible expansion of the Los Gatos Mineral Resource base within the property, as follows:

·Recognizance and in-fill drilling to test the following areas:

·Down-dip and along strike to the Northeast of the NW Block;

·Up-dip of the Central Block extending the high-grade mineralized zone along strike of the bulk sample area;

·At the hanging-wall detached blocks currently classified as Inferred Resources of the NW block;

·At the plunging mineralized concentration of the SE3 Block with additional potential of the Block’s extension; and

·Down-dip of the Central Block along the high-grade zone intercepted by drill holes GA-55, GA-66, and GA-243 to determine possible continuity;

·Additional infill drilling is recommended at the Amapola and Esther zones to delineate the possible extension of the identified Mineral Resources and assess full Resource potential.  If the results are positive, the identified Mineral Resources should be updated with a Scoping Study to determine if they may contribute to the Los Gatos economics;

·Complete detailed surface mapping and sampling in the area to define and prioritize other probable prospects within the project’s area;

·Geophysical surveys should complement the prospects prioritization prior to perform drilling exploration.

14.6.1  Standards

The following recommendations are to improve the current QA/QC protocols:

·Check sample standards relative to Ag, Pb, Zn, Cu, and Au. Correct assay records and database when errors are recognized;

·Clerical errors rather than actual analysis failure are the most frequent issue.  This can be corrected by comparing all five payable elements.  Spikes are visually observed when standard assays issues are encountered; and

·Keep records of sample results in the database for a quick review to confirm possible clerical errors.

·Submit with more frequency ore-grade standards from the mineralized zone:

·Identify on the sample submittal sheet the ore-grade standards to be tested directly using the assay methods OG62 or GRA21.  This action will avoid falling short of the 30-grams required for re-runs by the ME-ICP61 and 50 grams required to complete the assays; and

·Include additional sulfide standards that are as close as possible to grades of 250 g/t Ag, 2.5% Pb and 6% Zn, while assays for Au and Cu are of secondary importance and should only be evaluated after the economic metals when selecting ideal standards.

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·Generate several custom standards from material collected from the area of the bulk sample, 500, 50-gram standards could be made from 25 kilograms of material. Internationally certified laboratories offer custom standard services.

14.6.2  Blanks

Tetra Tech recommends improve blank samples use by:

·Inserting blanks more frequently after assaying high-grade samples to ensure the sample preparation equipment has been cleaned properly;

·Review geologic cross-sections to optimize the insertion of blanks within the vein intersections;

·Currently, the intersected veins show approximately 1,400 2 m intervals; while the insertion of blank samples represents about 20% (280 samples) of the sampling population.

·Investigate the use of a coarser certified blank so that crushing, splitting, and pulverizing equipment at the preparation facility are operating in a similar condition to a normal interval core sample;

·Submit blank samples directly for re-run testing to test OG62 or GRA21 equipment.  Present protocols do not test potential contamination of OG62 or GRA21 equipment because the first test never triggers re-runs;

·Record the sample ID of the sample tested before each of the blanks.  This will enable assessment of blanks in the context of possible contamination from the sample preparation and the sample analyzed before the blanks.  Any poor blank performance following high-grade samples should trigger re-runs of several samples following the high-grade sample;

·Source certified or self-certify blank material;

·Blanks obtained from core are best because the lab is blind to the control sample and both the laboratory preparation and analysis are checked.  Blanks could be sourced from splits of andesite that have been tested and returned detection limit results for Ag, Pb, Zn, Au, and Cu.  If blank core cannot be sourced andesite outcroppings should be considered;

·Untested “barren” full core should not be used as blank material because of the risk that it contains low background levels of Pb, Zn, and Cu as seen in the current blanks; and

·If certifiable blank core is limited, blanks can be submitted as reduced weight (not the equivalent of a 2 m core split), with only enough material to produce a coarse reject and pulp.

·Ensure the blank material is stored away from the core preparation area and the blank material sample bagging is completed in a clean environment; and

·Request the laboratory’s internal blank control sample to potentially evaluate or rule out contamination.

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14.6.3  Duplicates

To improve duplicate use for QA/QC Tetra Tech recommends:

·Insert instream field duplicates in areas that visually appear to be high-grade.  Doing so will test the prep lab’s ability to homogenize and provide data to evaluate the nugget effect; and

·Insert an empty bag with a sample ID tag and instruct the lab to generate a pulp split (pulp duplicate).  Analyzing pulp duplicates will provide data to evaluate the reproducibility of the sample analysis and suggest an error range of analysis.

14.6.4  Umpire Sampling

It is industry best practice to have 5% of sample pulps tested at a third-party lab.  Aside from limited re-sampling completed as part of past NI 43-101 reports, it is not apparent that umpire sampling has been performed.

As discussed, a small percentage of the total drill hole database encounters the area of interest, because of this umpire sampling should be focused on resampling most of the pulps within the expected mining areas.  Re-analyzing half of the pulps from all vein samples would account for approximately 750 samples and provide additional confidence in the database.

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15.0  MINERAL RESERVE ESTIMATES

The Mineral Reserve estimate, effective July 2020, includes the portion of the Measured and Indicated Resource that can be mined economically.  Economic criteria and mining constraints (based on the selected mining methods) are applied to the Resource blocks to define mineable blocks.  Mineral Reserves are determined after applying dilution and recovery factors to these mineable blocks.

The Mineral Reserve presented has been calculated from the mine plan created from the September 2019 Resource update.  The Reserve estimate herein is consistent with the Canadian Institute of Mining Standards on Mineral Resources and Mineral Reserves.

15.1  Net Smelter Return

The parameters used in the calculation of NSR in the block model (including metal values, recovery factors, transportation costs, etc.) were provided by SSMRC and reviewed by Tetra Tech.

NSR = rev – (ref + treat + trans)

Where

rev = dollar revenue per tonne mined

ref = concentrate refining cost per tonne mined

treat = concentrate treatment cost per tonne mined

trans = concentrate transportation cost per tonne mined

NSR has been calculated from the revenue/cost variables and recoveries presented in Section 21.  The primary parameter used to determine stopes was a $75 NSR cutoff.  Projected revenues from the sale of silver, gold, zinc, and lead are based upon long term consensus prices of $18.99/oz Ag, $1,1472/oz Au, $1.09/lb Zn, and $0.87/lb Pb respectively.

A script was used to assign NSR values to the block model as a new variable.  The script was written so that NSR values were assigned only to Measured and Indicated Resource blocks.  All other blocks in the model were assigned a zero-dollar value.  Figure 15-1 presents a grade tonnage curve including all blocks in the model with an NSR value greater than zero.

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Figure 15-1:  NSR Value / Tonnage Plot

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15.2  Ore Body Description

Mineralization is confined to a series of veins that run parallel to a group of faults that run Northwest to Southeast across the property.  The veins lie, for the most part, on the footwall side of these faults.  The primary fault is the Los Gatos Fault.  The veins are steep in the Northwest Zone (NWZ) at 70°–85°.  In the Central Zone (CZ), Southeast Zone (SEZ), and Southeast Zone 2 (SEZ 2), the veins are 60° near surface, flattening to 45° at depth.  Figure 15-2 is an orthogonal view showing the vein orientation and flattening with depth.

Figure 15-2:  Orthogonal View showing Flattening Veins

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The veins and main Los Gatos Fault are offset by several younger cross-cutting faults that break the mineralized veins into five distinct zones.  The faulting involves both lateral offsets and relative rotation of each of the blocks.  This resulted in each block being reviewed separately for mining method, recovery factors, and dilution parameters.  Figure 15-3 is a plan view showing the mineralized veins and faulted offsets.

Figure 15-3:  Plan View of Mineralized Veins

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For this study, the zones marked “Southeast Zone” and “Southeast 2 Zone” in Figure 15-3 were combined and developed as a single entity called the SEZ because of their proximity to each other.  The more distal “Southeast 3 Zone” has been planned and developed separately as SEZ 2.

Figure 15-4 presents a cross-section through the block model showing blocks colored by NSR value.  Also shown is a typical diamond drilling section, wireframe shapes for the mineralized veins, and the Los Gatos Fault.

Figure 15-4:  Block Model Section

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15.3  Mining Method

Each of the mineralized areas has been reviewed separately, and a combination of safety, vein orientation, economic, and geomechanical considerations are used to select the most appropriate mining method for each.  In the NWZ, the mineralized veins are steeply dipping and offset from the Los Gatos fault zone.  This, combined with local rock conditions, allowed for the selection of longhole.  Blocks of stopes that are thicker than 9 m (hanging wall to footwall) are scheduled to be mined transversely.  Thinner areas (less than 9 m) are planned to be mined longitudinally.

A mix of overhand and underhand mining has been implemented, starting on the 1390 Levels in both the NWZ and the CZ.  Due to the underhand approach, sill pillars are designed at necessary intervals in both cut-and-fill (CAF) and longhole stoping (LHS) blocks.  The underhand longhole stoping areas require re-mining or undercutting of the bottom sill drifts compared to the overhand approach.  This method will be reviewed by the mine planning team throughout the mine life and evaluated with current conditions in the mine.

The stopes within each of the mining zones were created using the Maptek Stope Optimizer (MSO) from Alford Mining Systems.  The software was run with Vulcan to generate stopes according to a set of parameters specific for each mining area.

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15.4  Dilution and Recovery Estimates

15.4.1  Dilution

The dilution strategy was developed by Stantec’s geomechanical consultant.  Per the resulting recommendations, 2 m of hanging wall dilution was added to each transverse and longitudinal stope in the NWZ (see Figure 15-5).  This dilution was considered sufficient to cover the combination of hanging wall and footwall dilution expected in NWZ stopes.  The original stope geometries were used to create a 2 m slice to query the block model for the grade of the resulting dilution.  The in-situ tonnes and grade of the stope were combined with the tonnes and grade of the 2 m slice to calculate the diluted tonnes and grade for the stope.  A recovery factor was then applied (as detailed in Section 15.4.2 ) to yield diluted and recovered tonnes and grade for the stope.

Figure 15-5:  NWZ Longhole Stope Dilution

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A different dilution strategy was applied to the drift-and-fill stopes in the CZ, SEZ, and SEZ 2.  The stopes will be mined longitudinally along strike and were divided into two groups: stopes less than 9 m wide (which can be mined in a single pass), and stopes greater than 9 m wide, which require two or more drifts. In the case of the narrower stopes, 0.2 m of wall and a portion of the 0.15 m of back dilution were added to the in-situ stope tonnes (see Figure 15-6).  The grade of the slices on the walls and back were queried using shape files and the block model on a stope-by-stope basis.  If the stope was vertically below a planned stope, then the back dilution was removed from the planned stope above.  Stopes wider than 9 m will be mined in multiple passes (primary/secondary sequence), with only a portion of back and outside wall dilution applied to the “outside” stopes (see Figure 15-7).  A total of 4.7% of the final diluted stope volume was added to all the planned drift-and-fill stopes for overbreak in rock, with an additional 2.4% with a reduced density of 2.12 t/m3 and a zero grade to represent paste mucked during mining.

Figure 15-6:  Drift-and-Fill Overbreak

Figure 15-7:  Drift-and-Fill Dilution

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Dilution by mining method is summarized in Table 15-1.

Table 15-1:  Dilution by Mining Method

Mining Type	Dilution
Transverse Longhole	17.3	%
Longitudinal Retreat Longhole	28.5	%
Drift-and-Fill	7.1	%

Dilution will be refined through the addition of mine experience as the workings progress.  The addition of infill drilling will also allow for a better estimation of dilution for the short term mine plan going forward at the site.  It is recommended to continue to evaluate the dilution on a regular basis and adjust it as necessary for future mine plans.

15.4.2  Mining Recovery

15.4.2.1  Stope Recovery

Ore recovery estimates for longhole stopes have been based on the designed stope dimension of 12 m W × 20 m H × 12 m L (average), with the 5 m W x 5 m H access/drill drift volume removed.  This yields a planned stope volume of 2,580 m3.

Several factors influence recovery, including line of sight from the remote mucking stand, distance to the muck pile, and blasting issues.  Line of sight and maneuverability may prevent the LHDs from accessing muck from the front corners of the stope.  It is assumed that the maximum angle the LHD can operate from the drawpoint is be approximately 45°.  Also, cleanup at the back of the stope can be difficult to gauge and results in additional loss of ore.  Summing these yields a volume of 34 m3 due to mucking losses.  Comparing this to the stope design volume of 2,580 m3 produces a mining design recovery of 98.7%.

Production blasting in large excavations presents issues that affect ore recovery, such as oversized rock and un-blasted rock left on the walls.  These factors represent an average ore loss of 2.7%. See Figure 15-8 for a visual presentation of ore losses in a stope.

Figure 15-8:  Isometric View of Stope

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Mucking complications, blasting limitations, and unplanned ore losses result in an overall mining recovery from longhole stopes of 96%.

Mining recovery should be evaluated on a continual basis with data from the mine and incorporated into the new mine plans going forward.

15.4.3  Drift-and-Fill Recovery

The average drift-and-fill production heading has been assumed to be 5 m H × 5 m W.  Mining losses are expected to occur when rock cannot be removed from the floor.  This typically happens when the floor is uneven and some of the rock settles to smooth the floor or is lost into the underlying paste fill.  Non-recoverable ore is assumed to be 0.1 m in depth (see Figure 15-9).  The resulting ore recovery for Drift-and- Fill mining is 98% of the planned excavation.

Figure 15-9:  Recovery Profile (Drift-and-Fill)

Recovery by mining method is summarized in Table 15-2.

Table 15-2:  Recovery by Mining Method

Mining Type	Recovery
Transverse Longhole	96	%
Longitudinal Retreat Longhole	96	%
Drift-and-Fill	98	%

Mining recovery should be evaluated on a continual basis with data from the mine and incorporated into the new mine plans going forward.

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15.5  Mineral Reserves

The mining parameters as defined above are applied to the Resources to convert them into Reserves. After application of mining dilution and recovery factors to the stope tonnage, the data was transferred from Vulcan to Deswik scheduling software to develop a life-of- mine (LOM) development and production schedule.  Figure 15-10 below shows the mine plan with the 2019 Resource model Measured and Indicated blocks.  The material mined through June 2020 has been removed.  Reserves are the economically extractable portion of the Resource.  During the scheduling process, a small amount of additional dilution was added as development drifts were mined through individual stopes.  Typically, this is material from the hanging wall side of the ore zone as cross-cutting drifts pass through the ore.

Table 15-3 presents the total Mineral Reserve, effective July 1, 2020, for Los Gatos, which includes dilution and recovery factors.  The Reserves have been calculated at an NSR cutoff of $75 and exclude material mined at the project since mining started in November 2018.  The NSR cutoff is representative of the site operating conditions. It is noted that the Proven Reserves exceed the reported Measured Resources, this is due to the mining dilution exceeding the mineralized material lost due to the mining recovery factor, and the difference in cutoff parameters used in the Resource and Reserve estimates. The Resource estimate was reported at a cutoff of 150 g/t AgEq in September 2019. The Reserves, produced in July 2020, were reported at a $75 NSR cutoff, to represent current site operating conditions and updated metal prices. The $75 NSR cutoff allows for the economic extraction of additional mineralized material from the Resource that were estimated at a cutoff of 150 g/t AgEq.

Table 15-3:  Mineral Reserve

Zone	Classification	Tonnes	Ag (g/t)	Au (g/t)	Pb (%)	Zn (%)
NWZ	Proven	2,587,684	359	0.43	3.09	5.88
	Probable	492,892	333	0.34	2.86	5.88
CZ	Proven	3,767,456	314	0.31	2.55	5.32
	Probable	1,772,921	299	0.44	2.32	5.82
SEZ	Proven	5,751	148	0.16	3.69	7.23
	Probable	569,380	148	0.16	3.69	7.23
SEZ2	Probable	421,547	118	0.17	3.11	4.16
Total	Proven	6,360,890	332	0.36	2.77	5.55
Total	Probable	3,256,740	254	0.34	2.74	5.86
Total	Proven + Probable	9,617,631	306	0.35	2.76	5.65

Figure 15-10 shows the mine plan with the Resource model blocks. Measured blocks are shown in red and Indicated blocks in green. In general, approximately 66% of the Resources were converted to Reserves.

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Figure 15-10:  LOM plan showing mining, development, and the 2019 Resource model, with Measured blocks in red and Indicated blocks in green.

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Blocks in the September 2019 Resource model were flagged within the mine plan shapes used for calculating the July 2020 Reserves. These tonnes were excluded from the Resources, and the remaining undiluted Resources are reported in Table 15-4.

There are several factors to note when evaluating these numbers. The cutoff for reporting these Resources is 150 g/t AgEg and are undiluted. Reserves were reported at a $75 NSR cutoff and include recoveries and dilution.

Table 15-4: Resources exclusive of Reserves

Classification	Tonnes	AgEq g/t	Ag g/t	Pb %	Zn %	Au g/t	AgEq toz M	Ag toz M	Pb lbs M	Zn lbs M	Au toz K
Measured	1,319,893	442	181	2.4	4.5	0.39	19	8	71	131	16
Indicated	2,200,669	368	139	2.1	4.2	0.23	26	10	101	205	17
Measured & Indicated	3,520,562	395	154	2.2	4.3	0.29	45	17	172	337	33
Inferred	3,717,063	361	107	2.8	4.0	0.28	43	13	231	330	34

NOTES:

1) 150 AgEq g/t cutoff grade has been calculated using $18/toz Ag, $0.92/lbs Pb, and $1.01/lbs Zn,

2) Columns may not total due to rounding,

3) Mineral Resources are stated as undiluted.

4) One troy ounce (toz) is equal to 31.1035 grams (g) and one Tonne is equal to 2,204.62 lbs.

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15.6  Conclusions and Recommendations

15.6.1  Conclusions

The Reserves have been calculated from the September 2019 Resource update and excludes material that has been mined since the Resource update.  The general mining plan is currently a mix of overhand and underhand mining, with the CZ being mined completely from the top down.

15.6.2  Recommendations

It is recommended to re-estimate the project Reserves when additional Resources are estimated.  A new short term mine plan should be completed regularly with the updated Resources and additional information obtained during mining.  Dilution should be adjusted as additional mining data is obtained from the working areas of the mine and evaluated on a continual basis throughout the mine life.

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16.0  MINING METHODS

Mining methods have been determined for the site and are currently being utilized for production at the mine.

16.1  Geotechnical

16.1.1  Geomechanical Investigation

Available geotechnical holes were reviewed and processed to extract pertinent geomechanical information.  The information was then grouped by mineralized zone and rock type.  The information obtained included Q’, Rock Mass Rating (RMR), and point load test data converted to unconfined compressive strength (UCS).  From the RMR, Geological Strength Index (GSI) values were estimated and rock mass properties were obtained (Stantec, September 2015).  The geomechanical hole information was supplemented by processing the underground mapping performed by SSMRC geologists to complete a wedge analysis.  Furthermore, laboratory testing and in-situ stress measurements were completed on core in order to adjust in-situ rock properties and parameters used for numerical modeling.

16.1.1.1  Geomechanical Holes Compilation

The NWZ has 23 geotechnical holes available for analysis.  The CZ has 68 geotechnical holes available for analysis.  The SEZ has 25 geotechnical holes available for analysis.  Table 16-1 presents the values obtained for Q’ and RMR for each zone.

Table 16-1:  Rock Mass Rating

Rock	Northwest Zone		Central Zone		Southeast Zone
Type	Q’	RMR	Q’	RMR	Q’	RMR
Andesite	9.62	46.86	13.47	42.99	9.51	10.67
Mineralized Andesite	21.88	38.74	20.71	50.39	13.60	12.27
Fault Breccia	—	—	21.71	43.92	—	—
Dacite	—	—	10.28	41.89	—	—
Epiclastic	—	—	21.14	38.79	17.63	7.92
Rhyolite	16.57	43.94	27.87	49.37	16.70	12.16
Tuff	7.78	41.17	13.48	44.39	8.81	12.91

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Point load testing was completed on a total of 117 holes.  Point load test data was compiled for all the Los Gatos mineralized zones.  Table 16-2 presents the UCS values obtained from the point load test results.

Table 16-2:  Point Load Test Compilation Results for All Mineralized Zones

Rock	Point Load (UCS MPa)
Type	No. of Points	Mean	Std. Dev.	Min.	Max.
Andesite	1,008	24.29	24.01	0.000100	221.99
Mineralized Andesite	64	31.42	35.12	0.009000	189.50
Fault Breccia	8	14.60	23.78	0.000580	61.00
Epiclastic	1,085	4.49	19.45	0.000035	556.56
Rhyolite	304	29.57	38.46	0.002000	221.99
Tuff	405	30.02	25.65	0.001000	148.56

16.1.1.2  Field Data Compilation

To supplement geomechanical hole information, a total of 1,184 joints and faults were mapped by SSMRC geologists in the andesite portion of the ramp.  This allowed Stantec to complete a stereo-net and wedge analysis.  The stereo-net analysis identified four joint sets, as listed in Table 16-3.  Joint set number 1 is the most prominent.  A case could be made that joint set number 2 and number 3 represent only one joint set.  Therefore, the rock mass ranges from only three joint sets to three joints sets plus random (Barton Q joint set number of 9 to 12), which confirms the Jn values used for Barton geomechanical Q’ evaluation from the geomechanical core logging.

Table 16-3:  Joint Set Mean Orientation

Joint Set	Dip	Dip Direction
1	68°	248
2	74°	323
3	76°	351
4	66°	73

A wedge analysis was performed to confirm that the selected ground support stabilizes potential wedges from the roof or the walls of the ramp, as observed during site visits (Stantec, September 2015).  Four combinations of joint sets were analyzed for potential unstable wedges.  The two main azimuths of the ramp (235° and 25°) were used for the analysis.  A friction angle of 30° and no cohesion were assumed for all joint sets.

For joint set combination 1, a large roof wedge with an apex height of 40.59 m and weight of 1,692 tonnes was generated by the software.  This size of wedge is unlikely to be unstable considering horizontal clamping force.  Therefore, the apex height was reduced to 10 m (twice the ramp width).  With the combination of Swellex and shotcrete, the wedge was stable with a safety factor greater than 1 for both ramp azimuths.

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The wedge analysis allowed Stantec to establish the most favorable excavation azimuth with respect to wedge weight.  The most favorable azimuth for underground excavation in footwall andesite with respect to potential wedges formed by the mapped joint sets are 100° to 145° (280° to 325°) and 155° to 210° (335° to 30°), as illustrated in Figure 16-1.

Figure 16-1:  Most Favorable Azimuth for Underground Excavation

16.1.1.3  Modeling Parameters

Laboratory testing from the core was performed.  The laboratory results and previous point load testing results are summarized in Table 16-4.  The point load testing underestimated the strength of the rock units (except for the Los Gatos Fault Zone) when compared to the laboratory results.  Because samples sent to the lab are usually the strongest (such that they can handle transportation and preparation), it was judged appropriate to multiply the point load results for the footwall and the mineralized andesite by two to obtain strength values aligned with the lab results.  No adjustment was made for the hanging wall fault zone results since lab results and point load values are aligned.  Table 16-5 provides the adjusted rock mass properties used in the updated numerical model.

Table 16-4:  Comparison of UCS Lab Results to Point Load Testing Results

RockType	Hanging Wall Andesite (Los Gatos Fault)	Host Andesite(between ore vein)	Mineralization	FootwallAndesite	Tuff Dacite
Average UCS (MPa) Lab Results	14.70 (5 tests)	68.39 (5 tests)	74.65 (5 tests)	69.91 (3 tests)	55.79 (3 tests)
Standard Deviation UCS (MPa) Lab Results	8.98	34.79	22.65	15.21	17.82
Average Point Load UCS (MPa)	14.59	24.80	31.42	24.80	30.02
Standard Deviation Point Load UCS (MPa)	23.78	24.01	35.12	24.01	25.65

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Table 16-5:  Adjusted Rock Mass Properties used for Numerical Modeling per Lab Results

RockType	MineralizedAndesite	FWAndesite	HW Andesite(Fault)
Avg. Q’	20.7	13.60	21.71
RMR	57	54	37
GSI	57	54	37
UCS (MPa)	62.84	48.6	14.7
Hoek mi Number	25	25	25
Young Modulus Intact (GPa)	25.14	19.43	5.88
Hoek mb Number	5.383	4.836	2.635
Hoek s Number	0.0084	0.006	0.0009
Hoek a Number	0.504	0.504	0.514
Cohesion (MPa)	4.53	3.363	0.801
Friction Angle	40.55	39.65	34.5
Rock Mass Tensile Strength (MPa)	-0.098	-0.061	-0.005
Rock mass UCS (MPa)	5.67	3.689	3.045
Rock Mass Modulus (GPa)	11.37	7.52	0.76

In situ stress tensors (see Table 16-6) obtained from the Acoustic Emission Method (Villaescusa and Hogan, 2016) were used to complete the updated numerical model.

Table 16-6:  In Situ Stress Tensors

Stress TensorComponent	Value(MPa)	Orientation	Plunge
Major Principal Stress	3 + 0.0532*D	116°	9°
Intermediate Principal Stress	1 + 0.0422*D	25.5°	1°
Minor Principal Stress	0.0273*D	298°	82°

Note:  D is the depth below surface in meters.

The deposit is shallow with weak rock strength.  The rock mass is generally characterized as fair to poor. As a result, the most likely mode of failure for the rock mass is associated with gravity failure (unraveling) due to a loss of confinement, which allows blocks or wedges to move freely along existing weakness planes such as joints and faults.  The best indicator of loss of confinement is a low or negative minor principal stress.  Therefore, to evaluate the stability of the rock mass with 3D numerical modeling, the minor principal stress was used.

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16.1.2  Analysis and Design

16.1.2.1  Numerical Modeling

To achieve the optimum production mining rate, concurrent extraction of ore from three mineralized zones may be required.  In addition, sill pillars may be created within individual zones by production mining starting on two levels concurrently (Stantec, April 2016).

Numerical modeling was completed to meet the following objectives:

·Ensure that concurrent extraction of all three zones will not adversely impact regional stability of the Los Gatos deposit.

·Verify that the creation of a sill pillar within a zone is possible.

·Evaluate the impact of the Los Gatos Fault on hanging wall stability.

The modeling was performed in 2D and 3D using Rocscience’s Phase2 code and MAP3D code, respectively.

The 2D modeling was completed to increase the understanding of the impact of bulk mining in the Central Zone and, more specifically, to observe the impact on the hanging wall and the Los Gatos Fault behavior.  The 2D modeling results indicated that yielding extends along the fault beyond the excavated stopes and that hanging wall failure is expected, which is directly linked to the shallow dip of the deposit.

All 3D models were linear.  The model geometry is illustrated in Figure 16-2.  Backfill was not used in the model.

Figure 16-2:  Geometry Used in the 3D Model Looking towards the Hanging Wall

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The combination of competent footwall rock and ore with the major principal stress parallel to the strike of the ore body creates favorable mining conditions for the Cerro Los Gatos ore body.  The 3D numerical model did not indicate any significant negative impact from mining all three zones concurrently.  However, there is a small West extension to the CZ (the CZ West Extension) that may have some negative impact on the hanging wall pillar of the Northwest Zone.  In stopes of level 1410 unstable zones have been identified due to the proximity to the fault.

For all three zones, the Los Gatos Fault is affected by mining-induced stress changes, especially in the CZ, due to the shallow dip of the deposit.  The amount of influence of mine-induced stress on the Los Gatos Fault is dependent on the buffer zone between open stopes and the fault zone.  Where the buffer is thin (less than 5 m), there is more potential for failure and unraveling.  This implies a possible hanging wall instability problem that may cause excessive dilution or unraveling, especially if a bulk mining method is used.  Therefore, ore pillars are required where the fault is within 5 m of the hanging wall ore contact to mitigate potential fault failure.  Operational data has indicated that ground stability is reached with ore pillars of approximately 2 meters.  In some places the geomechanical conditions are favorable enough to mine all the ore of the vein until the contact with the fault.  Some areas where the ground is less stable, a pillar of 1.0 m is left to avoid unstable conditions during operations.

Central Zone

The 3D modeling of the Central Zone also indicates that waste pillars between veins will yield, causing potential instability problems in the hanging wall of individual veins.  The extraction sequence of the veins is important in that the veins closest to the fault should be mined first.  The likelihood of hanging wall instability problems occurring are reduced by mining with cut-and-fill methods.  This is due to the smaller excavation size and more effective installation of support in the hanging wall.  The sill pillar in the CZ is expected to yield, resulting in adverse ground conditions four to five cuts below the upper sill.  Operations in this zone have confirmed backfill deficiencies have caused unstable conditions on the hanging-wall zone, as it was experienced in stope 625, where the stability problems created blocks to cause the support to yield.  The predicted conditions in this zone have been accurate.

Northwest Zone

The 3D modeling of the NWZ indicates that the sill pillar should yield progressively as the lower mining front approaches the sill pillar.  This confirms that the primary/secondary chevron extraction sequence is sound.  Shifting of blast holes or unraveling may occur when drilling takes place in the yielded sill pillar.  Extraction of the NWZ has less impact on the Los Gatos Fault than the extraction of the CZ due to the steep dip of the Northwest Zone.  Instability issues with the hanging wall of the NWZ are expected to be less critical compared to the Central Zone.

Mining of the CZ West Extension has an impact on the Los Gatos Fault according to the numerical model.  As mining progresses in the NWZ, the pillar between the NWZ and the CZ West Extension is affected.  This may lead to stability problems and greater dilution of the hanging wall of the NWZ adjacent to the CZ West Extension.

The operating geotechnical conditions, however, have been better than the anticipated conditions.  This has allowed to excavate bigger dimensions in the stope length (more than 12 m), and the planned activities have considered excavating more in the vertical dimension of the stope (25 m height).  Some of the stability problems found in this zone are mainly due to drilling deficiencies, which have reached areas near the hanging-wall fault zone.

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Southeast Zone

The 3D modeling of the Southeast Zone did not identify any issues except for potential hanging wall failure.  Current operations have not developed any work in this zone at the time of generating this report.

16.1.2.2  Modeling Results

Stantec makes the following recommendations based on the numerical modeling results.

·All three zones may be mined concurrently.

·Bulk mining methods (Longhole Mining) may be used in the NWZ.

·A primary/secondary chevron extraction sequence for pillars is recommended for the NWZ.

·Cut-and-fill mining methods are recommended for the CZ and SEZ.

·Timely backfilling is required to minimize stability issues.

·Ore pillars are required where the fault is within 5 m of the hanging wall ore contact to mitigate potential fault failure.

In general terms, this model has had good results.  Some of the differences to this model and the actual operations are that the ground conditions are better than the estimated by the model.  Other differences to this model are more evident due to the presence of water, which has been more abundant than the predicted amount.

16.1.3  Stope Design Criteria

The geometry of the CZ, as well as the risk of water inflows from ungrouted diamond drill holes, dictates a cut-and-fill mining method.  The geometry of the NWZ, combined with the lower risk of water inflows, makes this portion of the ore body amenable to Bulk (Longhole) Mining (Stantec, December 2015).

16.1.3.1  Cut-and-Fill Mining

From the Rock Mass Ratings (RMR) (Bieniawsky, 1989) presented in Stantec’s report Geomechanical Hole Compilation and Future Data Requirements, the empirical unsupported span design curve was used to determine the span of excavation allowable for the CZ.  An RMR of 40 to 60 was used, which is one standard deviation (10) above and below the average RMR value of 50 for the mineralized andesite.  The results indicated that with ground support, a maximum drift span of 9 m is recommended.  For ore widths greater than 9 m, two or more drifts (drift-and-fill) are used to extract the ore.  When more than two drifts are planned, consideration must be given to installing additional support, such as shotcrete posts in the first drift, to support the effective span created by driving the adjacent drifts.  The cut height was fixed at 5 m.

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16.1.3.2  Bulk Mining

The Stability Graph Method (Potvin, 1988) was used to evaluate bulk stope dimensions and dilution potential (Clark and Pakalnis, 1997).  This method consists of comparing the hydraulic radius (Hr) of a stope surface (back, end wall, footwall, or hanging wall) to a stability number (N).  Table 16-7 lists the values used to determine Hr and N.

Table 16-7:  Parameters used to Establish Hr and N

Surface	Length orHeight(m)	Width orStrike Length(m)	Q’ Avg.	UCS (MPa)	InducedStress(MPa)	A	B	C	N
Stope Back	12	21	21.9	31.4	19	0.10	0.31	2	1.35
Stope End Wall A	20, 25, 30	21	21.9	31.4	15	0.10	0.51	4	1.45
Stope End Wall B	20, 25, 30	21	21.9	31.4	15	0.10	0.51	4	1.45
Stope Hanging Wall	20, 25, 30	12	9.6	24.3	15	0.10	0.20	5	0.96
Stope Footwall	20, 25, 30	12	9.6	24.3	15	0.10	0.20	5	0.96

The 20 m and 25 m high stopes plot in the Unsupported Transition Zone, and 30 m high stopes are on the limit between the Unsupported Transition Zone and the Stable with Support Zone for a stope width (along strike) of 12 m.  Back and end walls for all the stope dimensions studied fall within the Unsupported Transition Zone.

The recommended stope size is 20 m high (floor to floor) × 12 m wide (along strike) and 21 m long (hanging wall to footwall).  Hanging wall equivalent linear slough dilution is expected to be 2 m.  Additional measures could be used to minimize dilution, such as cable bolting the hanging wall, using pre-splitting blasting techniques along the hanging wall, backfilling stopes within three weeks from the first blast, and reducing the stope width (along strike).  If additional information from bulk sampling and exploration drilling proves positive, consideration may be given to increasing stope height to 25 m.

16.1.3.3  Cut-and-Fill Support

The temporary nature of the cut-and-fill stopes allows the use of mechanical bolts or PM12 Swellex with 7-gauge welded wire mesh.  Bolt length and pattern are a function of the drift span.  As mining progresses from the bottom up, a large failure zone could be generated in the back of the ore body from the hanging wall to the footwall, and considerations must be given to longer ground support length.  The length of the ground support required beyond the third cut has been established through a monitoring program of the stope back with either ground movement monitors and/or extensometers installed in the back central drifts of a given cut.  The minimum length of the instrument must be at least twice the length of the initial 2.4 m long ground support.  Instrumentation must be installed on every cut.

Proposed support for cut-and-fill stopes is as follows (Stantec, December 2015).

·7–9 m span:  2.4 m long mechanical bolts on a 1.0 m × 1.3 m staggered pattern or

·2.4 m PM12 Swellex on a 1.0 m × 1.0 m staggered pattern with 7-gauge welded wire mesh.

·5–7 m span:  1.8 m long mechanical bolts on a 1.0 m × 1.4 m staggered pattern or 1.8 m PM12 Swellex on a 1.0 m × 1.1 m staggered pattern with 7-gauge welded wire mesh.

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·4–5 m span:  1.5 m long mechanical bolts on a 1.0 m × 1.6 m staggered pattern or

·1.5 m PM12 Swellex on a 1.0 m × 1.6 m staggered pattern with 7-gauge welded wire mesh.

16.1.3.4  Bulk Mining Support

Top sills and bottom sills for Bulk Mining operations are subjected to blast damage and large mine-induced stress change.  Therefore, the ground support must be able to sustain the additional demand caused by the Bulk Mining operation.  Furthermore, the sills are usually open for more than 6 months, especially with a bottom- up sequence where the top sill is also used as a bottom sill.  A stiff support is usually recommended, such as fully grouted rebar (Swellex could be used in weak rock mass) and/or fully grouted cable bolts for wide spans.  Additional support consideration must be given to the brow of a stope where mucking takes place.

Proposed support for Bulk Mining stopes is as follows (Stantec, December 2015):

·5-6 m span:  1.8 m long fully grouted rebar on a 1.0 m × 1.5 m staggered pattern or 2.4 m PM12 Swellex on a 1.0 m × 1.5 m staggered pattern with 6-gauge welded wire mesh.

·6-8 m:  2.4 m long fully grouted rebar or 3 m long PM12 Swellex on the same pattern as described above for the 5–6 m wide span.  Any span greater than 8 m requires longer support, depending on the effective span of the excavation.

A 5–6 cm thick coat of shotcrete may be used in some instances to minimize damage to existing support, especially above the open stope.  Brow support may require a 6–8 cm coat of shotcrete or shotcrete arches.

16.1.4  Development

From the numerical modeling results, all major infrastructure and main access should be at a standoff distance of 20 m from the ore body footwall contact to minimize issues due to mine-induced stress.

16.1.4.1  Lateral Development Support

Empirical and deterministic design were used to establish ground support guidelines for footwall waste development (Stantec, December 2015).  The deterministic analysis was based on a dead weight approach. It is anticipated that due to the shallow depth of the proposed mine, the main mode of failure will be gravity failure where confinement is lost (relaxed zone around an excavation).

Two-dimensional numerical modeling was completed for a range of drift spans to establish the depth of failure to be used in the dead weight analysis.  Rocscience Phase2 software was used for this analysis.  The expected depth of failure was used to determine required bolt lengths considering 30 cm of anchorage for grouted rebar and 110 cm for Swellex.  The final selected pattern is a function of the required safety factor for the excavation of 1.5 for a permanent excavation.

All bolting must be completed through 6-gauge, galvanized, welded wire mesh with a 10 cm × 10 cm square pattern or through fiber-reinforced shotcrete.  The mesh must cover the roof, shoulder, and walls down to a minimum of 2 m from the floor.  Where shotcrete is required, shotcrete must cover the entire excavation down to the floor.

For the walls, 1.8 m long tendons (galvanized friction sets, grouted rebar, or Swellex) may be used.  If Swellex bolts are used, they must be coated to provide protection against corrosion.

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The support classes shown in Table 16-8 were established for the following four geomechanical domains:

·Support Class 1 for Domain 1:  Q > 10 (indicating good rock)

·Support Class 2 for Domain 2:  4 < Q > 10 (indicating fair rock)

·Support Class 3 for Domain 3:  0.1 < Q < 4 (indicating very poor and poor rock)

·Support Class 4 for Domain 4:  Q < 0.1 (indicating extremely poor rock)

Table 16-8:  Ground Support Class Description for 5 m Wide × 5.5 m High Drift

	Back Support		Wall Support
Support Class	20 mm Rebar2 m long	PM16 Swellex2.4 m long(1)	35 mm Friction Stabilizer1.8 m long(2)	SurfaceSupport
1	1.2 m × 1.4 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern) down to within 1 m of the floor	No. 6 gauge, galvanized, welded wire mesh with 10 cm × 10 cm squares within 1 m of the floor or 5 cm of fiber- reinforced shotcrete down to the floor
2	1.2 m × 1.4 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern) down to within 1 m of the floor	5 cm of fiber-reinforced shotcrete down to the floor
3	1.2 m × 1.4 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern) down to within 1 m of the floor	7 cm of fiber-reinforced shotcrete down to the floor
4	1.2 m × 1.4 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern)	1.2 m × 1.5 m staggered pattern (dice pattern) down to within 1 m of the floor	10 cm of fiber-reinforced shotcrete down to the floor

(1)The minimum Swellex required are PM16. PM24 may be substituted for PM16 since it has a greater capacity than PM16.

(2)The minimum requirement for the walls is 35 mm friction stabilizers; the same bolts used for the back may be substituted for friction stabilizers since they have a greater capacity.

Table 16-9 lists the requirements for wider drift spans.  Wedge analysis indicates that the current support system used in for footwall development is adequate to stabilize wedges (Stantec, July 2016).

Table 16-9:  Support System for Spans Greater than 5 m

	Bolt Length (m)			Bolting Pattern (m)
Span(m)	Rebar	Swellex	Cable Bolts	Rebar	PM16	PM24	CableBolts
5.1 to 6.5	2.4	3.2	4.5	1.2 m × 1.2 m pattern	1.2 m × 1.5 m pattern	1.6m × 1.6 m pattern	1.6m × 1.6 m pattern
6.6 to 7.3	3.0	3.2	4.5	1.1 m × 1.2 m pattern	1.2 m × 1.3 m pattern	1.5 m × 1.5 m pattern	1.5 m × 1.5 m pattern
7.4 to 9.1	3.3	4.1	4.5	1.0 m × 1.0 m pattern	1.0 m × 1.1 m pattern	1.3m × 1.3 m pattern	1.5 m × 1.5 m pattern

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16.1.4.2  Vertical Development

The vertical development consists of ventilation raises with raise bore machines.  At this stage of the design, ore passes are not planned.  Four ventilation raises have been completed at the time of writing this report.

To evaluate the potential stability of raise bored openings, an empirical method developed by McCracken and Stacey (1989) was used.  Table 16-10 provides the Q values used for selected Los Gatos rock unit that may be bored from underground through surface.  Figure 16-3 and Figure 16-4 illustrate different raise diameters and wall and face stability with respect to the selected rock unit.  As shown, depending on the Q value encountered, raises may or may not be stable.  Referring to the first 100 m of the decline below surface, a portion of raises near surface require support such as 5-7 cm of fiber-reinforced shotcrete.  Therefore, it is recommended to drill a pilot hole at selected raise location to evaluate the rock mass quality as a function of depth to better plan requirements for support (if any).

Table 16-10:  Q Value for Los Gatos Selected Rock Unit

Rock	Q Value
Type	Minimum	Average
Andesite	0.007	15.69
Rhyolite	0.040	27.55
Tuff	0.042	12.83

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Figure 16-3:  Raise Wall Stability

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Figure 16-4:  Raise Face Stability

16.2  Mine Access Design

16.2.1  Mine Access

The upper portion of the CZ (the bulk sample area on the 1400 Level) is currently accessible via the existing portal and main access decline.  Three internal ramps will be developed throughout the mine life, one for the NWZ, one for the CZ, and one for the SEZ and SEZ2.

16.2.2  Access Ramps

The main access decline connects the main surface portal to the Northwest and Central internal ramps at the 1420 Level near the bulk sample location.  Where possible, ramps are designed to maximize straight runs for safety and haulage efficiency, as well as to minimize wear on mobile equipment.

The Southeast ramp, which will service the SEZ and SEZ2, is an extension of the Central ramp and will not be required until midway through the mine life.

For ease of entry and exit, ramps have been designed at a 15% gradient, reducing to 13% in the turns, and leveling out to 4% at main intersections.  Curves have been designed with a minimum 25 m radius, passing bays are incorporated where required, and safety bays are included at 30 m spacing.

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16.2.3  Mining Sublevels

Sublevels in the NWZ are accessed from the NWZ access ramp on 20 m vertical intervals, which is the planned longhole stoping height.  For geomechanical stability, footwall drifts are set back a minimum of 20 m from ore contact.  The level entrance from the ramp are set back approximately 30 m from ore contact.  Stope access cross-cuts are planned for each individual transverse or longitudinal longhole stoping block.

Drift-and-fill stoping will be used in the CZ and SEZs.  Sublevels in these areas are accessed from the access ramps on 25 m vertical intervals, which is the cut-and fill stope height based on accessing five 5 m cuts from each sublevel.  Stope access cross-cuts (attack ramps) in the drift-and-fill areas are developed on a 200 m horizontal spacing along strike.

Level development follows the general strike of the various ore zones and typically includes excavations for sumps, electrical substations, remuck bays, paste fill lines, and raise accesses.

Sublevels in both the NWZ and CZ generally terminate at ventilation raises at both ends, one fresh air and one exhaust, permitting the intake of fresh air and the exhaust of contaminated air on each level.

Figure 16-5 illustrates typical level development requirements in the NWZ longhole area of the mine.

Figure 16-5:  Typical Level Development Northwest Zone

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Figure 16-6 illustrates typical level development requirements in the drift-and-fill areas of the CZ and SEZs.

Figure 16-6:  Typical Level Development CZ and SEZs

Development drilling is completed using electric/hydraulic jumbos.  The holes are loaded with an ANFO-based emulsion and the round is initiated with NONEL blasting caps.

Figure 16-7 presents two examples of drill patterns for typical development headings at Los Gatos.

Figure 16-7:  Typical Waste Development Drill Patterns

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16.2.4  Ventilation Raises

The following four raise bored ventilation raises were planned between the underground and surface and have been completed:

·North — 4 m Diameter Return Air Raise

·Central — 4 m Diameter Fresh Air Raise

·South — 4 m Diameter Fresh Air Raise

·Southeast — 4 m Diameter Return Air Raise

Two additional raises have been planned and include:

·North-a second raise is planned to start in the second half of 2020

·Southeast- a second raise is planned to be constructed in this zone midway through the mine life

The ventilation raises are lined with shotcrete.

During development of the mine ramps, a series of 2.5 m × 2.5 m drop raises are excavated between each level access with ramp advance.  These raises serve as exhaust air raises during ramp development.  The drop raises in the NWZ and CZ are being used for secondary egress, while the drop raises in the SEZ and SEZ2 will be required as exhaust air raises during the production phase of these mining zones.

16.2.5  Secondary Egress

Four raise bored egress raises have been completed between the underground and surface.

The egress raises do not require a shotcrete liner due to their small diameter and lack of installed services.  The raises will provide an escape route for evacuating personnel out of the mine in the event the main access decline from surface becomes blocked.  A man cage (or “bullet”) will be lowered from surface into the mine through the raise using a truck-mounted hoist.  The ventilation department is evaluating the option of adding two more Secondary Egress raises along the SEZ and SEZ2.

Drop raises located on all ramps are equipped with ladders to provide secondary egress in case the ramp becomes blocked.

16.3  Mining Methods and Sequence

The mining methods at Los Gatos includes two bulk stoping methods, transverse longhole and longitudinal longhole, as well as a more selective drift-and-fill method.  Based on geomechanical recommendations, the sublevel intervals in the longhole stopes are fixed at 20 m.  The drift-and-fill method is used for the CZ and will be used for the SEZs, while the NWZ is mined using transverse longhole and longitudinal longhole.

16.3.1  Drift-and-Fill Mining

Drift-and-fill mining is a flexible mining method that allows near-complete recovery of ore.  Mining of ore is completed with the same equipment used for mine development, and dilution from waste external to the ore zone is minimal.  Negatively, productivity is lower than longhole stoping due to the small blast sizes

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and sequencing of backfill.  Good control of drilling and blasting is also necessary to minimize backfill dilution.

Drift-and-fill mining uses a two-boom jumbo for face drilling.  Main haulage drifts are developed laterally in the footwall, on 25 m vertical centers, and offset approximately 100 m from the vein.  The veins are accessed from cross-cuts at approximately 200 m centers horizontally along strike (see Figure 16-8).  From the cross-cuts, attack ramps are driven initially at a -15% gradient to intercept the vein(s) at the bottom floor of that stope block.  Subsequent floors are mined in 5 m cuts, with a total of five cuts being taken from each attack ramp.  Once the vein is intersected, headings are driven in both directions to provide two working faces on each vein.  In the CZ, this is repeated on two active main levels to achieve the daily production rate.

Drift-and-fill mining typically consists of parallel drifts mined in a primary-secondary sequence across the width of the vein or deposit.  Once the first (primary) drift is mined, it is tight-filled with cemented backfill (paste or cemented rock fill).  This allows another drift (secondary) to be developed immediately adjacent to the backfilled drift.  Successive cuts are then mined with the backfill from the lower cut serving as the floor for the next 5 m cut. See Figure 16-8 for an example of drift-and-fill mining.

Figure 16-8:  Typical Drift-and-Fill Stope Schematic

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Figure 16-9 presents two examples of typical ore development (drift-and-fill) drill patterns.  The pattern on the left is a full-drift pattern and could range from 4 m to 6 m in width.  The pattern on the right is for “breasting,” where the rock below has been removed and subsequently backfilled, leaving a narrow void between the cut and the floor below.  “Burn cut” and lifter holes are not required when breasting.

Figure 16-9 :  Typical Ore Development Drill Patterns

After blasting, the broken ore is mucked from the face using a 7.5 m3 LHD to the remuck bays or directly to trucks for haulage to surface.  Once mucking is complete, the back and walls of the headings are scaled and suitable ground support is installed.  After the drift is mined to the predefined boundary of the stope, it is backfilled with cemented paste fill.

Average advance for the development headings is estimated at 3.85 m/d for single headings, 5.01 m/d for double headings, and 6.16 m/d for multiple headings.

Drift-and-fill is the selected mining method for the CZ and SEZs for the following reasons:

·Drift-and-fill is a “person-entry” mining system where personnel have full access to the excavation in case the drift intersects a water-bearing drill hole or structure (therefore, water control measures can be implemented).  This is not the case with a longhole stope.

·The geometry of the veins, specifically the relatively shallow dip, is more amenable to drift-and-fill than longhole stoping.

·The poor quality of the hanging wall in close proximity to the ore source inevitably results in higher dilution with longhole stoping (compared to drift-and-fill).

·Drift-and-fill is a selective mining method, allowing close geological control of mining.  This control typically results in higher ore grades and less dilution.

·Less definition drilling is typically required with a method that essentially follows the vein.

·Mining multiple veins would result in sufficient production from drift-and-fill stoping to meet ore tonnage requirements.

·Specialty equipment is not required, and training/skill requirements are reduced.

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16.3.2  Transverse Longhole Stoping

Transverse longhole stoping is a bulk mining method in which the long axis of the stope and access drifts are perpendicular to the strike of the mineralized zone.

Access levels are developed subparallel to the strike of the ore zones in the footwall of the deposit.  Access drifts for drilling and mucking are then developed from the levels to the top and bottom of each stope.

The mine design assumes a stope width of 12 m and height of 20 m for all transverse longhole stopes. Stope lengths vary depending on the thickness of the ore zone; however, the maximum stope length is restricted to 21 m by the hydraulic radius, while the minimum length has been determined to be 9 m.  (longhole stoping areas 9 m or less in length will be mined longitudinally).  The drilling level at the top of the stope and the mucking level at the stope bottom are driven 5 m wide × 5 m high with arched backs.  Stopes have been designed with a minimum hanging wall and footwall angle of 55°.

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Figure 16-10 illustrates the concept for a typical development and extraction sequence for a transverse stoping block, while Figure 16-11 shows the typical transverse longhole stope dimensions.

Figure 16-10:  Typical Transverse Stope Schematic View

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Figure 16-11:  Transverse Stope Dimensions and Drilling

Transverse longhole stoping is considered a bottom-up method whereby the lowest stopes are removed first.  The stope sequence is primary-secondary.  In primary stopes, the drill drift and mucking drift is driven perpendicular to the strike of the ore body.  Access to the secondary stopes consists of a “Y” drift developed from the primary access drift.

After completion of the development, longhole drilling is started.  All drilling is completed before blasting starts.  To start stope extraction, a slot is created at the far end of the stope.  This slot creates a void for the ensuing stope blasts.

To begin the slot sequence, a drop raise is required.  The raise is approximately 2.5 m high × 2.5 m wide × 15 m long (see Figure 16-12).  The drop raise is blasted in two lifts, with the first being half of the length of the drilled raise.

Figure 16-12:  Typical Slot Raise

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The drop raise is then expanded to form the slot (see Figure 16-13).  The first three stope rings are blasted into the drop raise to form a slot the width of the stope.  The slot provides a large enough void to begin stope production ring blasting, which is typically completed in increments of two to three over the length of the stope.

Figure 16-13:  Slot Ring Drilling

The transverse longhole stope mucking drift is slashed to the full 12 m stope design width immediately in advance of each stope blast.  Walls are slashed using a jumbo drill (see Figure 16-14).  After the slot is complete, the slashing of the mucking drift starts and retreats ahead of the production blasting (see Figure 16-15).  Once the slash is taken, personnel are not permitted in the area since it does not include entry-quality ground support.

Figure 16-14:  Production Ring and Slash Drilling

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Figure 16-15:  Slot Complete and Undercut Started

The remainder of the stope is drilled and blasted on retreat from the initial slot rings to the stope entrance (see Figure 16-16).

Figure 16-16:  Production Blasting Retreat

Stopes above the bottom stopes only require a drill drift since the mucking drift is now the drill drift from the stope below.

Secondary transverse longhole stopes are mined similar to primary stopes.  However, secondary stopes require the primary stopes on both sides be mined out and backfilled (and the backfill-cured) prior to blasting.

After the stope is mined, it is backfilled with paste except when waste rock is available for inclusion in the secondary stope backfill mix.  Until paste fill is available, cemented rock fill (CRF) is used for backfill.

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During the backfill portion of the stope cycle, a bulkhead is initially constructed at the stope entrance of the mucking drift.  Paste fill is then poured from the drill drift into the open stope (see Figure 16-17).  Once the paste filling is complete, it requires a cure time of 28 days prior to blasting adjacent stopes.  Other activities such as development and drilling in the adjacent stopes may continue during the backfilling and curing cycle.

Figure 16-17:  Typical Paste Backfilling

In general, transverse longhole stoping is used where the rock mass quality of the hanging wall limits the length of the open mining span.  This method requires more footwall waste development (footwall drifting, cross-cuts and drawpoints); however, since each stope has an independent access it has more flexibility with regards to sequencing and scheduling.

16.3.3  Longitudinal Longhole Stoping

Longitudinal longhole stopes are up to 9 m wide (from footwall to hanging wall).  In the NWZ, ore zones (or veins) that are wider than 9 m are mined using transverse longhole stoping.

Similar to transverse stoping, longitudinal stoping consists of a drill drift at the top of the stope and a mucking drift at the bottom.  These drifts are driven in the ore parallel to the strike of the ore body and developed 5 m wide.

This method requires the ore drifts be driven to either a defined boundary, a change in mining method (transverse), or the end of the mineralization.  Then, the stopes are mined in sequence from the end of the drift to the access (i.e., retreat mining).  Similar in process to transverse longhole stoping, each stope has a drop raise as a beginning point for the extraction retreat to the access crosscut.  In accordance with geomechanical design parameters, should the stope length be greater than 12 m along strike, it is necessary to stop, backfill, and re-establish a new drop raise before restarting the stoping sequence (see Figure 16-18).  Upon completion, all voids are typically paste filled and supplemented with development waste rock when applicable.

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Figure 16-18:  Longitudinal Stope Dimensions and Drilling

16.3.4  Backfilling

A CRF plant has been constructed during the pre-production period to ensure a source of cemented backfill for primary stopes during pre-production (mining ore to build up a stockpile) and early in the production period (until the paste plant is built).  The underground haulage trucks are equipped with ejector-style boxes and backhaul the CRF material from the surface plant to the underground stopes.  The paste plant is currently scheduled to be constructed in June 2021.

Secondary stopes are less demanding for structural backfill.  A combination of lower-strength (lower binder content) paste and development waste is used in secondary stopes.  When secondary stopes are available, it is intended that most or all waste development rock is used for backfilling to avoid hauling waste to surface.

16.3.5  Transverse Longhole Stopes

All primary transverse longhole stopes are backfilled with CRF until the paste plant is operational.  Dump blocks are required when trucks are dumping CRF into longhole stopes and must be advanced as the stope is filled.

A barricade is required on the lower access drift when backfilling.  Typically, a waste muck pile is adequate for CRF.  In the case of paste, it may be necessary to construct a shotcrete bulkhead.

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Paste fill is transferred into the stope by extending the steel delivery line with HDPE pipe to the edge of the stope.  As the pour continues the pipe can be advanced if required.

As an initial paste pour, a “plug” of paste deep enough to cover the lower entrance to the stope is poured and allowed to cure for two or three days.  Following that, the stope is filled and allowed to cure up to 28 days before mining occurs above or immediately adjacent to the filled stope.

16.3.6  Longitudinal Stopes

Filling of longitudinal stopes is similar to filling transverse stopes, but because there is no primary-secondary sequence, the fill requirements differ slightly.

When using CRF, only a portion the stope must be filled with a cemented product.  Once the stope is approximately 60% filled, the remaining unfilled back end of the stope can be filled with unconsolidated waste.  All that is necessary is to have a solid backfill face in place to blast against when re-slotting for the next stope.  The last fill in the stope can be 100% rock fill or low-strength paste.

When pouring paste in longitudinal stopes, the procedure is the same as that in transverse stopes.

16.3.7  Drift-and-Fill Stopes

CRF has been used initially in the primary drift-and-fill stopes.  Although the ejector-equipped trucks can dump in the stope without raising the box, it is necessary to have an LHD available (preferably equipped with a “rammer jammer”) to push up the CRF to ensure tight filling.

Paste filling in the drift-and-fill stopes is a relatively easy procedure.  An HDPE paste line is extended to the stope face, and a waste rock barricade is built at the entrance to the stope cut.  As the paste is poured, the line is intentionally broken, or pulled back, to ensure tight fill for the length of the stope.

Secondary drift-and-fill stopes can be filled with development waste and/or low-strength paste.

16.3.8  Production Rate

The production rate is 2,500 tpd from the second quarter of 2021 forward.  Year 2020 will include a reduced production rate, due to shut down through the global pandemic in spring 2020, with full production being reached at the end of the first quarter of 2021.

The Los Gatos Resource is split between the CZ (56.2%), NWZ (31.3%), and SEZs (12.5%), each of which are planned to be operated somewhat independently.

Mining of the SEZs is scheduled to start in 2025.  These zones typically have lower NSR values and consist of comparatively narrow veins and lower tonnes per vertical meter than the CZ and NWZ.  As a result, the SEZs are scheduled later in the mine life (starting in 2025) and are meant to replace ore production from the CZ as production in that zone drops off.

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16.3.9  Stope Design Parameters

Table 16-11 presents the stope design parameters used for the project.  Figure 16-19 presents the design span curves used for the drift-and-fill mining.

Table 16-11:  Stope Design Parameters

Parameter	Unit	Drift-and-Fill	Longitudinal Longhole	Transverse Longhole
Maximum Mining Width	m	9 m	9 m	12 m
Drift Height	m	5 m	5 m	5 m
Maximum Stope Length	m	200 m	12 m	21 m
Maximum Strike Length	m	n/a	12 m	12 m
Stope Height	m	5 m	20 m	20 m
Access Drift Gradients	%	-15% to +15%	flat	flat
NSR Cutoff Value	$US	$170	$170	$170
Dilution	%	7.1	28.5	17.3
Recovery	%	98	96	96

Figure 16-19:  Design Span for Drift-and-Fill

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16.4  Development and Production Schedules

16.4.1  Development Productivity Rates

Development crews drive multiple headings whenever possible, thus increasing utilization of crews and equipment.  Double-heading estimates are prepared based on a 30% increase over the single-heading rate for the same drift size.

Vertical development, such as internal vent raises and manway raises, uses several techniques (including drop raises, raise bores, and raise climbers) to achieve the most cost-effective development.  Vertical advance rates include ground support/lining and mobilization/demobilization of equipment and crews. As mining progresses, detailed information from the mine should be included into future planning.

16.4.2  Development Schedule

The LOM development schedule is presented in Table 16-12.  It is recommended that the schedule continues to be updated with the most current Resource and an updated mine plan as the mine life progresses.

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Table 16-12:  LOM Development Schedule

Waste Type	2H 2020	2021	2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	TOTAL
Lateral WasteMeters	2,145	4,680	5,100	5,100	4,599	2,393	654	828	556	411	817	383	27,666
Vertical WasteMeters	728	467	350	103	580	468							2,697
Total WasteMeters	2,873	5,147	5,450	5,203	5,180	2,861	654	828	556	411	817	383	30,363

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16.4.3  Production Planning Criteria

The Mineral Reserve estimate includes detail by mining method for each ore zone.  These detailed estimates were used to determine production sequencing.  The following general planning criteria were applied to determine priorities for initial production.

·Highest Grade

·Highest Productivity

·Lowest Mining Cost

Productivities were generated and added to the Deswik development and production schedule.  Stope productivities are based on a typical total stope cycle, including cable bolting, drop raising, long hole drilling, production blasting, remote mucking, fill fencing, backfilling, and delay for curing.  Table 16-13 lists the productivity rate for each mining method.

Table 16-13:  Productivity Rates

MiningMethod	Finished Size(m)	Productivity Rate (tonnes/day/stope)
Drift-and-Fill	6× 5	363.3
Longhole — Transverse	12 W x 14 L x 20 H	191.5
Longhole — Longitudinal	9 W x 12 L x 20 H	127.5

The LOM schedule has been generated using a series of software suites.  Stope shapes created with Vulcan and MSO were used as a starting point to design the underground workings.  Additionally, Vulcan was used to assign attributes to the design strings and solids that were carried forward into the scheduling suite (Deswik).

16.4.4  Production Schedule

The LOM production schedule is presented in Table 16-14.  As additional information is gathered from mining and exploration programs, the production schedule should be continuously updated throughout the mine life.  The schedule starts in July 2020.

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Table 16-14:  LOM Production Schedule

Item	2H 2020	2021	2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	TOTAL
Total Ore Tonnes	320,990	862,000	900,000	900,000	900,000	900,000	900,000	900,000	900,000	900,000	900,000	334,642	9,617,631
Diluted NSR ($)	201	218	244	227	231	195	180	158	170	182	159	151	195
Diluted Ag (g/t)	331	368	423	384	374	298	266	214	239	274	240	221	305
Diluted Pb (%)	2.23	2.70	2.97	2.69	2.86	2.85	3.02	2.97	2.77	2.79	2.41	2.12	2.76
Diluted Zn (%)	4.27	5.46	5.73	5.86	6.72	6.17	5.66	5.41	6.00	5.76	4.53	4.92	5.65
Diluted Au (g/t)	0.50	0.41	0.39	0.39	0.38	0.30	0.30	0.30	0.29	0.30	0.40	0.38	0.35

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16.5  Mine Equipment

16.5.1  Underground Mobile Equipment

Mobile equipment required for lateral development includes drill jumbos, LHDs, haul trucks, and ground support equipment.

Operating mobile equipment requirements are estimated for the projected LOM production and development schedules.

The mobile equipment required has been determined by the following criteria.

·      Where a fleet of equipment (e.g., jumbo drills, underground loaders, haul trucks) is required, additional units have been purchased to provide standby coverage associated with expected availability.  Mechanical availability of 80% was assumed for all mobile equipment.  For utility vehicles, where one unit or less is required, one unit has been purchased.

·      Geometry of the ore body may sometimes increase equipment requirements.  In this case, the Cerro Los Gatos deposit is more than 2,000 m along strike, which can place unreasonable demands on slower-moving equipment such as jumbo drills.

Table 16-15 provides a list of the underground equipment with maximum units required on site for the LOM.  The quantities exclude replacement equipment.  The overall operating quantities fluctuate over the LOM to match the production and development schedule requirements at any given time.

Table 16-15:  Underground Mobile Equipment — Maximum Units on Site

Mobile Equipment	Max Units
Electric/Hydraulic (E/H) Jumbo Drill Rig	6
E/H Bolter Drill Rig	5
E/H Production Longhole Drill	3
Air Utility Longhole Drill	1
5.7 m3 LHD	6
3.1 m3 LHD	2
40-tonne Capacity Haul Truck	7
Grader	1
Emulsion Charger	3
Boom Truck	4
Fuel and Lube Truck	1
Scissor Lift	2
Personnel Carrier	2
Shotcrete Sprayer	2
Transmixer	3
Telehandler	3
Total	50

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16.5.1.1  Electric/Hydraulic Jumbo Drill Rigs

During steady-state operations, development advance in waste and ore (including the drift-and-fill stopes) is expected to average approximately 9,000 m/yr.  Based on estimated advance rates, as well as the potential travel requirements with a Resource more than 2 km long, it was determined that three operating units plus one spare are required.

16.5.1.2  Electric/Hydraulic Bolter Drill Rig

Based on geomechanical recommendations, pattern bolting, and occasional screening is required in development headings and in the drift-and-fill areas.  Bolting mainly consists of 2.4 m long Swellex-type friction bolts. Bolters should also be equipped for the installation of rebar (or DYWIDAG) bolts.  At steady-state production, four operating bolters plus one spare are required.

16.5.1.3  Electric/Hydraulic Production Longhole Drill

Production drilling in the longhole stopes is completed with appropriately sized, rubber-tire-mounted, top-hammer drill rigs.  One operating plus one spare machine are required.

16.5.1.4  Air Utility Longhole Drill

Drilling of service and drain holes, and possibly slot raises, are completed using a small, air-operated, top-hammer longhole drill rig.

16.5.1.5  Diesel 14-tonne LHDs

Based on development and production requirements in the schedule, four 14-tonne LHDs plus one spare are required.

16.5.1.6  Diesel Haulage Trucks

Truck haulage to surface is required for the scheduled ore production rate of 2,500 tpd plus a portion of the waste from development.  It has been assumed that, whenever possible, waste is delivered to secondary stopes underground for disposal as backfill.

Seven operating 40-tonne capacity trucks, plus one spare, are required during steady-state operations.  The trucks are equipped with ejector boxes for transport of CRF from surface as well as moving development waste for backfill into drift-and-fill secondary stopes.

16.5.1.7  Graders

One low-profile grader is required to maintain underground roadways.

16.5.1.8  Explosives (Emulsion) Chargers

Two emulsion chargers (plus one spare) are required for loading the waste development headings, longhole stopes, and drift-and-fill ore rounds.  There is a single explosives magazine on site, approximately 1 km from the portal, and no other magazines are expected to be permitted.  Emulsion is picked up at the magazine and brought underground, as needed, by the same vehicles used for charging the blastholes at the working faces.  The chargers are equipped with a boom and man basket for loading horizontal blastholes of mining faces up to 6 m high.  The base emulsion and sensitizer are combined at the face

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while being pumped into each blasthole.  Detonators and ancillary explosives are transported to the working faces using explosives utility vehicles.  Any unused explosives, ancillary explosives, or detonators must be returned to the magazine at the end of each shift.

16.5.1.9  Boom Trucks

Essentially a flat deck with a crane, boom trucks are used to delivery heavy loads (e.g., pipe, shotcrete bags, rock bolts, or lumber) as well as for construction and maintenance activities.

16.5.1.10  Fuel and Lube Truck

A fuel and lube truck is required to fuel and service all equipment unable to return to the SatStat fueling areas at frequent intervals.  This equipment typically includes jumbos, bolters, and longhole drills.

16.5.1.11  Scissor Lifts

Scissor lifts are required to install pipe, ventilation ducting, hanging fans, power cables, and other service activities.  With an average development advance of seven rounds per day throughout the production period, two scissor lift trucks are required.

16.5.1.12  Personnel Carriers

Personnel carriers are required for the approximately 50 personnel traveling underground per shift. Some personnel typically travel underground in tractors or light vehicles.  The listed carrier can accommodate up to 28 passengers; therefore, two units are required.

16.5.1.13  Shotcrete Sprayer

Based on geomechanical recommendations, it is estimated that 20% of all lateral development, 60% of ramp development, and all infrastructure excavations require shotcrete.  In addition, shotcrete is required for miscellaneous construction items such as backfill barricades.  Two units provide adequate capacity and redundancy for this critical activity.

16.5.1.14  Transmixers

Shotcrete is mixed in the existing surface batch plant and delivered underground using transmixers.  Two operating plus one spare unit have been included in the estimate.

16.5.1.15  Telescopic Handler (Telehandler)

A telehandler is essentially a forklift with a telescopic boom that can extend forward and upward from the vehicle.  These machines are the primary movers of supplies underground.  Three units have been included in the estimate.

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16.5.2  Surface Mobile Equipment

Table 16-16 provides a list of surface equipment with maximum units required on site for the LOM.  The quantities exclude replacement equipment.

Table 16-16:  Surface Mobile Equipment — Maximum Units on Site

Mobile Equipment	Max Units
Utility Vehicle/Pickup/Tractor	8	*
Surface Loader	1
Dozer	1
HDPE Pipe Fusion Machine	1
Total	11

* Included in surface equipment but suitable for underground use.

16.5.2.1  Utility Vehicles

Tractors are used to transport some personnel during shift changes and for nipping materials and general transport throughout the mine.  All tractors are equipped with a cargo/man—carrying compartment in the back.  Some tractors are also equipped with man lifts for facilitating services installations, constructing bulkhead, surveying, geological mapping, loading of development rounds, etc.

Utility vehicles are used by personnel for quick transport between headings and are the preferred mode of transport for supervision and technical support staff.

The following crews have been issued tractors and/or utility vehicles for use during the shifts.

·Development blasters

·Backfill crew

·Mechanics

·Electricians

·Production blasters

·Warehouse

·Managers/shifters and technical support staff

16.5.2.2  Surface Loader

The surface loader is used to load crushed waste rock into the CRF plant aggregate feed hopper, move surface skid-mounted equipment, and clean spills on surface roads.

16.5.2.3  Excavator

A mini excavator is used for excavating trenches for mechanical and electrical services.

16.5.2.4  Forklift

The forklift is required to transfer materials on surface between the surface facilities and to load materials onto the utility vehicles for transport into the mine.

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16.5.2.5  Scissor Lift

A scissor lift is used for servicing equipment located above the reach of maintenance personnel.  Examples of this are overhead crane, hose reels, interior lights in the maintenance facility and warehouse, and yard lights.

16.5.3  Fixed Equipment

A list of major fixed equipment by category is presented below.  Detailed scope for each facility is outlined in Section 18.0 .

16.5.3.1  Surface Fixed Equipment

Surface infrastructure supporting the underground mine and associated equipment are as follows:

·Cemented Rock Fill Plant

·Aggregate feed hopper

·Aggregate feed hopper discharge conveyor

·Aggregate transfer conveyor

·Cement silo

·Twin shaft continuous mixer

·Cement silo discharge screw conveyor

·Cement weigh hopper

·Cement weigh hopper discharge screw conveyor

·Super wash pump

·Super wash oscillator

·Plant water feed pump

·Air compressor

·Aggregate feed hopper vibrator

·Cement silo dust collector

·Super wash water tank

·Plant water tank

·Plant air receiver

·Instrument air dryer

·Admix system and pump

·Building dust collection system

·Dust collection screw conveyor

·Cement silo rotary air lock valve

·Dust collection rotary air lock valve

·Overhead bridge crane

·Cemented paste fill plant

·Air compressor

·Plant air receiver

·Instrument air dryer

·Instrument air receiver

·Tailings surge tank

·Tailings surge tank mixer

·Tailings surge tank pump

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·Vacuum disc filter

·Vacuum disc filter discharge conveyor

·Paste mixer (pug mixer)

·Paste hopper

·Paste pump

·Flush pump

·Cement silo

·Cement silo rotary valve

·Cement silo weight belt feeder

·Cement silo discharge screw conveyor

·Fresh water tank

·Fresh water tank pump

·Waste tank

·Waste tank pump

·Floor sump pump

·Vacuum pump

·Filtrate receiver

·Filtrate receiver pump

·Surface ventilation fans

·Inlet bell and screen

·Discharge cones

·Motor

·Discharge silencer and transition

·Fan brake

·Ultrasonic flow monitor

·Static pressure sensor

·Ductwork package (including the following)

·Expansion joints

·Transition ducting

·Ducting components

·Raise plate and elbow connection

·Refrigeration chillers and plant

·Refrigeration machine

·Cooling towers

·Water pumps and piping

·Air coolers

·Fans

·Gearboxes

·Drive shafts

·Motors

·Torque tubes

·Oil/lubrication pump and pipe

·Vibration accelerometer

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·Office/dry building

·Air handler

·Heating and cooling (HVAC)

·Hot water tank

·Water pumps

·Sprinkler distribution system

·Surface maintenance shop with wash and welding bay

·Oil pumps

·Eye wash stations

·Foam generator

·Retractable hose reel and drum pumps

·Shop roll-up door

·Tools for main shop

·25t, 10t, 5t bridge cranes

·5t monorail crane

·Separator tank / oil skimmer

·Pressure washer

·Surface fuel station

·Oil pumps

·Eye wash stations

·Fuel/oil piping

·40,000 l double-wall, skid-mounted fuel tanks

·Retractable hose reel and drum pumps

16.5.3.2  Underground Fixed Equipment

Underground infrastructure and associated equipment are as follows:

·Mine dewatering

·Dewatering pumps — primary

·Dewatering pumps — secondary

·Portable dewatering pumps

·Underground mine ventilation

·Facility and auxiliary fans

·Development and production ventilation fans

·Substation ventilation fans

·Fire doors

·Regulators

·Underground satellite shop

·Oil pumps

·Eye wash stations

·Fuel/oil piping

·Retractable hose reel and drum pumps

·Shop roll-up door

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·Tools for main shop

·Jib cranes

·Pressure washer

·Safety and miscellaneous

·Refuge chambers

·Secondary egress raise

·Electrical equipment

·Ventilation substation electrical equipment

·Dewatering substation electrical equipment

·Ventilation raise fans electrical equipment

·North, South, and Central refrigeration plants

·North, South, and Central development and production mine load centers (MLCs)

·Shop, wash bay, and fuel bay electrical

·Miscellaneous lighting and auxiliary panels

16.6  Ventilation

The purpose of the mine ventilation system is to provide air in sufficient quantity and quality to ventilate the underground workings, and most importantly, to maintain working conditions at an acceptable level and in accordance with mine regulations.

The design principles that have been incorporated for the Los Gatos project are proven and internationally recognized.  The prime objective is to provide a safe and healthy environment in all underground workings and to comply with the regulations of the Mexican standard NOM-023-STPS-2012 Underground Mines and Open-Air Mines — Occupational Health and Safety Conditions.

The Los Gatos project is a trackless underground mining operation. It is highly mechanized with LHDs, trucks, drill rigs, personnel carriers, etc.  Ore transport is via trucks to surface.  Development of the main accesses and sublevels is through drill-and-blast operations utilizing LHDs, trucks, and drill rigs.  For this reason, the ventilation system design base case is based on requirements for the mobile equipment fleet and has been adjusted for heat load calculations and required refrigeration.  As a pull system arrangement, fresh air is downcast via the main ramp and fresh air raises and exhausted through the main exhaust raises.

The mine development and production plan, along with schedules and equipment, have been reviewed to determine the airflow requirements for the various stages of the mine life.

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16.6.1  Ventilation Method and Design Criteria

The ventilation system is designed as a push system.  The two main intake raises (Central and South) provide fresh air into the mine, while the air is exhausted via the North Vent Raise, Southeast Vent Raise, and the main access decline.  Fresh air is distributed into the mine workings through a series of ramps and haulage levels.  Regulators are used to direct air to the active mining levels.  Where possible, exhaust and intake raises have been placed at the ends of the mining sublevels to provide flow-through ventilation.

Air quantities are based on a diesel emission dilution of 2.13 m3/min/hp (Mexican regulations) and additional flow for mine-cooling purposes.  The diesel engine horsepower ratings are based on technical specifications for the selected major equipment.

To address the health and safety of personnel, and as a general guideline in design, air velocities are designed to be between 0.25 m/s and 6.5 m/s.  A 20% leakage factor is also included.

The design criteria for preparing the ventilation models for Stages 1 through 3 are listed in Table 16-17.

Table 16-17:  Ventilation Design Criteria

Item	Design Value
All surface vent raises	4 m Ø
Internal vent raises	3 m Ø
Ramp dimension	5 m × 5.5 m
Footwall drift dimension	5 m × 5 m
Ventilation duct	1,220 mm Ø (48 in.)
Average drift friction factor	0.016 kg/m3
Raise bore friction factor	0.005 kg/m3
Steel duct friction factor	0.0037 kg/m3
Airflow requirement for diesel-powered units	2.13 m3/min/hp
Maximum velocity in drifts	7.6 m/s
Surface elevation mean sea level	1,550 m

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16.6.2  Airflow Requirements

16.6.2.1  Early Production

The installation of the first refrigeration plant is currently in progress, as well as the first of the permanent fan installations.  The anticipated required ventilation flowrate is ~500,000 cfm during the first phase of full production for elevations above the 1350 Level.  Two exhaust fans operating at 94.4 m3/s (200 kcfm) and 0.5 kPa each provide the required flow for this stage.  Both fans are located on surface at the Central Vent Raise and operate in a bifurcated orientation to achieve a total flow of 188.8 m3/s (400 kcfm).  See Figure 16-20 for the ventilation arrangement for this stage.

Figure 16-20:  Early Production Ventilation Arrangement

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16.6.2.2  Stage 3:  Full Production-LOM

The full production ventilation stage was modeled with active production crews in each mining area.  Fresh air is delivered to these areas via the Central Vent Raise with two surface fans and the South Vent Raise, also with two fans.  Air is exhausted via the portal/decline, Northwest Vent Raise, and Southeast Vent Raise.  The two fans on the Central Vent Raise operate in parallel with an operating point of 94.4 m3/s (200 kcfm) and 0.9 kPa per fan.  The surface fans on the South Vent Raise supply air to the central and Southern portion of the mine and have an operating point of 141.6 m3/s (300 kcfm) and 2.5 kPa each.  See Figure 16-21 for the ventilation arrangement for this stage.  The ventilation requirements are based on the diesel-powered equipment fleet for both production and development mobile equipment.

Figure 16-21:  Full Production Ventilation Arrangement

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This equipment list is summarized in Table 16-18.  Brand names have been included for reference only.

Table 16-18:  Full Production Airflow Determinations

Item	Total Units	Unit hp	Required m3/s	Required cfm
Jumbo — Atlas Copco 282	4	160	22.7	48,000
Bolter — Atlas Copco MD Boltec	5	160	28.3	60,000
Longhole Drill — Atlas Copco Simba M4	2	160	11.3	24,000
LHD — Cat R1700	5	353	62.5	132,375
LHD — Cat R1300	1	165	5.8	12,375
Haul Truck — Cat AD45	7	589	145.9	309,225
Grader — Cat M12	1	145	5.1	10,875
Emulsion Charger — Marcotte M40	3	193	20.5	43,425
Boom Truck — Marcotte M40	2	193	13.7	28,950
Fuel/Lube Truck — Marcotte M40	1	193	6.8	14,475
Scissor Lift — Marcotte M40 RL4000	2	193	13.7	28,950
Personnel Carrier — Marcotte M40	2	193	13.7	28,950
Shotcrete Sprayer — Normet Spraymec 6050WP	2	120	8.5	18,000
Transmixer — Normet MF500	3	161	17.1	36,225
Telehandler — Gehl FS12-42	3	115	12.2	25,875
Subtotal			387.8	821,700
Leakage (20% allowance)			77.6	164,340
Total			465 m3/s	986,040

16.6.3  Mine Air Cooling

The cooling design for the Los Gatos project is based on achieving an average stope and development reject WB (wet bulb) temperature of 28.5°C, using a surface intake air temperature of the high month average of 24.5°C and 33°C (WB and DB [dry bulb], respectively).  This section relates to the in-mine heat loads and the associated cooling requirements.  Primary heat load components are as follows:

·Virgin rock temperature (VRT) and related thermal gradient.

·Artificial heat loads (e.g., trackless equipment, fissure water, broken rock, pumps, auxiliary fans, personnel, lighting).

·Ambient summer fresh air intake temperatures.

·Mining depth and auto-compression.

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Each system (or combination thereof) contributes either positively or negatively towards a mine’s heat load and related cooling and refrigeration needs.  Taking into consideration the geographic location of the project and the trackless mining system, primary heat load contributors are as follows:

·VRT:  Gradient = 2.5°C/100 m with a surface starting temperature of 25.0°C and the deepest mining level (1175 masl) rock temperature range between 33°C and 37°C.  Heat load contribution is approximately 0.4 MW.

·Trackless Equipment:  Overall vehicle duties were calculated to be approximately 8.6 MW, and the related heat load in conjunction with the utilization of the equipment was calculated to be approximately 4.1 MW.

·Auto-compression Heat Load:  Surface to shaft depth related to approximately 1.1 MW.

·Other In-Mine Heat Loads:  Approximately 0.7 MW from auxiliary fans and pumping systems.

·Fissure Water Heat Load Contribution:  An estimated 0.7 MW of heat could be applied to the underground mine environment.  This assumes an inflow of 20 l/s at 54°C and leaves the mine via the dewatering system at 44°C.

·Backfill and Other Heat Loads (personnel, lighting, etc.):  Approximately 0.1 MW.

The calculated overall heat load is approximately 6.5 MW, including auto-compression.  The fresh air cooling capacity calculated is 1.7 MW.  As a result, the mechanical cooling required to maintain the reject temperature of 28.5 °C is 4.8 MW.  Including provisions for 20% energy loss, the refrigeration plant design capacity will be nominally 6.0 MW.  Although not included in the calculations, a precooling tower may be required for the underground service water to ensure it does not constitute an additional heat load on the underground environment.  Mine cooling design criteria and assumptions are summarized in Table 16-19.

Table 16-19:  Cooling Design Criteria and Assumptions

Criteria/Assumption	Value
Average Monthly High	33 °C
Average Relative Humidity	47%
Average Monthly High WB Temperature	24.5 °C
Max Allowable WB Temperature	28.5 °C
Surface Rock Temperature	25 °C
Geothermal Gradient	2.5 °C/100 m
Rock Thermal Conductivity	3 W/mC
Rock Wetness Fraction	0.25
LHDs	5
Haul Trucks	7

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The average monthly high and relative humidity come from statistical data and the Q451-02-010 Los Gatos Design Criteria from previous work.  The surface rock temperature comes from measured values from the diamond drill chamber; an accurate gradient was not available, so a standard geothermal gradient of 2.5°C/100 m was used.

The major pieces of diesel equipment (five LHDs and seven haul trucks) were placed in the model as point sources.  These point sources of heat were distributed into the different mining zones.  Figure 16-22 identifies the location of diesel equipment for the simulation.

Figure 16-22:  Heat Model Equipment Locations

Using 28.5°C as the design reject WB temperature, the simulation results show areas in which the WB temperatures exceeds the allowable threshold, supporting the calculations.  This indicates that cooling is required during the hotter months of the year.

To reduce the amount of equipment and infrastructure underground, a single 6 MW cooling plant is currently under construction on the surface near the South Vent Raise.  This cooling plant will chill water and send it to bulk air coolers (BACs) on the Central Vent Raise and the South Vent Raise.  These BACs will be located on the intake air stream of the Central and South Vent Raises and chill the air using a direct-contact spray of chilled water from the cooling plant.

The 6MW cooling plant supplies 2 MW of cooling capacity via a BAC to the Northwest Zone and 4 MW to the Central Zone and South Zone.

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These locations are shown in Figure 16-23.

Figure 16-23:  Heat Simulation Results

The model inputs will be updated as additional data on the rock temperature and water sources becomes available throughout the mine life.

16.7  Backfill

There are three main backfill products at Los Gatos, as follows.

·Cemented Rock Fill (CRF)

·Cemented Paste Fill (CPF)

·Waste Rock

CRF is essentially a concrete mix with a comparatively large rock particle size.  Crushed waste rock (-75 mm in size) is mixed with cement and water in a batch process.  The resulting mix is transferred underground to the mining areas via production mine trucks equipped with ejector boxes.

CPF is made by combining a portion of the mill tailings with binders (cement and fly ash) and a controlled amount of water to achieve a thick, mud-like (toothpaste) consistency.  The resulting product is transferred underground via boreholes and distributed to the stopes through heavy-wall pipe.  Pumping is typically required for paste delivery when the stopes being filled are outside a 45° cone (from the surface delivery point) and gravity is not sufficient to “push” the paste.

Unconsolidated waste rock will be used to backfill portions of the secondary transverse longhole stopes, drift-and-fill stopes, and longitudinal longhole stopes.

CPF typically provides a better final engineered product compared to CRF.  There is little or no product segregation during delivery and it is easier to tight fill in the drift-and-fill areas.  When mucking in the drift-and-fill areas, the paste tends to serve as a better marker for the floor (smoother and more visual compared to waste rock), which tends to reduce floor dilution.  A better backfill product also tends to reduce dilution when blasting up against backfill when mining secondary stopes.

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CRF will be used until the CPF is constructed.  Once the paste fill plant becomes operational, CPF will become the primary backfill product; however, the CRF plant will remain as a backup for the life of the mine.  Development waste rock will be used to displace the other two backfill products whenever possible.

16.7.1  Distribution

16.7.1.1  Cemented Rock Fill

Haulage trucks equipped with ejector boxes is loaded with CRF at the CRF plant on surface and transports the material into the mine.  Dedicated LHDs with rammers are required full time to push fill in the drift-and-fill stopes.  Dump blocks have been installed for the trucks during backfilling of the longhole stopes.

16.7.1.2  Cemented Paste Fill

Paste backfill is pumped from the paste plant on surface into the paste boreholes to the paste bay in the mine.  The paste bay is located by the Central Ventilation Raise on 1390 Level.  From the paste bay, paste is distributed throughout the mine in 150 mm diameter Schedule 80 pipes that transitions to 150 mm diameter HDPE pipe at the discharge locations.  Costing for heavy-wall pipe (for paste distribution) has been included in all internal ramp development.  An opportunity exists to reduce costs by transferring paste level to level using boreholes.

16.8  Conclusions and Recommendations

16.8.1  Conclusions

The mining operations at Los Gatos are being carried out through various methods, including cut and fill, and long hole stoping.  In the recently started mine development differences in water inflows and temperatures were noted from the FS.  Due to the unexpected water inflows and heat, the mine experienced significant challenges for continuing to excavate below the 1390 MSL level.  Production has started underground at Los Gatos.  Mining is being performed in the CZ and NWZ areas, starting at the 1390 level.  Due to the lack of development progress through the wet zones, a combined overhand and underhand approach has been chosen in order to optimize the mining sequence based on the existing excavation extents in order to meet the production and grade requirements of the mine.  Mining is currently focused on high grade zones that were near current development.  The CZ zone is producing mineralized material using cut and fill methods and will continue downward, as well as north and south from the current mining area in the upper center of the zone.

Due to the underhand approach, sill pillars are designed at necessary intervals in both cut-and-fill and longhole stoping blocks.  The underhand longhole stoping areas require re-mining or undercutting of the bottom sill drifts compared to the overhand approach.  This approach will be evaluated throughout the mine life to identify opportunities to implement overhand mining methods, for optimization of underground waste placement.

The equipment necessary for ventilation and mine cooling is currently being constructed.  The refrigeration plant is currently under construction.

Backfill will continue to be complemented with waste rock from the underground advances, as well as CRF.

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16.8.2  Recommendations

Continue to update the mine plan throughout the mine life and evaluate mining methods to ensure they are appropriate.

Continue to monitor the ventilation requirements and performance throughout the mine life, including temperature.

It should also consider updated geotechnical information to define the pillar size requirements.

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17.0  RECOVERY METHODS

This section summarizes the process plant design to be applied to the crushing, grinding, flotation, and dewatering facilities and the cyanide destruction installation for a 2,500 tonnes per day, lead, zinc and silver ore processing facility located in Chihuahua, Mexico.

The crushing plant is processing the run-of-mine (ROM) ore by using a primary jaw crusher to reduce the ore from a nominal 24 inches to a P100 of minus 175 mm.

The grinding circuit is a semi-autogenous (SAG) mill-ball-mill grinding circuit with subsequent processing in a flotation circuit.  The SAG mill operates in closed circuit with a vibrating screen.  The ball mill operates in closed circuit with hydrocyclones.

Cyclone overflow, the grinding circuit product, is fed to the flotation plant.  The flotation plant consists of lead and zinc flotation circuits.  The lead flotation circuit consists of rougher flotation and three-stage cleaner flotation.  The zinc flotation circuit consists of rougher flotation and five-stage cleaner flotation.

High levels of fluorine have been encountered in the ore from the mine since the original feasibility study was completed and since initial construction of the mill was completed.  The fluorine removal has required alterations to the milling process and an additional deep-froth flotation cell has been added to the zinc circuit, and one additional deep-froth flotation cell will be added to the zinc and lead circuits during 2020, to remove more fluorine, a deleterious mineral for sales.  The addition of this equipment should allow for the production of penalty free concentrate in the future.

Both final lead and zinc concentrates are thickened, filtered, and stored in concentrate storage facilities prior to loading in trucks for shipment.

Zinc rougher flotation tailings and zinc 1st cleaner scavenger tailings are combined to become the final tailing.  Tailings thickener underflow (100%) is pumped to a cyanide destruction facility.  After detox, forty percent (40%) of final tailings are pumped to backfill plant and the remaining (60%) are pumped to the tailings storage facility.

Plant water stream types include lead process water, zinc process water, fresh water, and potable water.

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The overall flowsheet is shown in Figure 17-1.

Figure 17-1:  Process Plant Overall Flowsheet

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17.1  Primary Crushing

Run-of-mine (ROM) ore is transported to the crushing plant area by rear-dump trucks and dumped into an open stockpile for manual metal removal.

The primary crushing line consists of a dump hopper, grizzly screen, rock breakers, crusher and associated dust collection and transfer equipment.  Run of Mine (ROM) ore is dumped into the dump hopper using a front-end loader.  The grizzly screen oversize feeds the jaw crusher.  Two mobile rock breakers are available, one to service the crusher or screen and another one to service ROM area stockpile.  The crusher reduces the size of run-of-mine ore from maximum 610 mm to approximately 100% passing 175 mm.  Crushed ore drops onto a belt conveyor that transports the crushed ore to a covered crushed ore stockpile.

Crushing production rate is monitored by belt scale mounted on the conveyor.  Tramp iron is removed using a magnet is located at the discharge of the primary crusher discharge conveyor.  A metal detector has been installed over conveyor.  Dust is controlled in the dump pocket with water sprays and dust collector vents positioned at the conveyor transfer points.  An air compressor and instrument air dryer have been installed for operation and maintenance.  A mobile crane is used for maintenance of the primary crusher.

17.2  Crushed Ore Conveying, Transport and Storage

Primary crushed ore is stockpiled in a covered dome.  A reclaim tunnel has been installed beneath the stockpile.  The stockpile contains approximately 3,000 tonnes of “live” ore storage and a total of 18,000 tonnes of storage.  When required, ore is moved from the “dead” storage area to the “live” storage area by front-end loader or bulldozer.

Ore is withdrawn from the coarse ore reclaim stockpile by variable speed belt feeders.  The feeders discharge to the transfer conveyor belt.  The transfer conveyor discharges to the SAG mill in the grinding circuit.  The ore reclaim rate is monitored by a belt scale mounted on the conveyor.

Dust control in the stockpile area is performed by the wet type dust collector systems.  There are two dust collector systems, one to control dust at the discharge of the stockpile feed conveyer, and another to control dust in the ore reclaim tunnel.

17.3  Grinding

Ore is ground to rougher flotation feed size in a semi-autogenous (SAG) mill primary grinding circuit and a ball mill secondary grinding circuit.

The SAG mill operates in closed circuit with a vibrating screen.  Water is added to the SAG mill to produce a slurry and the ore feed size is reduced as it traverses the SAG mill.  The SAG mill discharges onto a double deck screen with 6.35 mm bottom openings.  Screen oversize is recirculated to the SAG mill feed chute by a series of conveyors.  Screen undersize flows by gravity to the cyclone feed pump box.  A belt scale mounted on the recycle conveyor monitors the SAG mill recycle rate.  The target SAG grind is 80% passing 1,381 microns.

Secondary grinding is performed in a ball mill.  The ball mill operates in closed circuit with hydrocyclones.  Ball mill discharge is combined with vibrating screen undersize in the cyclone feed pump box and is

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pumped to hydrocyclone clusters.  Combined slurry is pumped using variable speed horizontal centrifugal slurry pumps (one operating and one standby) to the cyclone clusters.  Hydrocyclone underflow flows by gravity to the ball mill.  Hydrocyclone overflow (final grinding circuit product at 80% minus 45 microns) flows by gravity to the tramp oversize screen positioned prior to flotation circuit.

Cyclone overflow is sampled by primary samplers and analyzed by the lead and zinc on-stream analyzer for metallurgical control prior to flotation.  Cyclone overflow from cyclone cluster is also monitored for particle size distribution by a particle size monitor.

Zinc sulfate (ZnSO4) and sodium cyanide (NaCN) is added into the ball mill.

Grinding balls are added to the SAG mill and ball mill by ball loading systems.  Air compressors and an instrument air dryer provide service and instrument air for operations and maintenance.  An overhead crane has been installed for maintenance of the grinding mills.

17.4  Lead Flotation and Regrind

Hydrocyclone overflow flows by gravity to the lead flotation circuit.  The lead flotation circuit consists of one row of rougher cells and one row of cleaner cells.  The rougher row consists of five (5) 70 m3 tank type rougher flotation cells with a drop between each cell.  The lead rougher concentrate is sampled by a rougher concentrate primary sampler and pumped (one operating pump and one spare) to the lead regrind mill circuit.  Reground lead rougher concentrate flows by gravity from the lead cleaner conditioning tank to the lead first cleaner flotation cells.  The lead cleaner row consists of 10 flotation cells; three 10 m3 forced air first cleaner cell, two 10 m3 forced air first cleaner scavenger cell, three 10 m3 forced air second cleaner cells, and two 10 m3 forced air third cleaner cells.  The lead first cleaner concentrate is pumped (one operating pump and one spare) into the second cleaner flotation cells and the lead first cleaner scavenger concentrate is pumped (one operating pump and one spare) to the lead first cleaner conditioning tank.  Lead rougher tailings and lead first cleaner scavenger tailings flows by gravity into the zinc rougher conditioning tank.  The lead second cleaner concentrate is pumped to the lead third cleaner flotation cells.  The lead third cleaner concentrate flows by gravity to the lead concentrate thickener.

The concentrate samples cut by the samplers are analyzed for process control by the lead and zinc on-stream analyzer.  Tailings from rougher flotation cells and first cleaner scavenger cells are combined and sampled with primary samplers and analyzed by the lead and zinc on-stream analyzer.

Lead rougher concentrate is pumped to the lead regrind cyclone feed pump box and combined with the regrind mill discharge.  The combined slurry is pumped using fixed speed horizontal centrifugal slurry pumps (one operating and one spare) to a hydrocyclone cluster.  Overflow from the regrind cyclone cluster (final regrind circuit product) is sampled for particle size distribution analysis by the lead regrind cyclone particle size monitor, analyzed by the lead and zinc on-stream analyzer and flows by gravity to the lead cleaner conditioning tank and cyclone underflows by gravity to the lead regrind mill.  Product from the regrind mill reports to the lead regrind cyclone feed pump box.

Air compressors, air receivers, and instrument air dryer have been installed for general plant operation and maintenance.

A bridge crane has been installed for maintenance of the flotation and regrind equipment.

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17.5  Zinc Flotation and Regrind

Lead rougher tailings and lead first cleaner scavenger tailings flow by gravity to zinc rougher conditioning tank.  The zinc flotation circuit consists of one row of rougher cells and one row of cleaner cells.  The rougher row consists of five 70 m3 tank type rougher flotation cells.  The zinc rougher concentrate is sampled by the zinc rougher concentrate primary sampler and pumped (one operating pump and one spare) to the zinc regrind mill circuit.  The zinc cleaner row consists of 16 flotation cells; one bank of four 10 m3 forced air first cleaner flotation cells, two 10 m3 forced air first cleaner scavenger flotation cells, three 10 m3 forced air second cleaner flotation cells, three 10 m3 forced air third cleaner flotation cells, two 10 m3 forced air fourth cleaner flotation cells and two 10 m3 forced air fifth cleaner flotation cells.  Tailings from the zinc rougher cells is combined with zinc first cleaner scavenger tailings in a zinc tailings box and flows by gravity to the tailings sample box, then to the tailings thickener.

Reground zinc rougher concentrate flows by gravity from the zinc cleaner conditioning tank to the zinc first cleaner flotation cells.  The zinc first cleaner concentrate is pumped (one operating pump and one spare) into the zinc second cleaner flotation cell while the zinc first cleaner scavenger concentrate is pumped to zinc first cleaner conditioning tank.  The zinc second cleaner concentrate is pumped to the zinc third cleaner flotation cell.  The zinc third cleaner concentrate is pumped to the zinc fourth cleaner flotation cell.  The zinc fifth cleaner concentrate is pumped to the zinc fourth cleaner flotation cell.  The zinc fifth cleaner flotation concentrate is pumped to the zinc concentrate thickener.

The concentrate samples cut by the samplers are analyzed for process control by the lead and zinc on-stream analyzer.  Tailings from rougher flotation cells and first cleaner scavenger cells are combined and sampled with primary samplers and analyzed by the lead and zinc on-stream analyzer.

Zinc rougher concentrate is pumped to a zinc regrind hydrocyclone feed pump box and combined with the zinc regrind mill discharge.  The combined slurry is pumped using fixed speed horizontal centrifugal slurry pumps (one operating and one spare) to the zinc regrind hydrocyclone cluster.  Overflow from the zinc regrind cyclone cluster is sampled by sampler for particle size distribution analysis by the zinc regrind cyclone particle size monitor, analyzed by the lead and zinc on-stream analyzer and flow by gravity to the zinc cleaner conditioning tank and underflows by gravity to the zinc regrind mill.  Product from the regrind mill reports to the zinc regrind cyclone feed pump box.

17.6  Lead Concentrate Dewatering

Concentrate from the lead third cleaner flotation cells is pumped to a lead concentrate thickener.  The concentrate thickener overflow is pumped back to the thickener feed for dilution and thickener spray bar to control froth, or to the lead process water tank.  The concentrate thickener underflow is pumped (one operating pump and one spare) to an agitated storage tank, (with24 hours of retention capability) and then to a pressure filter.  Filter cake is discharged to a covered stockpile.

Concentrates, both lead and zinc, are reclaimed by front-end loader onto highway haulage trucks.  A truck scale is located near the concentrate loadout area.

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17.7  Zinc Concentrate Dewatering

Concentrate from the zinc fifth cleaner flotation cell is pumped to a zinc concentrate thickener.  The concentrate thickener overflow is pumped back to the thickener feed for dilution and thickener spray bar to control froth or to the zinc process water tank.  The concentrate thickener underflow is pumped to an agitated storage tank, (with24 hours of retention capability) and then to a pressure filter.  Filter cake is discharged to a covered stockpile.

17.8  Tailing Dewatering

Tailings from the zinc rougher flotation row and tailings from the zinc first cleaner scavenger flotation row flow to a high rate tailings thickener.  Thickener overflow is pumped from the tailings thickener overflow tank to the lead process water tank.  Thickener underflow from the tailing thickener is pumped by variable speed horizontal centrifugal slurry pumps (one operating and one stand-by) to the tailing dam (60%) and backfill plant (40%).

17.9  Cyanide Destruction (SO2/Air Process)

Cyanide in the tailing thickener underflow is neutralized by the cyanide destruction circuit.  Then the detoxed tailings are pumped to the backfill plant facility.

17.10  Reagent

Reagents requiring receiving, handling, mixing, and distribution systems include:

·Sodium Cyanide (NaCN)

·Zinc Sulfate (ZnSO4·7H2O)

·Aerophine 3418A (Promoter)

·Copper Sulfate (CuSO4·5H2O)

·Sodium Isopropyl Xanthate (SIPX)

·Methyl Isobutyl Carbinol (MIBC, frother)

·Flocculant

·Sodium Metabisulfite, Na2S2O5 (MBS)

·Lime

17.11  Water System

17.11.1  Fresh Water

Fresh water is supplied from four to five wells, plus one backup well located on the property, as well as mine dewatering.  Fresh Water from the wells is pumped to a fresh/fire water tank.  The fresh water distribution system provides fresh water for process requirements such as process water makeup, reagent mixing and gland water.  There are onsite three (3) dedicated water tanks, one each for Gland water, Mill water, and the third potable water.  Controls have been installed to ensure flow to the process water system when the raw water system is operating.  From the fresh water tank, low pressure process water flows to the systems that do not require high pressure.  Booster pumps have been installed to provide

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high pressure water to the systems that require it including pump gland water.  Gland water is provided for sealing each pump without return.  Pumps and a control system have been installed at the fresh water tank to provide pressure to the fire water system.

17.11.2  Process Water

17.11.2.1  Process Water — Lead Circuit

The lead process water tank receives overflow from the lead concentrate thickener, tailing thickener and water reclaimed from the tailing dam.  The lead process water is used as makeup water in the primary cyclone feed sump.  Fresh water can be added to the lead process water tank if necessary.  This lead process water is not suitable for general distribution throughout the process plant.  Water is reclaimed from the tailing dam using reclaim water pumps.

17.11.2.2  Process Water — Zinc Circuit

Overflow from the zinc concentrate thickener and lead process water excess overflow is recycled to the zinc process water tank and used as makeup water in the zinc flotation circuit.  Fresh water can be added to the zinc process water tank.

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18.0  PROJECT INFRASTRUCTURE

18.1  Existing Infrastructure and Services

18.1.1  Location

The Los Gatos Project is located in the Municipality of Satevó, Chihuahua, Mexico, approximately 160 kilometers Southwest of the state capital of Chihuahua City and about 8 km West of San José del Sitio, Chihuahua.

18.1.2  Site Access Roads

The access road from Chihuahua, Chihuahua, Mexico is newly paved.  A portion of the road from San José del Sitio has been rerouted to the mine site to minimize interference with the river that runs by the mine property.

There is one bridge crossing of the Santo Toribio creek, with the purpose of minimizing risks due to crossing.

18.1.3  Buildings

MPR has already established the following facilities:

·Exploration camp with cafeteria onsite for 60 people;

·Exploration sample preparation and core/sample storage facility in San José del Sitio;

·Concrete batch plant that is currently in use for the mine underground requirements;

·Small vehicle maintenance shop;

·Emergency generators for the mine underground facility;

·Exploration waste dump;

·Explosive magazine; and

·Several preliminary access roads to the project facilities.

·Processing facilities for grinding, floatation, reagents, concentrate storage, and flocculant storage.

·Process buildings for truck scale, truck wash and truck sampling

·Assay laboratory

·Mill area change house

·Maintenance shop

·Process area office

·Mill area administration & engineering building

·Multipurpose room

·Training room

·Infirmary

·Warehouse

·Hazardous storage

·Security guard gates

·Lunch room

·Owner’s camp for 350 people

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The site buildings are pre-engineered structural steel construction.  The buildings have R38 insulation in the roof and R19 in the walls.  The steel has a primer and special paint coating to prevent rust.  The roof and wall panels are standard 24-gauge steel.  The buildings include overhead and manway doors.  The buildings meet the local requirements for the project.  For processing, major equipment is supported on concrete piers or pedestals as required to minimize vibration.  The building foundations are on separate foundations to minimize vibration transfer into the building structure.

The Owners camp consists of 192 men’s staff rooms with two beds in each, 64 women’s staff rooms with two beds in each, 48 supervisor rooms and 56 manager/visitor rooms.  The manager and supervisor rooms have only one person per room and include a bathroom in each room.  The men’s staff rooms have six group bathrooms and the women’s staff rooms have four group bathrooms.

In addition, the camp includes the following buildings: multipurpose room, kitchen and cafeteria, laundry, infirmary, store to sell employees small items, and maintenance building.

The camp has 100% backup emergency power in the event of a power outage and includes fresh and potable water.

18.1.4  Communications

The site is currently serviced by a satellite dish-based internet and TV connection.  Upgrades have been made to the communication services.

Communication for the surface facilities are provided via handheld radios at the same frequency as the mine leaky feeder.  This system provides communication capabilities with both the surface and underground personnel.

To provide communication for control systems, CCTV, telephone, power monitoring, and data to all the surface facilities of the mine, a 48-strand, single-mode, fiber-optic cable backbone ring has been installed from the controls alarms room in the office / dry facility.  This included fiber junction boxes at the following surface locations:

·Office / dry facility

·Maintenance facility

·Southeast raise ventilation plant

·Central raise ventilation plant

·Cemented paste fill plant

·Cement rock fill plant

·Cooling plant

·Fuel facility

·Compressor plant

To provide communication for fire systems, a separate 6-strand, single-mode, fiber-optic cable backbone ring connects the controls alarms room in the office / dry facility.  This includes fiber junction boxes at the same surface locations listed above.

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18.1.5  Personnel

As a part of the company’s commitment to adding value to the local communities and building capacity in Mexico, over 99% of the operations workforce is from Mexico.  Most of the workforce is being sourced from local towns, with skilled labor from the nearby cities of Chihuahua and Hidalgo del Parral.

18.1.6  Power Supply

Power to the Cerro Los Gatos Mining project site is supplied via a 115-kV utility transmission line.  This originates from the ‘San Francisco de Borja’ substation in Satevó (Chihuahua), where a new 115 kV connection has recently been installed.

This connection includes a 115-kV switch and feeder, a 7.5 MVAR capacitor bank with circuit protection, and substation area modifications for installing new equipment.  The 115-kV utility transmission line is approximately 66 km long.  The connection was completed with minor changes to the original layout, due to land negotiations.

A 115-kV loop has been installed near the Cerro Los Gatos main substation.  This loop includes 115kV disconnects, high voltage connections to the 115 kV — 13.8 kV substation type transformers, and utility metering.  Power distribution on site is monitored and metered at respective facilities for power management.

18.1.7  Power Distribution

Upon review of the proposed equipment list, a total electrical load of approximately 24 MW was determined.  This electrical load comprises of about 12 MW for the process plant and water management, 8.6 MW for the underground mine and surface facilities, 0.9 MW for the camp, and 2.7 MW contingency for motor starting capability and future additions.  The electrical load is summarized in Table 18-1.

Table 18-1:  Mine Power Requirements by Area

Area	Electrical Load MW
Underground Mine	5.0
Mine Surface Facilities	3.6
Crushing	0.4
Grinding	5.5
Flotation + Reagents	2.8
Filtration	0.8
Tails + Cyanide Destruction + Lime + Flocculant	0.7
Water (Dewatering wells + cooling)	1.4
Ancillary Buildings	0.2
Camp	0.9
Contingency (Future addition)	1.7
Reserve Capacity for motor starting	1.0
Total	24.0

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The utility power (115 kV) entering the mine site has been stepped down to 13.8 kV at the mine main substation, via two 15/20 MVA, 3-phase, 60 Hz transformers.  These transformers provide power to a 13.8 kV switchgear, consisting of two main circuit breakers and a tie breaker feeding two buses for distributing power.  One bus provides power via duct-bank to the SAG mill and ball mill.  An overhead power distribution system to the primary crushing, mine surface facilities, ancillary buildings, water management and camp has also been provided.  The other bus provides power via duct-bank to the process plant electrical rooms and overhead power line to the underground mine disconnect.  The main substation switchgear has individual feeder circuit breakers designated for the above-mentioned loads, thereby maintaining isolation between the mine, the mills, the process plant, the camp and other surface facilities to avoid voltage drops due to motor starting or power supply interference due to maintenance shutdowns.

On site power distribution voltage is rated 13.8 kV (13800V).  The power distribution within facilities is rated for medium voltage distribution at 4.16 kV (4160V) and low voltage distribution at 480V.  The process plant is the most significant load and, for that reason, the distribution switchgear for the process plant is located near the main substation that distributes power to the process plant electrical rooms housing motor control for respective process loads.  The distribution switchgear includes two 7.5 MVA transformers, stepping down 13.8 kV to 4.16 kV.  One transformer is designated for the SAG and ball mills, whereas the other transformer provides power to the mill building electrical room and filter building electrical room.  All other areas have their individual distribution switchgear stepping down from 13.8 kV to 4.16 kV or 480V.

The process plant includes two electrical rooms, one in the mill building and the other in the filter building.  The mill building electrical room provides power and control to stockpile, grinding, flotation and reagents.  The filter building electrical room provides power and control for filtration, thickening, and ancillary buildings near the process plant.  The water management system includes dewatering wells (19 new + 1 existing), cooling tower and water supply pumps to the process plant and camp.

All project facilities have power at 13.8 kV delivered to them, via overhead power lines or at 4.16 kV via underground duct-bank.  Depending on the load, the medium voltage of 4.16 KV is utilized directly for feeding motors greater than 300 HP and it is further stepped down to a 480 VAC, 3-phase, 3-wire system for feeding motors below 300 HP.  The 480 VAC systems are further stepped down to 208Y/120VAC to feed lighting loads and general office equipment (receptacles, computers, printers, etc.) or 24VDC to feed instrumentation control requirements.  Power distribution design follows the federal, state and local standards.

The mine site has been provided with a grounding grid to which all building steel, equipment, etc. are connected for safety.  This grounding grid consists of a #4/0 AWG bare copper conductor buried below ground connecting all electrical equipment.  All above-ground connections except connections to building steel are mechanical type connections so that equipment can be removed or replaced easily.  All underground connections including those to building steel are of the thermo-weld type.  A test well is provided for periodically measuring / testing the resistance of the ground grid.  Lightning protection is installed on overhead pole lines, building structures, etc. and grounded separately from equipment grounding.  Grounding design follows the federal, state and local standards.

Lighting is of the high intensity discharge type.  High pressure sodium type light fixtures or LED are utilized for exterior areas and high bay interior applications.  Metal halide lighting fixtures are utilized indoors for low bay application and where color rendition is a factor.  Fluorescent lighting fixtures are used in interior applications such as office lighting, electrical rooms, etc.  All areas are equipped with emergency light

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fixtures utilizing battery packs which provide a minimum of 90 minutes of illumination.  Lighting levels are designated by the Illumination Engineering Society (IES) published guidelines.

A computer-based data gathering system, supervisory control and data acquisition system (SCADA), have been incorporated in the control and monitoring of all process operations.  The SCADA system uses remote termination devices to channel appropriate control and monitoring signals from field locations back to the central processing unit (CPU) computer where an operator can physically operate equipment from his computer workstation.  The configuration of the SCADA is based on the latest industrial standards.  A programmable logic controller (PLC) system has been installed in respective areas, gathering information from the input and output signals from instruments and motor control equipment installed for monitoring, control and safety.  The SCADA processes and records all communications with respective PLCs.  An uninterruptable power supply (UPS) provides power to each PLC.

Standby diesel generators have been provided to handle emergency situations at the process plant, water management system, and camp.  These standby generators are rated 480V, 1500 kW each and are connected via automatic transfer switch at respective locations.  Depending on the switchgear voltage, a step-up transformer has been installed for 13.8 kV or 4.16 kV connections, as needed.  The standby power is provided to critical equipment like flotation cells, thickeners, select pumps and other equipment that may affect the process production line should they stop operating.  The UPS provides backup power to the control system and emergency light fixtures utilize battery packs, in case of power supply fails.

18.2  Site Development

The Project to date includes development at the following major locations:

·The mining areas

·The mine surface facilities, including the fuel facility, office/dry facility, explosives magazine, Compressor plant, ventilation and cooling, cemented rock fill plant, cemented paste fill plant, and service water

·The crushing plant area

·The overland conveyor and primary crush stockpile

·Mill area

·Filter area

·Thickener area

·Cyanide destruction area

·Overland pipelines

·Fresh water pumping, storage and distribution

·Access and internal roads

·Power line tie into the local utility

18.2.1  Mine Surface Facilities

Surface facilities that have been completed or are under construction include:

·Maintenance facility

·Fuel facility

·Office / dry facility

·Explosive magazine

·Compressor plant

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·Ventilation and cooling

·Cemented rock fill plant

·Cemented paste fill plant (June 2021)

·Service water

·Electrical

·Communications

18.3  Water Source

Raw water to meet potable and non-potable water demand is supplied by groundwater pumped from dewatering wells.  The well water is directed to a cooling tower to lower the temperature from the expected 50 to 60°C to 40°C prior to use.  Analyses of groundwater from existing dewatering indicates the water quality is suitable for the proposed potable and non-potable uses.

The total raw water demand is estimated to be approximately 2,605,600 liters per day, including approximately 100,000 liters per day for potable use (see Section 18.3.1 ), and approximately 1,209,600 liters per day for process water and 1,296,000 liters per day for drill cooling and underground activities.  The mill process water demand accounts for approximately 660,000 liters per day of reclaim water obtained from the tailings storage facility being recycled back into process.  The 2,605,600 liters per day of water would be supplied from the dewatering wells.  Additional information regarding the dewatering wells is provided in Section 24.0 .

18.3.1  Potable Water Supply

Potable water is used for drinking water, cleaning, change rooms, laboratory water and safety showers.  Potable water is not required for the process requirements.

The majority of the potable water is required at the Owners camp and the mine and mill change rooms.  The potable water required at the camp is provided by a modular potable water system.

The water requirements for specific areas of the project are noted below:

·Camp site area 350 people — 70,000 liters / day

·Mine portal area change room — 15,000 liters / day

·Mill portal area change room — 15,000 liters / day

·The potable water storage tank is sized for 100,000 liters storage or 24 hours retention time.

18.3.2  Raw Water Distribution System

Based on the water balance study and hydrological assessment, there is adequate raw water available from the mine dewatering pumps.  The water coming up from underground is piped to the cooling tower location.  The water from underground is 50°C and is cooled to 40°C prior to being pumped into the raw water tank near the well sites.  Construction included a raw water storage and distribution system for the project that transports water to the mine portal, owner’s and contractor’s camp, surface mine facilities, crusher area, mill area, and paste plant area.

In order to minimize the number of services, fire water is provided via the raw water system.  There is sufficient elevation and tanks have been located such that fire water supply is available by gravity.  This

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provides sufficient pressure to provide a secure fire service.  A minimum volume of water is held in the raw water tanks at all times.

18.3.3  Process Water Supply

Process water is prepared at the process plant and is recycled to the extent possible.  Makeup water is kept to a minimum.  Process water quality is monitored and used for process plan makeup, process plant sprays, and process plant washdown.

18.3.4  Sewage Waste Water Treatment

Sewage water treatment systems were included to handle waste from the kitchen, bathroom and shower facilities.  Four separate systems were included for the owner’s camp, mine portal area, primary crusher area, and mill area.

18.4  Waste Disposal

For storage and management of landfill disposal, a building has been constructed with a zone for non-hazardous waste and a second zone for hazardous waste; these zones are adjacent.  Waste disposal facilities have been constructed in accordance with the Mexican Regulations landfill requirements.

The storage and management of this area is per Environmental Area of MPR and must coordinate with the Construction Area and Operation Area.

The hazardous waste is collected and disposed of by a certified and authorized company under the Mexican Regulations.

18.5  Underground Infrastructure

The underground infrastructure includes the following items:

·Mine dewatering

·Materials handling

·Electrical power and distribution

·Compressed air

·Service water

·Service bay

·Fuels and lubricants

·Refuge stations

·Sanitary facilities

·Communications

18.5.1  Mine Dewatering

During mine production, dirty water is collected in sumps located on each level of the NWZ, CZ, SEZ, and SEZ2.  The dirty water in the mine originates from the following sources:

·Drill water

·Mine service water

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·Fissure (ground) water

·Backfill seepage

·Backfill flush water

The above sources of dirty water total 30 l/s (475 USgpm) from three areas (the NWZ, CZ, and SEZs) for a total 90 l/s (1,427 USgpm).

All mine roadways are driven on a gradient and include a ditch system to allow dirty water to flow to a series of collection sumps.  Decline ramp headings require face pumps to transfer water from the face to the sumps.

Two types of sumps have been planned for the mine, borehole sumps and level dewatering sumps, in addition to the main dewatering sump.  The main sump has been completed to the FS specifications.  A solids separation system is included.  In addition to the original four sumps planned, an additional sump has been installed in the area intended as the Preventative Maintenance shop.  This temporary sump is being utilized to handle dewatering that exceeds the main sump’s designed capacity.  It will continue providing backup capacity for the main sump as needed.  These sumps are described in the following subsections.

18.5.1.1  Borehole Sumps

Borehole sumps collect water from the level and from borehole sumps on levels above and transfer the water via gravity flow to a lower sump through a pair of 150 mm boreholes located at the entrance to the sump.  This is illustrated in Figure 18-1.

Figure 18-1:  Borehole Sump

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18.5.1.2  Level Dewatering Sumps

Level dewatering sumps consist of a tank, agitator, and a pump mounted on a skid.  Level dewatering sumps are strategically placed to collect water from the borehole sumps and other level dewatering sumps and pump it to the main dewatering sump located off the decline ramp at 1390 Level.  A typical level dewatering sump is shown in Figure 18-2.

Figure 18-2:  Level Dewatering Sump

18.5.1.3  Main Dewatering Sump

The main dewatering sump (shown in Figure 18-3) consists of a reinforced concrete dam, two agitators, and two pumps capable of pumping a combined maximum of 130 l/s (2,061 US gpm) of dirty water from 1390 Level to surface through a 250 mm diameter pipe.  Provisions have been made for adding a third pump in the future.  The mine dewatering system is designed to maintain fines (slimes) in suspension throughout the system and pump water and fines to surface.  A solids separation system, known as a MudWizard, was included in the installation of the main sump.

Figure 18-3:  Main Dewatering Sump

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18.5.2  Materials Handling

Materials to support development and production (bolts, screens, drilling rods, etc.) are transported into the mine using a boom truck with a flat deck.  Maintenance materials for the pumps and underground equipment are hauled into the mine via maintenance service trucks.

Explosive trucks (the same vehicles used for charging the blastholes at the working faces) deliver the explosives from the surface magazine to the working faces in the mine.  Explosive utility vehicles transport the detonators and ancillary explosives from the surface magazine to the working faces.

18.5.3  Electrical Power and Distribution

Two 13.8 kV feeders provide power to the main switchgear located near the main dewatering sump on the 1390 Level.  One feeder delivers permanent power from surface through a borehole, and the other feeder delivers 13.8 kV power from the backup generators on surface through the decline.

Power for underground distribution from the main switchgear is via four feeders.  One feeder is for the main dewatering sump, booster fan, and service bay.  The remaining three feeders feed 13.8 kV to each of the mining zones:  NWZ, CZ, and SEZs.  A 15 kV tap box with 200 A load-break elbow is installed on each level of the three mining zones.

18.5.3.1  Main Dewatering Sump, Booster Fan and Service Bay

The main dewatering sump and service bay switchgear is 800 A / 4.16 kV and is designed to feed the two 261 kW main dewatering sump pumps, the 125-kW booster fan, and the service bay distribution transformer at 4.16 kV.  Medium voltage starters for the pumps and fan are in the switchgear, and the shop transformer is fed from a fused disconnect switch.

The transformer in the service bay transforms the voltage from 4.16 kV to 480/277 V, which feeds a 600 A / 480 V motor control center (MCC).  This MCC feeds additional main dewatering sump equipment, the refuge chamber, and the service bay.

18.5.3.2  Level Dewatering Sumps

The electrical distribution system follows the mining zones:  NWZ, CZ, and SEZs. For each zone, a 15-kV tap box is located on each level.  The tap box is designed for use with 200 A load-break elbows and up to four connections per tap box.

Levels with a level dewatering sump have a 200 A fused load-break switch, 225 kVA, 13.8 kV-480/277 V transformer and a 400 A distribution panel installed close to the sump.  The 200 A load-break switch is fed from the 15-kV tap box located on that level.

18.5.3.3  Development Mine Load Centers

Development MLCs are required in all mining zones.  The MLCs consist of a 15-kV fused load-break switch, 750 kVA, 13.8 kV-480/277 V, and distribution.  The distribution is for a production drill, jumbo, bolter, and development fans.  The MLCs are fed from the 15-kV tap box located on that level.

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18.5.4  Compressed Air

The compressed air system at Los Gatos consists of mine-wide distribution from the main surface facility.  There is an existing permanent surface air compressor plant consisting of two compressors, each with a capacity of 488 cfm (for a total of 976 cfm).  The expected compressed air demand from the mine is outlined in Table 18-2.

Table 18-2:  Mine Compressed Air Demand

Item	Quantity	Demand (cfm)	Utilization (%)		Actual Demand (cfm)
Jackleg Drills	2	186	5	%	19
Shotcrete Machine	1	777	81	%	631
Air Tuggers	2	155	25	%	78
Vent Doors	3	52	20	%	31
Refuge Stations*	2	31	100	%	62
Maintenance Shop	3	83	81	%	202
Total	—	1,284	—		1,023

* Air required in combination lunchroom/refuge stations only.

An additional 488 cfm compressor has been added to the surface plant, increasing the capacity to 1,551 cfm and providing backup for the existing two units.

Compressed air is distributed throughout the mine via 150 mm diameter pipe on the decline/ramps and branches off with 100 mm and 50 mm diameter pipe throughout the mining levels.  Mobile drilling equipment (e.g., jumbos, cable bolters, production drills) have on-board compressors to provide the primary compressed air requirements.

18.5.5  Service Water

The underground water consumption estimate is based on the amount of water anticipated to be used by equipment and underground processes.  The equipment and underground processes are outlined below:

·Jumbo drills

·Bolters

·Shotcrete sprayers

·Longhole drills

·Diamond drills

·Raise bores

·Jackleg drills

·Hose and nozzle (dust control and cleanup)

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A summary of the underground water consumption estimates is outlined in Table 18-3.

Table 18-3:  Underground Process Water Requirements

Facility Description	Average Flow l/s (US gpm)		Peak Flow l/s (US gpm)
Development Crew #1	2.0 (32.1	)	5.3 (84.1	)
Development Crew #2	2.0 (32.1	)	5.3 (84.1	)
Production Crew #1 (Development)	2.0 (31.5	)	5.9 (94.1	)
Production Crew #2 (Development)	2.0 (31.5	)	5.9 (94.1	)
Production Crew #3 (Blasthole)	2.3 (35.8	)	5.4 (86	)
Production Crew #4 (Blasthole)	2.1 (33.6	)	5.4 (86	)
Raise Crew	0.6 (9.1	)	4.2 (66	)
Diamond Drilling	0.2 (2.7	)	0.5 (8	)
Raise bore Drilling	0.3 (5.0	)	0.6 (10	)
Construction Crews	0.1 (1.3	)	0.9 (15	)
Underground Infrastructure (Misc.)	0.2 (3.3	)	1.3 (20	)
Leakage	1.3 (21.6	)	3.3 (53	)
Total	15.1 (239.6	)	36.6 (580	)*

* Peak flows for the various facilities occur at different times in a 12-hour shift.  The water balance in Section 24 identifies Hour 3 of the 12-hour shift as having the peak flow of
36.6 L/s (580 US gpm).

A 150 mm diameter process water line has been included in the design from the mine portal to supply the underground mine via the ramp.  As required, each mining zone is equipped with pressure reducing valve (PRV) stations to limit the line pressure supplying the local levels.

All PRV stations include a full-time, in-line duty PRV; a bypass line with a full-time spare PRV; and isolation valves.  This configuration ensures that water supply can be maintained with minimal interruption.  Large face diameter pressure gauges are provided on upstream and downstream sides of all PRV stations for visual confirmation of the operating condition of the PRV and setting valves.  A pressure safety valve (PSV) is included at each PRV assembly to ensure that over-pressure conditions do not occur in level distribution.  Air bleed connections are provided on the assemblies to allow air removal during the pipeline filling process.

One booster pump station is required in the NWZ and SEZ to allow for adequate water pressure at the face in the upper levels.  The booster pump stations are standalone units mounted on a common steel skid.  A process water booster pump station generally consists of the following:

·Pumps (one operating, one standby)

·Steel water reservoir

·Piping

·Controls

·Electrics

·Instrumentation

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18.5.6  Service Bay

A service bay has been constructed close to the base of the Central Raise, as shown in Figure 18-4.  This service bay is used to complete minor repairs and oil changes on mobile equipment that does not typically travel to surface (e.g., jumbos, bolters, and longhole drills).  This saves time by not requiring equipment be taken to the surface maintenance shop for these minor maintenance activities.

Figure 18-4:  Service Bay Location

Mexican mining regulations require maintenance facilities to have two independent exits, be constructed with noncombustible coating or fire-resistant materials, and have adequate ventilation.  Fire doors (including fused-link activation) have been installed at the entrance to the service bay, and a second means of egress is provided in the ventilation bulkhead at the rear of the service bay.  The bulkhead contains a man door to provide escape into the exhaust air and also has an air regulator to control the airflow through the service bay.

The service bay is equipped with the following services and equipment:

·Compressed air

·Lighting

·Telephone

·Concrete floor

·Jib cranes

·Work benches

·Parts storage

·Lubricant storage

·Hose reels

·Waste fluid storage

·Fire suppression

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A layout of the service bay is shown in Figure 18-5.

Figure 18-5: Service Bay Layout

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18.5.7  Fuels and Lubricants

Los Gatos is a trackless mining operation.  All mobile equipment is diesel driven and has Tier 4 engines operating on diesel fuel with sulfur content less than 15 ppm.  The quantity of fuel stored underground is not to exceed 72 hours of consumption.

Diesel is transported underground via a SatStat storage and handling system.  Fuel stations have been strategically placed in exhaust air drifts to support development and production.  The locations may change as development and production activities advance.  A total of three SatStats have been budgeted for the mine.

Forklift handling pockets are built into the units, as well as pick points for handling ease.  SatStat systems have various options for combined or individual storage and dispensing, and all systems are equipped with fire suppression.  Dispensing systems, including SatStat, are equipped with Wiggins-type coupling connections, and mobile equipment have corresponding coupling connections.  These types of connections limit contamination, spillage, and fire hazards.

Lube and fuel vehicles service fixed, and mobile equipment normally stationed at the working face or vehicles that do not typically travel to surface (e.g., jumbos, bolters, and longhole drills).

18.5.8  Communications

Communication throughout the mine are provided via leaky feeder radio and a voice over internet protocol (VOIP) telephone as a secondary system.  An allowance for an emergency dispatch system is provided to allow one-way, mine-wide emergency communication from surface to all cap lamps equipped with the personnel emergency dispatch system pager.  An allowance for a basic vehicle dispatch system is also included.

To provide communication for control systems, closed-circuit television (CCTV), telephones, power monitoring, and data collection for equipment in the mine, a pair of 48-strand, single-mode, fiber-optic backbone cables have been routed from surface throughout the underground to connect various pieces of mechanical and electrical equipment in the mine.  The pair of fiber-optic cables have been routed back to the controls alarm room in the office/dry facility on surface.

A separate 6-strand, single-mode, fiber-optic cable backbone has been provided from the controls alarm room in the office/dry facility to the booster fan in the mine to provide communication for the fire system.

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18.5.9  Refuge Stations

Underground portable and permanent refuge stations are required to house personnel in a secure, hazard-free location during emergency conditions.  The maximum distance for personnel to walk to a refuge station is 750 m.  There is currently one portable refuge station on site, and four more 16-person portable refuge stations plus two permanent refuge stations have been budgeted.  Permanent refuge stations, which also serve as lunchrooms, are located in the NWZ and the CZ

Refuge stations have been designed to meet Official Mexican Standard NOM-023-STPS-2012, Underground Mines and Open Pit Mines — Occupational Health and Safety Conditions.  The stations have 96 hours of autonomy and include the following:

·Uninterruptible power supply of up to 96 hours (without reliance on mine power)

·Breathable air (oxygen) supply

·Self-rescuers (quantity equal to the capacity of the station)

·Emergency food and water rations

·Carbon dioxide and carbon monoxide scrubbers

·Communications equipment

·Didactic materials to be used during confinement

·Air conditioning equipment

·Gas monitoring for inside and outside environment

·Toilet

·Lighting

·Blast rating to 5 psi

·Signs prohibiting smoking

·Simple to operate under emergency conditions

·First aid equipment (per MHSA Regulation No. 24.1.1)

·Lifting lugs, skid base, and forklift slots for portability

There are two means of warning underground mine personnel in the event of an emergency:

·Stench gas (ethyl mercaptan) will be introduced into the compressed air line and mine ventilation air so that it quickly travels throughout the mine.  Miners are trained to stop working and report to the nearest refuge station (or, in some cases, exit to surface) once the distinctive odor is detected.

·Another warning signal device is included in underground cap lamps that will be activated, signaling all personnel to stop work and proceed to the nearest refuge station as detailed in the area’s ventilation and rescue plan.

Self-contained rescue packs are issued to all employees working underground.  The employees are trained in the use of the self-rescuers.  The self-rescuers are serviced and checked every 12 months and replaced after 10 years in service.

18.5.10  Sanitary Facilities

Costs are included for four latrines and a mobile service vehicle for cleaning the units.  One latrine is located in each of the mining zones.

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19.0  MARKET STUDIES AND CONTRACTS

The Los Gatos project is a Joint Venture holding rights to 16 concessions (see Table 4-1) through its 100%-owned Mexican subsidiary company, Minera Plata Real S. de R.L. de C.V. (MPR).  The Los Gatos Joint Venture is 51.5% owned by Sunshine Silver Mining & Refining Corporation and 48.5% owned by Dowa Metals & Mining Co., Ltd.  MPR holds the rights to the concessions of Los Gatos and Paula Adorada through exploration agreements with purchase options.  These have been duly executed and recorded in the Mexican Public Registry of Mines (Vazquez, Sierra, and García, S.C) for the mining concessions title number 231498 dated March 4, 2008.

MPR has a royalty agreement with La Cuesta International S.A. de C.V. under the terms of the document, Contrato de Exploración, Explotación y Promesa — La Cuesta International, S.A. de C.V. and Minera Plata Real, S.A. de C.V., dated April 2006.  Under the terms, MPR paid a royalty payment of US $40,000 per year during the preproduction period.  When production is initiated a net smelter return (NSR) royalty would start at 2% on production from the Los Gatos concession.  This is reduced to 0.5% upon all payments reaching $10 million) and 0.5% net smelter return from lands within a one-kilometer boundary of the Los Gatos concession.  Upon commencing production, payments under the royalty agreement were deferred until March 31, 2021 with an annual interest rate of 4.5% applied to the outstanding balance.  During the deferral period, MPR pays a royalty payment of $100,000 per year until January 2021.  The agreement has no expiration date; however, the Company may terminate the agreement upon 30-day official termination notification.

Indicative smelting and refining terms have been prepared for the zinc and lead concentrates from Los Gatos.  These terms are summarized in Table 19-1 and Table 19-2 below.

Table 19-1:  Indicative Smelting and Refining Terms for Zinc Concentrate

Description	Treatment Charges
Base Rate	$299.75/tonne of zinc concentrate (dry)

Elements	Payable Terms
Payable Metals
Zinc	8% deduct; maximum pay for of 85.0%
Silver	If greater than 93.3 g/t, then 93.3 g/t deduct; maximum pay for 70%.
Penalties
Arsenic	If %As is greater than 0.1%As, a penalty of $3.00 per tonne of concentrate for every 0.1%As greater than 0.1%As.
Fluorine	If F ppm is greater than 500 ppm, a penalty of $12.00 per tonne of concentrate, if 300-500 ppm, a penalty of $9.00 per tonne of concentrate, if 200-300 ppm, then a penalty of $6.00 per tonne of concentrate, if 100-200 ppm, then a penalty of $3.00 per tonne of concentrate.

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Table 19-2:  Indicative Smelting and Refining Terms for Lead Concentrate

Description	Treatment/Refining Charges
Base Rate — Concentrate	$175/tonne of lead concentrate (dry)
Scale Rate — Concentrate	None
Gold — Refining (base rate)	$20.00/ounce of payable gold
Gold — Refining (scale rate)	None
Silver — Refining (base rate)	$1.50/ounce of payable silver

Elements	Payable Terms
Payable Metals
Lead	Pay lower of 3% deduction or 95%
Silver	Pay lower of 50 gm/tonne of concentrate deduction or 95%
Gold	Pay lower of 1.0 gram per tonne of concentrate deduction or 95%
Penalties
Fluorine	If less than 4,000 ppm, $1.50 per tonne concentrate, per 100 ppm, if greater than 4,000 pm, $50/tonne of concentrate

MPR has obtained metal price forecasts for the four payable metals from Los Gatos to include:

·Zinc

·Lead

·Silver

·Gold

The metal price forecasts were obtained from the following analysts:

·Barclays (Gold, Silver, Zinc: 16-Apr-20, n.a. for lead)

·Royal Bank of Canada (RBC) (Gold, Silver, Lead: 15-May-20, Zinc: 15-April-20)

·Macquarie Research (18-May-20)*

·United Bank of Switzerland (UBS) (18-May-20)*

·Cantor Fitzgerald (Gold: 12-May-20, Silver, Zinc, Lead: 8-May-20)

·Bank of Montreal (BMO) (29-May-20)

·Canaccord Genuity (20-May-20)

·Morgan Stanley (18-May-20)*

·TD Securities (Gold, Silver: 23-April-20, Zinc, Lead: 10-Jan-2020)

*Supplied by a third-party provider Energy & Metals Consensus.

Table 19-3 contains the average metal price forecasts for the payable metals for 2021 and the long-term forecast.

Table 19-3:  Payable Metal Price Forecasts for Los Gatos

Metal	Units	Consensus Long-Term Price
Gold	Ozs	1,472
Silver	Ozs	18.99
Lead	lb	0.87
Zinc	lb	1.09

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These metal prices were used in calculating the metal revenues in the economic analysis.

MPR has no lease agreements for the Los Gatos project.  Smelting and refining terms assumptions used in the Technical-Economic Model (TEM) are indicative of current rates and have been provided to Tetra Tech by MPR.

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20.0  ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1  Introduction

MPR retained ASI (Asesores en Impacto Ambiental y Seguridad, S.C.) to develop all the necessary studies to comply with the Mexican regulations related to the completion of a Manifestación de Impacto Ambiental (MIA — Environmental Impact Statement) and a Risk Study.  Tetra Tech has been involved in the process and provided technical support to ASI for the completion of this report.

The mining and processing operations at the Los Gatos Project are located approximately 120 km South of the state capital of Chihuahua City, and approximately 100 km North/Northwest of the historic mining city of Hidalgo del Parral.  The Project has been developed over an area of 390.7 Hectares, all located within surface lands owned or controlled by MPR.  This area includes mine operations, a processing plant, a tailings storage facility, and infrastructure to support the mine.  The purpose of the facility is to mine and process ore, with the final products being lead and zinc concentrates, as well as silver as a byproduct.

This Project is intended to be an ore and concentrate producer.  The ore will be mined from an orebody by underground mining methods over a period of 11 years (with an additional two years planned for onsite construction).  The ore mined from underground is processed onsite at a processing plant.

The processing plant has been designed to process 2,500 tonnes of sulfide ore per day, which is 912,500 tonnes per year.  Pb and Zn concentrates are produced, as well as a silver byproduct.  The tailings material from the plant is treated via the Inco cyanide destruction process, and then used for either cemented paste for backfill, or be transported to the tailings storage facility.

Waste rock from the mine is moved to a 4-hectare area authorized by the environmental authorities for this purpose.  A portion of the waste rock produced by mining operations is used for mine backfilling as a concrete rock fill (CRF).  The remaining waste rock is stockpiled on the surface before being used as construction material.

Necessary infrastructure will be built in the 390.37 hectares project area.

20.2  Regulations

The Secretary of Environment and Natural Resources (SEMARNAT) regulates the environmental aspects of mining projects, and issues permits once the EIS is approved, according to Art. 28, Frac. III, VII and X of the General Law of the Ecologic Equilibrium and Environmental Protection, and by the Art-5 Section L), Frac. I and III, Section O) Frac. I and Section R), Frac. I of the Regulations for Environmental Impact Assessment.

The Project is also regulated by the following regulations and legal orders:

·Political Constitution of the Mexican United States

·National Development Plan 2013-2018

·State Development Plan 2010-2016

·Satevó Municipality Development Plan 2013-2016

·General Territorial Ecology Program

·Important Hydrological Regions

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·General Law of the Ecologic Equilibrium and Environmental Protection (LGEEPA)

·Regulations to the LGEEPA for Environmental Impact Assessment

·Regulations to the LGEEPA for Prevention and Control of Atmospheric Contamination

·Regulations to the LGEEPA for Register of Emissions and Transfer of Contaminants

·General Law for Sustainable Forest Development

·Regulations to the General Law for Sustainable Forest Development

·General Law for the Wild Life

·General Law for Prevention and Integral Management of Residues

·Regulations to the General Law for Prevention and Integral Management of Residues

·Law of National Waters

·Regulations to the Law of National Waters

·Mining Law

·General Law of Climate Change

·Law of the National Bureau for Industrial Security and Environmental Protection of the Hydrocarbon Sector

Additionally, there are several Mexican Official Norms (NOMs) that apply to the Project:

·NOM-059-SEMARNAT-2010

·NOM-045- SEMARNAT-1996

·NOM-043-SEMARNAT-1993

·NOM-080-SEMARNAT-1994

·NOM-081-SEMARNAT-1994

·NOM-052-SEMARNAT-2005

·NOM-054-SEMARNAT-1993

·NOM-141-SEMARNAT-2003

·NOM-161-SEMARNAT-2011

·NOM-087-SEMARNAT-1995

·NOM-001-SEMARNAT-1996

·NOM-003-SEMARNAT-1997

·NOM-004-CNA-1996

·NOM-011-CNA-2000

·NOM-035-SEMARNAT-1993

·NOM-138-SEMARNAT/SS-2003

The Project, as designed, complies with these regulations and official norms related to environmental matters.

20.3  Baseline Studies

Before the start of a project, an environmental baseline has been established.  The purpose of baseline analysis is to characterize the components, deterioration processes and state of conservation under current, pre-operational conditions to serve as a basis for the identification of environmental impacts.  This allows for the establishment impact prevention and mitigation measures, and for generating a predictive model of future scenarios.

Additionally, with the environmental diagnosis, it is intended to identify any environmental problems in the area of influence of the project (AP).

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The Environmental System (ES) of the project has been defined by the micro basin of the San José river for reference, which covers an area of 20,225.98 ha (202,2598 km2).  The ES land categorization approach allows the analysis of components within a territorial system (Gómez Orea, 1999), associated with human activities, their distribution in space, and the regulation of their behavior.

Subsequent Permits and Modifications have been obtained by Los Gatos from SEMARNAT and other Federal, State and Municipal Authorities to maintain updated environmental permits (2018-2019) as follows:

SEMARNAT issued a modification to the authorized MIA-R approved under document SGPA/DGIRA/DG/05121-2017 for Cerro Los Gatos project mining exploitation in Satevó, Chihuahua, under official notification SGPA/DGIRA/DG/01914 dated on March 15, 2018 for approval of the environmental modification to the impacted area for execution of the project including 211.0841 ha to 268.8450 ha, as well as an increment from 95 to 99 mine workings areas.  This authorization expires on July 17, 2041.

Subsequently, on November 28, 2018, SEMARNAT issued a second modification authorization for the mining project in environmental matters through document SGPA/DGIRA/DG/09272, in which the impacted environmental area is incremented from 268.8450 ha to 325.086 ha, and the number of mine working areas is modified from 99 to 133.  This authorization expires on July 17, 2041.

SEMARNAT approved resolution for the exemption of presentation of MIA for the workings related to the expansion of the “Compacted dirt road San José del Sitio — Mina Los Gatos,” which was authorized under document No. SG.IR.08-2018/097.

SEMARNAT also issued on May 04, 2018 a resolution for modification of the trajectory for the “Power Line 115 KV Los Gatos,” which included detour within a distance of approximately 2 km (1.24 miles) from the original trajectory of the line, by resolution document No. SG.IR.08-2018/093.  This authorization expires on September 4, 2037.

A Preventive Report for Direct Mining Exploration including diamond drilling for the nominated zones “Los Gatos NW-CE-SE, Cascabel Fault and El Valle Vein” was approved by SEMARNAT on May 7, 2019 through document No. SG.IR.08-2019/070.  This permit does not have a set expiration date but will be concluded when the ecological conditions evaluated and stated in the study change.

CONAGUA issued concession N. 06CHI141265/24FADL16 for discharge of residual waters from the mine’s ramp, stopes and underground mine workings for deepening of the “Cerro Los Gatos” mine into the Santo Toribio creek.  This concession dated on August 31, 2018 by CONAGUA approves discharge of residual water including a volume of 8.0 liters per second.  This authorization expires on March 9, 2026.

CONAGUA issues resolutive document on July 16, 2019 for the application presented for modification of the concession No. 06CHI141265/24FADL16, which is based on the increment of the volume discharged to the Santo Toribio creek, from 8.0 to 120 l/s, and including the change of site for discharge to 500 m down drain from the original authorized point of discharge of the mentioned concession.  This authorization expires on March 9, 2026.

Cerro Los Gatos obtains from the Department “Gerencia de Aguas Superficiales e Ingeniería de Ríos, (GASIR)” which depends of CONAGUA the permit No. 4494 and dated on January 18, 2019, which authorizes the construction and operation of Tailings Storage Facility No. 1, which is projected for holding 7.6 Mm3 of mining tailings.  This permit includes approval for construction of the TSF in four stages, with a period for construction of 9 years.  This authorization expires on January 18, 2028.

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SEMARNAT issued on May 27, 2019 to the Los Gatos mining operation approval of the “Licencia Unica Ambiental (LAU)” (Unique Environmental License) under document No. LAU-CHIH-001-2019, which encompasses and approves installation and operation of the Los Gatos mining operation, with indefinite time limit considering an operating capacity of 2,500 tonnes per day.  This permit is unique for the authorized operating capacity and for the legal operator.  This authorization does not have a set expiration date, if the metallurgical process is not changed or the 2,500 tpd production capacity is not increased.

Los Gatos obtains authorization for the “Plan de Manejo de Residuos Peligrosos” (Management Plan for Handling Dangerous Residues) from the “Dirección de Gestión Integral de Materiales y Actividades Riesgosas” (DGGIMAR) a SEMARNAT agency, under document No.08-PMG-I-3405-2019, which registry validates and approves a total annual production of 144.642 tonnes of dangerous residues within the installations.  This permit does not have a set expiration date as long as the hazardous waste declared (144.642 tpy) is not increased.

Los Gatos obtained the registry as generating enterprise, as well as the Plan for Handling Residues for Special Handling on April 9 and 11, 2019 from the Secretaría de Desarrollo Urbano y Ecología (SEDUE) (Secretary of Urban Development and Ecology) from the State of Chihuahua.  This registry validates and approves that Los Gatos mining operation will generate 468.747 tonnes of residues annually, which require special management.  This permit does not have a set expiration date as long as the hazardous waste declared (468.747 tpy) is not increased.

During the months of June and July 2018 Los Gatos applied for and obtained from the “H. Ayuntamiento Municipal Constitucional de San Francisco Javier Satevó” (Municipality of Satevó) the following permits and approvals:

·Land use permit, which expires July 23, 2023;

·Authorization and approval for the initiation of construction of mining workings and infrastructure, which is indefinite; and

·Official alignment and number, which is indefinite.

20.4  Environmental Setting

The Project is located approximately 120 km South of the state capital of Chihuahua City.  The Project is located within the Sierra Madre Occidental Physiographic Province, particularly within the Sub-provinces of the Great Plateau and Canyons of Chihuahua, and Sierras and Plains of Durango, which cover the greater part of the ES.

20.4.1  Climate and Precipitation

Within the ES defined by the micro basin of the San José river, the present climate group is defined as the dry climates “B” and semi-dry climates “BS1”, and the climatic subtypes semi-arid semi-warm and mild semidry.  The area has an average temperature of 17.5°C and an average rainfall of 433.2 mm according to the nearby climatological stations.

The project climate is relevant, as prior to the execution of the project, and area of 390.37 hectares occupied by desert scrub microphyll was cleared.  This clearing, together with the emission of suspended particles and greenhouse gases, can contribute to the impact of the micro-climate of the environmental system where the project is located.

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20.4.2  Air Quality and Atmosphere

Within the ES, no fixed sources of pollutant emissions were identified.  This system belongs to the municipality of Satevó and is considered as rural according to the municipal classification and the size of its villages.  San José del Sitio, with a population of 264 inhabitants, is the most important population for the ES.  Due to the low number of inhabitants, the presence of mobile sources of emissions are quite limited.

According to the National Institute of Statistics and Geography in Mexico (INEGI), most of the area is covered by natural grass shrub vegetation; however, according to observations made in the field, there is also the presence of introduced grass for cattle grazing.  Livestock is the preponderant activity within the ES.

Raising livestock is a greenhouse gas producing activity, mainly for nitrous oxide, methane and carbon dioxide, and is considered that the ES does not present an optimal state of conservation.

Dust is produced by vehicles and machinery on access roads and internal roads of the Project.  Ore crushing, dismantling and stripping, and hauling of borrow material for the construction of the mining infrastructure, including the tailings storage facility, contributes to site-wide dust.  This may affect the atmospheric quality of the environmental system.

20.4.3  Geology

Geologically, the ES is composed of a range of seven surficial lithological units plus soils, although only five of them cover 92.05% of the total surface area of the ES and the other two lithological units together cover only the 0.89%, the remaining component is alluvial that cover 7.06%.  The most important are the conglomerates, andesite, acid rhyolite-tuff and sandstone-conglomerate intercalations plus granodiorite.

20.4.4  Soils

Edaphically, the ES is represented by different soil types such as Leptosols, Chernozem, Luvisol, Fluvisol, Kastañozem and Regosol; however, according to the fieldwork carried out, it was found that the natural grassland has been affected by the establishment of introduced grass for cattle breeding that is carried out extensively.

Cattle grazing generates soil compaction and eventually soil erosion.  According to the estimates made, the ES has an average erosion of 10.05979 t/hectares/year, which is considered light, although there are areas with moderate erosion.  While the area of the project an average erosion of 7 t/hectares/year was estimated, and it is also considered light.

20.4.5  Regional and Site Hydrology

20.4.5.1  Hydrology

The ES is part of the Bravo-Conchos RH-24, located particularly within the Conchos-Presa de la Boquilla basin and the Río Conchos-Valley sub-basin of Zaragoza, with the Rio Conchos basin contributing International waters, which gives this component of the ES a special importance.

Within the ES, there are series of intermittent water runoffs that originate the Santo Toribio, El Yeso and El Salto streams, indicated as permanent streams in the INEGI cartography.  However, these streams are

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typically dry during the dry season due to the reduced volume of precipitation in the area and the geology, which favors infiltration.

According to the estimates, the following parameters are included in the ES a precipitation volume of 100,651 m3/year, with an evapotranspiration of 81,551 hm3/year, and a drained volume of 13,531 hm3/year, which means an infiltration of 5,564 hm3/year.

The discharge of industrial effluents (that could contribute to high levels of contaminants to the surface water) were not identified within the ES, since this is a rural area where irrigation agriculture is very limited, and the discharge of municipal waters is limited to the discharge of the drainage of the locality of San José del Sitio, located near the limits of the micro basin of the San José river.

The potential impact generated by the Project refers to the risk of water pollution derived from Project actions, without the application of mitigation measures; the modification of the runoff in the area of influence of the project, mainly by the construction of the TSF, and the alteration of the runoff and infiltration patterns of the rainwater by the installation of the mining infrastructure.

20.4.5.2  Hydrogeology

The ES is part of the aquifers identified as San Pedro High River and San Felipe de Jesús, which covers most of the ES area.  According to CONAGUA, the San Felipe de Jesús aquifer presents a deficit of -0.045088 Mm3 of water per year.

According to studies conducted on site, a large amount of groundwater resource has been found to be at two levels:  perched zones, which are isolated and disconnected from the main aquifer, normally appear from 3 m deep, and the “San Felipe de Jesús” aquifer located starting at depths between 67 and 245 m.

The San Felipe de Jesús aquifer is not classified as vulnerable to contamination based on the Norm Annex 2 of NOM-141-SEMARNAT-2003, which describes the method to evaluate the vulnerability of an aquifer to pollution.

Based on chemical analysis of water samples from dewatering and monitoring wells in the area, groundwater quality does not exceed any of the standards established for water use or water discharge, with the exception of total coliforms in samples from most wells, fecal coliforms in samples from two wells, and total trihalomethanes in samples from three wells.  The water has neutral pH, low to moderate total dissolved solids, and does not exceed any of the other regulated compounds.

According to the database “Public Register of Water Rights” (REPDA) of the National Water Commission (CONAGUA 2016), the nearest uses identified by the REPDA that were located in the micro basin are one for groundwater approximately 3 km from the project area and one for surface water, just upstream of the project area.

The planned impacts to this resource refer to the extraction of groundwater by pumping of groundwater intercepted in the mine and mine dewatering wells to allow for mining activities, which is sent to the tailing storage facility.  Part of this water is used for processing, which is recirculated in closed circuit and some is used in camps and offices.

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20.4.6  Flora

The biotic environment of the ES, according to INEGI (2015), is made up of three components. 93.93% of the area is covered by a primary natural pasture and secondary shrub vegetation, 4.48% is covered by agricultural and livestock activities and only 1.59% of the total area is covered by a secondary shrub vegetation of oak forest.

According to the field work, the presence of a microphytic desert scrubland was found within the Project area (PA), and the corresponding inventory was made both for the area affected by the Project and in the ecosystem located within the ES.

According to the inventory, 95 species of flora within the ES and 84 species in the PA were identified, being that the microphyllous desert scrubland has a great similarity between the ES and the PA, especially since in the two areas only one species was listed in the NOM-059-SEMARNAT-2010, under the category of threatened, being the walnut Juglans major.

Modifications to the vegetation is one of the areas where potential environmental impacts have been identified.  An area of 390.37 hectares will be modified, which is currently covered by a microphyllous desert shrubs and will be affected by the construction of the mining infrastructure and mining and Mineral processing and/or other activities.

Regarding the state of conservation of the PA, the vegetation generally does not correspond to pristine ecosystems due to extensive livestock activity and for the establishment of more productive introduced grasses to support cattle production.

20.4.7  Fauna

Inventories of fauna within the PA and ES were conducted. Based on these inventories, 14 species of herpetofauna; 99 species of avifauna; 36 species of mammal fauna; and 9 species of ichthyofaunal were observed in the ES, 21 species of which are listed in the NOM-059-SEMARNAT-2010, while in the PA 88 species were identified: 10 of herpetofauna, 46 of avifauna, 26 of mammal fauna and 6 of ichthyofauna.

The wildlife is another component of the ES, which has some environmental impacts.  The project will affect an area of 390.37 hectares covered by a microphyllous desert scrubland.  This could cause the migration of fauna to more protected sites and could cause loss of habitat.  To minimize this potential impact, mitigation efforts will be directed towards this resource.

20.4.8  Socioeconomic

The ES is located totally within the municipality of Satevó, state of Chihuahua, finding that within the ES there are 16 small villages.

It has identified a population of 452 people within the ES, noting that the data provided are approximate, since the mobility of the population is constant and the most recent official records available are from the year 2010.

The activities carried out within the ES are the breeding of livestock, agriculture and recently mining due to the execution of the mining exploration project that has been carried out since 2009 by the company Minera Plata Real, S. de R.L. of C.V., identified as “Los Gatos”.

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20.5  Identification, Description, and Assessment of Environmental Impacts

It is estimated that the Project will generally generate a series of potential project-environment interactions of a negative and positive nature based on data obtained from the environmental studies, the opinion of experts, and the various environmental impact assessment techniques used.

Of the 589 interactions identified, 359 correspond to negative impacts, which means 60.95% of the total, while 230 were identified as positive interactions, representing 39.05%.

Of the negative interactions, it was found that most of the components 102 (28.41%), were related to the atmosphere, 71 (19.78%) socioeconomics, 65 (18.11%) water, and 37 (10.31%) soil and ecosystem factors.  The other negative interactions (23.39%) are found in geomorphology, flora and fauna.

In relation to the project stages, it was found that of the 359 negative interactions, 153 (42.62%) will occur during the construction stage; 100 (27.86%) in the operation stage; 73 (20.33%) in the preparation stage of the site; and 33 (9.19%) in the closure or post-operation stage of the project.

Regarding the positive impacts, the majority of the 168 (73.04%) are related to the socio-economic factor, and the remaining 26.96% in other components of the ecosystem.

As for the project stages, of the 230 positive interactions, 20 (8.70%) will occur during the site preparation stage; 52 (22.61%) during the construction stage; 53 (23.04%) during the operation stage; and 105 (45.65%) in the closure or post-operation stage.

Of the 589 interactions found in the matrix of identification of potential project-environment interactions, the list was narrowed down to 42 potential impacts based on the Project and its interactions with the environment.  Of the 42 potential environmental impacts, 24 correspond to negative impacts and 18 to positive impacts.

Examining the results of the analyses, the environmental impacts were identified, determining which are significant or relevant without the application of mitigation measures.  The impacts identified correspond to water pollution and the loss of individuals of flora and fauna in conservation status.  These impacts can be prevented and mitigated, making the Project, in environmental terms, viable in all of its stages.

The project complies with the provisions of Article 35 of the LGEEPA according to the identification and evaluation of impacts presented.  The evaluation shows the possible effects of the project activities will not put the structure and function of the micro basin of the San José River delimited as the PA or the ES at risk.  This is based on the analysis of possible interactions that the Project might have with environmental components and processes of the environmental system at different geographic scales.

In this context, it was analyzed and concluded that:

·There are processes whose occurrence is greater than the ES delimited by the micro basin of the San José River and are named supra-regional, such as the climate, the geological structure and the hydrological cycle.  Consequently, the project does not generate effects that could alter these macro processes.

·Several components with a certain degree of importance from an environmental perspective were identified, such as water and vegetation cover with species of flora and fauna listed in NOM-059-SEMARNAT-2010.  Although the vegetation of the impacted area has some ecological importance, this type of vegetation is not protected by any specific norm.  The project does not affect the existence of such vegetation or the integrity of the ecosystem,

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since no species are compromised because the areas of distribution is greater than the surface required by the Project and the ES itself.  In addition, the individual species of concern will be relocated and/or replaced as appropriate.

·Interactions between project activities and environmental components and processes were identified as potential environmental impacts.  In particular, the lowering of the ecosystem quality, the loss of natural components of the ecosystem, soil impermeability, and alteration of water runoff and infiltration patterns were recognized as impacts with a high incidence rate, but were not considered relevant, since the impacts above are only represented on 1.93% of the surface of the ES.

The impact assessment is compliant with the approach requested by the LGEEPA and the REIA as detailed below:

·Qualify the effect of impacts on the ecosystems, with regards to the relevance of possible effects on their functionality (Article 44, section II of the REIA).

·Develop this qualification in context of an ES and an area of influence of the project (Article 12, section IV of the REIA), so that the evaluation refers to both the ES and the PA where the project is intended to be located.

It was found, by the regulations established by the REIA, the ES integrated by the micro basin of the San José River will generate a non-relevant impact for the removal of 390.37 hectares of vegetation.  In comparison to the surface area of the ES, these areas represent only 1.93% of the total area, showing the impact is not significant.  This ensures the function or continuity of ecosystem processes in the environmental system is not affected.

The approach to maintaining the integrity of the ecosystems present in the ES during and following mining and Mineral processing is to reduce and avoid impacts that eliminate habitats and/or species and preserve the conditions that allow the mobility and the viability of species.  By understanding the load capacity of an ecosystem, such as the capacity to be used or managed without compromising its basic structure and operation, the design of the project ensures these two conditions.

By understanding the capacity of an ecosystem to tolerate or adapt to stressors, such as the capacity to be used or managed without compromising its basic structure and operation, the design of the Project ensures these two conditions are satisfied.

The conclusions of the environmental impact assessment indicate that the functional integrity of the ecosystems is respected, since the relevant environmental components will not be significantly affected.  In the case of species under some category of risk, their areas of distribution are greater than the ES.  For water pollution, considered as a relevant impact, it is not planned to discharge process or mining water to natural effluents, even though it is considered to implement water treatment practices to be discharged to the tailings dam, which is built with liner and geo membranes; plus, the recirculation of process water.

The level of conservation of regional biodiversity demonstrates that the project will not cause any species to be declared as threatened or endangered, and the habitat of individuals of flora and fauna will not affect the species, according to Article 35-III, subsection b) of the LGEEPA.

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Finally, as a result of the above conclusions the project will not generate significant and/or relevant effects to the ES, such as:

·Ecological imbalances;

·Damage to public health;

·Irreversible effects to the ecosystems of the ES.

20.6  Preventive and Mitigation Measures for the Environmental Impacts

The set of measures designed for the prevention and mitigation of environmental impacts is included within the Environmental Management System (EMS), as part of the company’s commitments to prevent and mitigate the environmental impacts of the Project.

The objective of the EMS is to provide the basis for monitoring and adopt control mechanisms to ensure the Project’s development and associated activities, as well as prevention and mitigation, are in compliance with the criteria of sustainability and environmental protection as indicated in the environmental norm.

The program, considered to be a guiding instrument for the company, includes the descriptive information necessary to guide the work of the supervision team assigned to the Project.

Therefore, the EMS for the Project allows the company to execute an integrated approach to address and mitigate negative effects of the Project to the environment meeting the following central objectives:

·Implement the impact management measures to prevent and mitigate the possible effects derived from the potential environmental impacts expected in each stage of the Project in compliance with the precepts established in Article 30 of the LGEEPA and 12, section VI of the REIA.

·Propose technically and economically feasible actions, in a way that the implementation can be monitored through an environmental monitoring program.

·Implement actions that allow attention and strict compliance with the terms and conditions established by SEMARNAT in the REIA.

·Identify and implement monitoring methods capable of illuminating impact mechanisms and indicators and evaluating, verifying and documenting compliance with and effectiveness of the adopted environmental measures.

·Ensure that, in relation to the environment, each activity during the Project life is carried out according to the plans and approaches authorized by the SEMARNAT.

·Determine the effectiveness of the environmental protection measures that have been proposed and, if necessary, correct them.

·Properly apply the methods of registering and documenting actions to validate compliance.

·Identify, report, adjust, and correct any deviations in Project development or implementation of environmental measures.

·Manage, in a timely manner, the necessary financial resources for the implementation of environmental measures and ensure their timely availability.

To comply with the guidelines established in article 30 of the LGEEPA, the following specific objectives are presented.

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20.7  Environmental Forecast

The environmental forecast for the micro basin of the San José River delimited as an ES for the Project, is the result of the application of the KSIM Environmental Quality Change Simulation Model and is described below.

20.7.1  Atmosphere

The air quality in the ES is predicted not to vary significantly for a period of 25 years, considering the deterioration trends without the Project compared to the insertion of the Project and mitigation measures.

20.7.2  Soil

The soil is an environmental component that is altered within the ES, mainly by the change of land use in forest lands, for the construction of mining infrastructure.  However, the proposed mitigation measures are effective to minimize the impact.

20.7.3  Water

Water is an indispensable element for the development of a mining project, and an important component of the ES that is affected by the execution of the Project, however, according to the Simulation Model KSIM, the proposed mitigation measures reduce the level of impact.

20.7.4  Geology and Geomorphology

Ore extraction, construction of the tailing storage facility and borrow material areas have a direct effect to the ES, even with the application of the proposed mitigation measures.  According to the simulation, the negative impact is low, and the mitigation measures contribute to reduce modifications to the ES.

20.7.5  Flora

With regard to vegetation, the project focuses on areas of native vegetation, particularly on a microphyllous desert scrubland, so according to the KSIM Simulation Model, it indicates the existence of partial modifications on this component, nevertheless, its tendency and environmental quality and its effects are controllable through the practice of minimizing the affected area or increasing forestation areas.

20.7.6  Fauna

A series of potential negative interactions were identified involving fauna.  The proposed mitigation measures allow a minimum impact to fauna, and therefore does not require additional measures to those proposed within the EMS.

20.7.7  Ecosystem

The impact on the ecosystem is negative, but it has a great similarity with the evolution of their natural characteristics or attributes, where the actions of the evaluated project do not produce substantial

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modifications at the level of ES.  Therefore, no additional measures of attenuation to the environmental are expected.

20.7.8  Socioeconomic

This component has both potentially negative and positive interactions due to the demand for goods and services, generation of jobs, and a change in the use of the territory.

In summary, it is expected that the implementation of the project will have no significant negative impact in the ES delimited by the San José River basin, but with the execution of the project, the positive impacts will support the development of mining in the state of Chihuahua.

20.8  Closure Plan

The Los Gatos Project (Project) is subject to laws and regulatory requirements related, either directly or indirectly, to the closure and reclamation of mines in Mexico, as well as the use and protection of land, and water and wildlife resources.  Closure of the Project and reclamation of surface disturbances attributed to the Project must be consistent with the provisions of the Mining Rights of Public Lands Under Article 27 of the Mexican Constitution (i.e., the government may establish compulsory measures for the preservation and restoration of land and ecological equilibrium), Article 98 (i.e., land reclamation should consider all the necessary actions to prevent land degradation), and Article 35 (i.e., reclamation conditions defined in the approved Manifestación de Impacto Ambiental — MIA, or Environmental Impact Statement, are binding) and the General Law of Ecological Equilibrium and the Protection of the Environment.  These laws establish that holders of mining concessions in Mexico are subject to compliance with all Mexican general regulations and Normas Oficiales Mexicanas (NOMs) that relate to ecological equilibrium and environmental protection, in particular the NOMs for wastewater discharge into national waters and resources (i.e., NOM-001-SEMARNAT {Secretaría de Medio Ambiente y Recursos Naturales}-1996), remediation of hydrocarbon-contaminated soils (i.e., NOM-138-SEMARNAT/SSA-2003), solid waste and landfills (i.e., NOM-083-SEMARNAT-2003), and tailings and mine waste management and closure (i.e., NOM-141-SEMARNAT-2003 and 157-SEMARNAT-2009, respectively).

The primary governmental agencies with direct or indirect regulatory authority over mine closure and reclamation include:

·Procuraduría Federal de Protección al Ambiente (PROFEPA, Office of the Federal Attorney General for Environmental Protection)

·SEMARNAT (Secretariat of Environment and Natural Resources)

·Comisión Nacional del Agua (CONAGUA, National Water Commission)

Water regulation, including the use of the nation’s water or the right to discharge wastewater, is carried out by CONAGUA.  Whereas land use is handled by local agencies in charge of the zoning and registration of land ownership, as well as SEMARNAT delegations, responsible for issuing land-use change permits for projects that will involve alteration of forested areas.  PROFEPA is the agency responsible for enforcing SEMARNAT regulations.  PROFEPA’s main activity is to deal with complaints, conduct inspections and, in general, verify compliance with all federal environmental laws and regulations.  It imposes penalties for violations of environmental laws and regulation and monitors compliance with any preventive and mitigating measures issued by it.  PROFEPA also conducts environmental audits.  To our knowledge MPR has not experienced significant conflicts with these agencies listed above as it related to Project

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development, environmental permitting and compliance, and mine closure and reclamation planning and design.

Pursuant to Chapter VIII of the Ley General del Equilibrio Ecológico y la Protección al Ambiente (LGEEPA) regulations, insurance and bonds are considered (but not specifically required) for closure and reclamation activities.  However, in recent mine closure reviews, LGEEPA authorities have been more diligent regarding bond requirements for mining companies (Tetra Tech, 2015).

Tetra Tech was retained by MPR, to develop a Feasibility Closure and Reclamation Plan (CRP) and closure cost estimate for the Project in support of this Feasibility Study NI 43-101 Technical Report (FS) for mining operations.  Tetra Tech also prepared a conceptual closure and reclamation plan for the MIA or Environmental Impact Statement for the Project.  Tetra Tech is not aware of other published closure and reclamation plans for the Project.

The CRP includes approaches and plans to address the closure and reclamation of Project-related disturbance in accordance with regulatory requirements discussed above and sound scientific and engineering practices, and industry-standard practices for mine closure and reclamation.  The CRP was developed by Tetra Tech based on input, designs, and analyses from key technical experts and design teams involved with the planning and design of mining, process, waste and fluids management, and operating systems for the FS and MIA.

The closure and reclamation activities described in the CRP are anticipated to be modified as more information becomes available.  As such, mine closure and reclamation planning will be dynamic and adapted as mine plans and site conditions change through the mine pre-production and production periods.  Closure and reclamation activities are planned to be completed following to cessation of mining and ore processing; however, opportunities may arise during the production period to concurrently reclaim facilities that are no longer necessary for mining (e.g., borrow areas, temporary camp, temporary rock storage).

Prior to the initiation of mining activities, existing vegetation was cleared and soils that are suitable for reclamation that can be practicably salvaged using heavy equipment have been salvaged from planned surface disturbance and stockpiled.

In general, grading has been completed to blend disturbed areas (with the exception of the TSF) into the surrounding topography and to generally re-establish the previous or natural drainage patterns to convey surface water towards the Northeast and the Santo Toribio Arroyo.  The site has been regraded to the degree practicable using storm water drainage controls constructed and maintained during the pre-production and production period.  At closure, most culverts will be removed or buried, and drainage re-established.  Riprap or other armoring methods are anticipated to be necessary to limit scour and head cutting along portions of re-established drainages.

Prior to and throughout the duration of disturbance activities, erosion and sediment control Best Management Practices (BMPs) will be installed, monitored and maintained.  These BMPS may include but are not limited to: chemical soil stabilizers, wind fences, tillage furrows, mulch, silt fences, erosion control blanket, check dams, coir log, etc.  The effectiveness of these BMPs have been evaluated on areas reclaimed on an interim and concurrent basis (if any).  In addition, the proliferation of noxious and invasive weeds is controlled.

Post-closure monitoring and maintenance will provide assurance that the reclaimed Project-related disturbance meet the Project closure and reclamation goals.

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The planned closure and reclamation activities for major Project facilities are briefly described below.

20.8.1  Tailings Storage Facility

At closure, residual process water, process circuit rinsate, rinsate from tank decontamination and contact water in the TSF seepage collection basins will be pumped to the TSF and drain to its supernatant pond.  The mine dewatering pumping and cooling system is anticipated to continue operation until underground equipment, fuels and chemicals stored underground are removed, which is anticipated to be completed in the first year of the closure period.  Therefore, during this time mine water is anticipated to be pumped to the TSF as well.  As such, the TSF supernatant pond following the production period is anticipated to be a combination of the following: mine water; tailings slurry and seepage collected in over- and under- drains; dilute and mixtures of reagents and chemicals potentially remaining within the process circuit following cyanide destruction; rinse solution produced from the decontamination of processing equipment and tanks; and rainwater and runoff from the limited catchment area below the TSF diversions.  At this time, it is anticipated the water quality of the TSF supernatant pond following the production and closure periods will not meet applicable NOM-001 standards for discharge of wastewater into national waters and resources.

Following termination of tailings and mine water deposition in the TSF, the mine water, tailings and reclaim pumps, pipelines, spigots and barge will be removed and salvaged or disposed.  The run-on diversion above the TSF will remain in place to protect the TSF from the PMP and lesser events.  The downstream slope of the TSF embankment will be armored with durable rock to limit erosion of the embankment.  The application of plant growth medium to and revegetation of the impounded surface of the TSF is anticipated to occur through the closure and post-closure period as the supernatant pond passively evaporates and tailings consolidate, allowing equipment access.  Passive evaporation of most of the supernatant pond is anticipated to take 30 to 40 years following termination of tailings and mine water deposition in the TSF.

Surface drainage features are anticipated to be constructed on the impounded surface of the TSF (if necessary) to control runoff conveyance towards the Southeastern side of TSF.  A spillway channel will be constructed for the TSF at or following the closure period.  Following reclamation of the TSF, a pond may persist or temporarily form on the impounded surface of the reclaimed TSF following storm events.

Salts and other constituents dissolved in the TSF supernatant pond will deposit on the tailings surface and concentrate in pond water as evaporation occurs, while rain incident to the impounded surface of the TSF and runoff from the limited catchment area below the TSF diversion will dilute the supernatant pond.  If necessary, mitigation measures to address the ingestion of supernatant pond water by livestock and wildlife and control the capillary rise of salts into the plant growth medium covering the TSF will be developed and executed.

The TSF seepage will be collected in the seepage collection system and pumped back into the TSF supernatant pond.  The TSF seepage collection and pumping system will be sealed (if appropriate), decommission and removed and reclaimed when seepage volumes are de minimus or seepage quality is demonstrated to meet applicable NOM-001 standards for the discharge of wastewater into national waters and resources.  The TSF seepage collection and pumping system is assumed to remain through the closure and post-closure periods for approximately 35 years following termination of the tailings and mine water deposition.

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20.8.2  Temporary Rock Storage

Following removal of mine rock, the Temporary Rock Storage (TRS) will be graded to blend into the surrounding topography.  The area will be graded to promote surface water drainage towards engineered armored channels designed to convey the 100-year peak storm flow.  Plant growth medium (PGM) will be applied to the surface and the area will be revegetated.

20.8.3  Underground Mine and Openings

Underground equipment, fuels and chemicals stored underground will be removed from the mine prior to closure of the decline and ventilation shafts. Inert debris from the demolition of site-wide facilities (see below) that is not salvaged as scrap will be used to backfill the decline to above the predicted steady-state groundwater elevation in the decline following closure.  The portal will be sealed to prevent access.  Plant growth medium will be applied to the portal area and the area will revegetated.

Closure of the ventilation shafts is anticipated to include a reinforced concrete slab with a minimum of 3 meters of backfill on top of the slab, followed by application of PGM and implementation of revegetation activities.

20.8.4  Process and Ancillary Facilities

Process and ancillary buildings and facilities will be decommissioned, demolished and disposed of onsite in the decline and primary crusher area or in off-site solid waste landfills.  Some ancillary buildings and facilities are anticipated to remain open for four to five years following cessation of mining activities to support final closure and reclamation activities on the site.  Some of the processing and ancillary facilities and buildings may remain to support an industrial PMLU, if viable.  Following facility decommissioning, demolition and removal, concrete foundations will be broken up and covered in place.  Building and facility yards will be graded to blend into the surrounding topography and to generally re-establish the existing drainage patterns, which conveys surface water towards the Northeast.  PGM will be applied to the surface of the regraded process and ancillary facilities and the area will be revegetated.

20.8.5  Borrow Areas and Exploration Camp

Closure and reclamation of the borrow areas and the exploration camp and associated infrastructure is anticipated to occur following the production period, however, opportunities may exist to close and reclaim these facilities during the production period.

20.8.6  Seepage, Stormwater, Sedimentation and Cooling Basins

When final reclamation and water quality criteria are satisfied, accumulated sediments in the bottom of seepage (contact water) collection, stormwater, sedimentation and cooling basins will be removed, if necessary, and the basin area will be reclaimed.

20.8.7  Road and Conveyor Corridors

Most of the Project road and conveyor corridors within the Project area will be reclaimed during the closure period unless needed to support an industrial PMLU, if viable.  Select ancillary roads will be needed for site monitoring and maintenance until final reclamation standards are achieved.

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Road and conveyor corridors that will be reclaimed will be deep ripped to reduce compaction.  Grading will be limited to the main access road and where regrading may be needed to re-established drainages.

20.8.8  Utility and Pipeline Corridor

The surface disturbance associated with the installation of utilities (e.g., electric, natural gas, water and sewer) will be reclaimed on an interim basis during the production period.  Electrical utilities (i.e., substations and powerlines) are assumed to be valuable to the local community following the closure period and will therefore be de-energized and remain in place.  Surface and large shallow pipes or culverts will be removed and disposed.  At closure power and pipeline corridors will be deep-ripped and revegetated.

20.8.9  Well Abandonment

Monitoring and mine dewatering wells will be abandoned in accordance with applicable rules and regulations when no longer needed to prevent cross contamination between aquifers, and the required shallow seal will be placed to prevent contamination by surface access.

The estimated cost for closure and reclamation of the Project-related surface disturbance and facilities is presented with a 6% contingency in Table 20-1.

Table 20-1:  Summary of Closure and Reclamation Costs

Cost Category	Cost Estimate (USD)
Earthworks and Recontouring	$	2,920,000
Revegetation and Stabilization	$	1,222,000
Detoxification/Water Treatment and Disposal of Waste	$	2,240,000
Structure, Equipment and Facility Removal, and Miscellaneous Cost	$	4,915,000
Monitoring and Maintenance	$	358,000
Construction Management and Support	$	596,000
Indirect Costs	$	2,631,000
Total	$	14,884,000

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21.0  CAPITAL AND OPERATING COSTS

Project economics are based primarily on inputs developed by MPR, Tetra Tech, SGS, and Stantec, as well as current operating conditions.  All costs and economic results are presented in constant 2020 U.S. dollars.  Project economics using conventional pro forma cash flow are presented in this report indicate the following results, assuming no gearing (100% equity): Mine Life: 11 years;

·Pre-tax present value (PV5.0%): $764.7 million;

·Post-tax present value (PV5.0%): $653.2 million;

·Taxes Paid: $148.2 million;

·Sustaining project capital is $267.3 million; and

·Initial project capital was $315.7 million.

Quantities and values are presented using standard metric units unless otherwise specified.  No escalation has been applied metal prices, to capital or operating costs.  No gearing is assumed in the analysis.   Cash flows are discounted on an end-of-year basis.

Technical economic tables and figures presented require subsequent calculations to derive subtotals, totals, and weighted averages.  Such calculations inherently involve a degree of rounding.  Where these occur, they are not considered to be material.

21.1  Principal Assumptions

Parameters used in the analysis are shown in Table 21-1.  These parameters are based upon current market conditions, quotes, and benchmarks against similar existing projects.  The construction schedule accounts for 30 months of pre-production activities.  All pre-production tasks have been executed on time, and the cost is included in the current Technical-Economic Model (TEM) to keep the same economic parameters from the original FS.  The Los Gatos Project (the Project) operations are planned to operate at a sustained rate of 2,500 tonnes of ore per day (tpd) after first quarter 2021.  Mineral Resources that are not Mineral Reserves have not demonstrated economic viability.

Table 21-1:  TEM Principal Assumptions

Description	Parameter	Unit
General Assumptions
Mine Life	11	years
Operating Days	360	days/yr
Ore Production	2,500	tpd
Payable Metals (Life of Mine)
Zinc Concentrate
Zinc	678,737	klb
Silver	7,032	koz
Gold	0	koz
Lead Concentrate
Lead	441,755	klb
Silver	64,930	koz
Gold	46	koz

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Description	Parameter	Unit
Metal Assumptions (Long-Term)
Silver Price	$18.99	$/oz
Gold Price	$1,472	$/oz
Zinc Price	$1.09	$/lb
Lead Price	$0.87	$/lb
Financial Assumptions
Royalty	2.0% / 0.5%
Federal Tax	30%
Mining Tax	7.5%
Gearing	None
Technical Assumptions
Diesel	MXN16.81	MXN/L
Gasoline	MXN16.06	MXN/L
Power	MXN2.14	MXN/kWh
Process Recoveries
Zinc Concentrate
Silver	12.7%
Gold	10.2%
Zinc	66.6%
Lead	5.6%
Lead Concentrate
Silver	72.5%
Gold	53.7%
Zinc	6.6%
Lead	79.6%

Projected revenues from the sale of silver, gold, zinc, and lead are based upon long term consensus prices of $18.99/oz Ag, $1,472/oz Au, $1.09/lb Zn, and $0.87/lb Pb respectively.  Smelter assumptions used in the TEM are indicative of current rates and have been provided to Tetra Tech by MPR and are detailed in Section 19.

Treatment (smelting) and refining charges are $299.75/tonne-concentrate and $175/tonne-concentrate for zinc and lead concentrate, respectively.  Smelter charges also include penalties for fluorine, averaging $12/tonne concentrate for zinc and $50/tonne concentrate for lead.  Freight and insurance costs to the smelting plants average $11.04/tonne concentrate and $2.90/tonne concentrate for zinc and lead concentrate, respectively.

The Project is subject to a:

·30% federal tax;

·7.5% Mexican special mining tax; and

·The La Cuesta Royalty which is generally paid at increments of $40 thousand per year during pre-production, 2.0% up to $10 million, and 0.5% up to $15 million ($14.415 million is the remaining balance as of June 30, 2020).

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The following exchange rates are used for the Project:

·22.74 Mexican Pesos to 1 United States Dollar (USD);

·1.36 Canadian Dollars to 1 USD;

·0.89 Euros to 1 USD; and

·107 Japanese Yen to 1 USD.

Metallurgical test work supports the process recovery rates presented in Table 21-1.

21.2  Life of Mine

21.2.1  Underground Mining

Underground production will have a 11-year life of mine (LOM), generating 9,618 kt-ore.  Production over the LOM is summarized in Table 21-2.  Details describing the underground operations are in Section 16.

Table 21-2:  LOM Production from July 1, 2020

Description	Value	Unit
Development
Lateral Development	27,666	m
Vertical Development	2,697	m
Waste
Waste tonnes	1,446	kt
Ore
Development Ore	339	kt
Production Ore	9,279	kt
Total Ore	9,618	kt
Grade
Silver	305	g/t
Gold	0.35	g/t
Zinc	5.7%	%
Lead	2.8%	%
Contained Metal
Silver	94,461	koz
Gold	108	koz
Zinc	1,198,959	klb
Lead	585,343	klb

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21.2.2  Processing

The Project process facility has been designed to treat a nominal 2,500 tpd of ore.  The LOM mill feed is shown in Table 21-3.  Over the project life, 80 million ounces of silver, 69 thousand ounces of gold, 878 million pounds of zinc, and 499 million pounds of lead are produced in the concentrate.  Details describing the process operations are in Section 13.

Table 21-3:  LOM Mill Feed from July 1, 2019

Description	Value	Unit
ROM Feed	9,618	kt
Contained Metal
Silver	94,461	koz
Gold	98	koz
Zinc	1,198,959	klb
Lead	585,343	klb
Concentrate Produced
Zinc	650	kt-dry
Lead	346	kt-dry
Concentrate Grade Produced
Zinc	55.7%	—
Lead	61.0%	—
Metals in Concentrate (LOM)
Silver	80,481	koz
Gold	69	koz
Zinc	877,638	klb
Lead	498,713	klb

21.3  Capital Costs

LOM capital cost (including sustaining capital) requirements are estimated at $267 million as summarized in Table 21-4.  Initial capital of $316 million, was spent to construct the project and commence operations. The Project construction was completed on time and on budget.

The overall cost estimate includes sustaining capital and mine development costs.  Capital cost estimates utilized MPR-provided database information for:

·Mine and Surface Infrastructure

·Process Plant and Infrastructure

·Tailings Storage Facility

·Waste Rock Storage Facility

·Water Management

·Environmental

·Reclamation

·Owner’s Costs

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Table 21-4:  LOM Capital Costs

Description	Units		Sustaining Capital
Direct Costs
Mine & Surface Infrastructure	$	000s	266,398
Direct Costs	$	000s	266,398
Indirect Costs
Mine & Surface Infrastructure	$	000s	932
Indirect Costs	$	000s	932
Total Sustaining Capital	$	000s	267,330

Based on the available information used in the report, under Association for the Advancement of Cost Engineering (AACE) guidelines, the current study is considered a Class 2 estimate.  The accuracy is estimated to be (±15%), and for the mine in production, 0% contingency was included in the estimate.

21.4  Operating Costs

Operating costs are presented as a summary.  LOM operating costs were developed from current operating conditions and are summarized in Table 21-5.  Operating costs are estimated to be $83.58/t-milled over the LOM.

Table 21-5:  Mining Operating Costs

Description	Unit Cost ($/t-ore mined)	Unit Cost ($/t-milled)
Mining Operating	83.58	83.58

The mine operating cost estimate includes all sill development, stope production, and drift-and-fill production and associated indirects.  The owner performed all preproduction and production ore operating work.  Operating costs include all labor, material, mobile and fixed equipment operating, and power consumption costs.  The accuracy of these operating costs is estimated to be within +/- 15%.

Costs have been calculated by year, based on the development and production schedule and the surface and underground infrastructure designs.

The operating costs were combined for inclusion in the TEM and include:

·Processing Plant and Infrastructure

·Tailings Storage Facility

·Waste Rock Facility

·Water Management

·Environmental

·Reclamation

·General and Administration Costs

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21.5  Taxes and Royalties

21.5.1  Royalties

MPR has a royalty agreement with La Cuesta International, S.A. de C.V. and Minera Plata Real, S.A. de C.V., dated April 2006.  Under the terms, MPR paid a royalty payment of US $40,000 per year during the preproduction period.  Upon production, a 2% net smelter return (NSR) royalty from the Los Gatos concession and 0.5% NSR royalty from lands within a one-kilometer boundary of the Los Gatos concession began.  The royalty is reduced to 0.5% upon all payments reaching $10 million.  Upon commencing production, payments under the royalty agreement were deferred until March 31, 2021 with an annual interest rate of 4.5% applied to the outstanding balance.  During the deferral period, MPR pays a royalty payment of $100,000 per year until January 2021.  The agreement has no expiration date; however, the Company may terminate the agreement upon 30-day official termination notification.

The maximum royalty payment for this agreement is set at $15 million.  As of June 30, 2020, $14.4 million remained for future royalty obligations.

21.5.2  Taxes

21.5.2.1  Mexican Federal Tax

Corporate Mexican federal income tax is applied at a rate of 30%.  An operating loss carryforward is used to offset taxable income, thereby reducing taxes owed.

21.5.2.2  Mexican Special Mining Tax (SMT)

The SMT is assessed at 7.5% to pre-tax income (revenue minus operating costs and depreciation).

21.5.2.3  Mexican Precious Metals Tax

MPR does not sell precious metals as a product for revenue, but instead develops the underground mine to provide and sells access to the ore.  Accordingly, and as supported by its tax advisors, the Mexican precious metals tax is not applicable to the project economics.

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22.0  ECONOMIC ANALYSIS

Projected cash flows and economics are based on amounts for the calendar years beginning July 1, 2020 through the end of mine life in 2031.  The level of accuracy of the estimate is considered ±10%.  The following economic analysis includes Measured and Indicated Mineral Resources, which have been converted into Mineral Reserves.  The economic model is presented on an unlevered, post-tax, present value (PV) basis.  Valuation estimates presented in this technical report should be adjusted for existing LGJV current liabilities, receivables and long-term indebtedness.

22.1  Net Smelting Return

The estimate of the net smelting return (NSR) is summarized in Table 22-1.  Technical parameters supporting these estimates are shown in Section 2.0.

Table 22-1:  NSR

Description	Unit		Value
Zinc Concentrate
Payables
Zinc	$	000s	739,823
Silver	$	000s	133,538
Gold	$	000s	—
Subtotal	$	000s	873,361
Treatment Charges	$	000s	(194,917	)
Refining Charges	$	000s	—
Penalties	$	000s	(4,131	)
Freight	$	000s	(106,226	)
Zinc NSR	$	000s	568,088
Lead Concentrate
Payables
Lead	$	000s	384,327
Silver	$	000s	1,233,017
Gold	$	000s	67,025
Subtotal	$	000s	1,684,370
Treatment Charges	$	000s	(60,510	)
Refining Charges	$	000s	(98,305	)
Penalties	$	000s	(7,195	)
Freight	$	000s	(27,868	)
Lead NSR	$	000s	1,490,491
Net Smelting Return	$	000s	2,058,579

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22.2  Economic Results

The economic model is presented on an unlevered, post-tax, present value (PV) basis.  Valuation estimates presented in this technical report should be adjusted for existing LGJV current liabilities, receivables and long-term indebtedness.  Economic results are summarized in Table 22-2.  The analysis suggests the following conclusions, assuming no gearing:

·Mine Life: 11 years;

·Pre-tax present value (PV5.0%): $764 million;

·Post-tax present value (PV5.0%): $653 million;

·Taxes Paid: $148 million;

·Sustaining project capital of $267 million; and

·Initial project capital of $316 million (completed 2019).

Table 22-2:  TEM Results

Description	Unit Cost ($/t-milled)			LOM Value ($000s)
Net Smelting Return	$	214.04		2,058,579
La Cuesta Royalty as of June 2020	$	(1.50	)	(14,415	)
Net Revenue	$	212.54		2,044,164
Operating Costs
Mine & Surface Infrastructure	$	(83.58	)	(803,835	)
Operating Costs	$	(83.58	)	(803,835	)
Operating Margin	$	128.96		1,240,329
Capital Costs
Sustaining Capital Costs	—			(267,330	)
Capital Costs	—			(267,330	)
Pre-Tax Cash Flow
Cash Flow	—			978,867
PV5.0%	—			764,690
Post-Tax Cash Flow
Cash Flow	—			830,653
PV5.0%	—			653,166

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22.3  Sensitivity

Project sensitivity at a post-tax unlevered present value basis is shown in Figure 22-1 to Figure 22-4.  As shown below, the Project is most sensitive to silver price.

Figure 22-1:  Sensitivity Analysis for Silver Price

Figure 22-2:  Sensitivity to Gold Price

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Figure 22-3:  Sensitivity to Zinc Price

Figure 22-4:  Sensitivity to Lead Price

259

								END PROD<<
Description	Qty		Unit Cost		units	Total or Avg.		2020 1		2021 2		2022 3		2023 4		2024 5		2025 6		2026 7		2027 8		2028 9		2029 10		2030 11		2031 12
Market Prices
Silver	—		1		$/oz	18.99		18.99		18.99		18.99		18.99		18.99		18.99		18.99		18.99		18.99		18.99		18.99		18.99
Gold	—		1		$/oz	1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00		1,472.00
Zinc	—		1		$/lb	1.09		1.09		1.09		1.09		1.09		1.09		1.09		1.09		1.09		1.09		1.09		1.09		1.09
Lead	—		1		$/lb	0.87		0.87		0.87		0.87		0.87		0.87		0.87		0.87		0.87		0.87		0.87		0.87		0.87
Production
Mill Ore	—		—		kt	9,618		321		862		900		900		900		900		900		900		900		900		900		335
Silver	—		—		koz	80,481		2,914		8,680		10,423		9,468		9,225		7,341		6,556		5,278		5,886		6,764		5,925		2,022
Gold	—		—		koz	69		3		7		7		7		7		5		5		5		5		5		7		3
Zinc	—		—		klb	877,638		22,124		75,888		83,244		85,156		97,664		89,620		82,180		78,569		87,149		83,625		65,858		26,561
Lead	—		—		klb	498,713		13,425		43,770		50,246		45,490		48,301		48,103		51,100		50,219		46,879		47,203		40,682		13,295
Cash Flow
Gross Revenue
Silver	—		—		USD000s	1,528,328		55,332		164,826		197,924		179,795		175,181		139,414		124,496		100,231		111,773		128,444		112,521		38,393
Gold	—		—		USD000s	101,570		4,874		10,719		10,593		10,630		10,315		7,833		7,950		8,038		7,794		8,079		10,897		3,849
Zinc	—		—		USD000s	956,625		24,115		82,718		90,736		92,820		106,454		97,686		89,576		85,640		94,992		91,151		71,785		28,952
Lead	—		—		USD000s	433,880		11,679		38,080		43,714		39,577		42,022		41,849		44,457		43,691		40,785		41,067		35,393		11,566
Gross Revenue	—		—		USD000s	3,020,404		96,000		296,343		342,967		322,822		333,972		286,782		266,479		237,600		255,344		268,741		230,595		82,760
Net Smelting Return
Lead Concentrate	—		—		USD000s	756,744		51,857		183,233		165,023		127,045		115,153		114,434		0		0		0		0		0		0
Zinc Concentrate	—		—		USD000s	285,012		15,850		59,431		63,776		51,274		52,261		42,420		0		0		0		0		0		0
Net Smelting Return	—		—		USD000s	1,041,757		67,708		242,664		228,798		178,319		167,4149		156,854		0		0		0		0		0		0
Royalties
La Cuesta Royalty	—		—		USD000s	(14,415	)	(1,354	)	(4,150	)	(4,853	)	(1,127	)	(1,144	)	(963	)	(822	)	0		0		0		0		0
Royalties	—		—		USD000s	(14,415	)	(1,354	)	(4,150	)	(4,853	)	(1,127	)	(1,144	)	(963	)	(822	)	0		0		0		0		0
Net Revenue	—		—		USD000s	1,027,342		66,353		238,514		223,945		177,191		166,270		155,891		(822	)	0		0		0		0		0
Operating Costs
10 — Mining	—		1.0		USD000s	(803,835	)	(29,178	)	(73,874	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(29,342	)
Operating Costs			1.0		USD000s	(803,835	)	(29,178	)	(73,874	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(29,342	)
Operating Margin	—		—		USD000s	223,507		37,175		164,639		149,340		102,587		91,665		81,286		(75,427	)	(74,605	)	(74,605	)	(74,605	)	(74,605	)	(29,342	)
Capital Costs
10 — Mine & Surface Infrastructure	—		1.0		USD000s	(267,330	)	(28,581	)	(62,705	)	(38,302	)	(40,720	)	(29,375	)	(18,645	)	(13,545	)	(11,280	)	(9,495	)	(6,130	)	(5,345	)	(3,207	)
Capital Costs			1.0		USD000s	(267,330	)	(28,581	)	(62,705	)	(38,302	)	(40,720	)	(29,375	)	(18,645	)	(13,545	)	(11,280	)	(9,495	)	(6,130	)	(5,345	)	(3,207	)
Working Capital
Beginning Balance	—		—		USD000s	(149,063	)	0		(14,775	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	0
Ending Balance	—		20	%	USD000s	(160,767	)	(5,836	)	(14,775	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(14,921	)	(5,868	)
Change	—		—		USD000s	11,704		5,836		0		0		0		0		0		0		0		0		0		0		5,868
Cash Available for Debt Service
Operating Margin	—		1,090,868		USD000s	1,240,329		37,175		129,497		163,207		149,730		153,050		117,121		102,892		81,061		92,809		105,707		82,250		25,831
Project Capital (Equity)	—		—		USD000s	(267,330	)	(28,581	)	(62,705	)	(38,302	)	(40,720	)	(29,375	)	(18,645	)	(13,545	)	(11,280	)	(9,495	)	(6,130	)	(5,345	)	(3,207	)
Income Tax	—		—		USD000s	(148,214	)	(1,175	)	(2,579	)	(3,150	)	(21,878	)	(27,334	)	(16,203	)	(15,411	)	(8,787	)	(15,898	)	(19,657	)	(12,593	)	(3,549	)
Working Capital	—		—		USD000s	5,868		5,836		8,939		146		0		(0	)	0		(0	)	(0	)	0		0		(0	)	(9,053	)
Pre-Tax Cash Flow	—		—		USD000s	978,867		14,430		75,731		125,051		109,010		123,675		98,476		89,347		69,781		83,314		99,577		76,905		13,571
Cumulative	—		—		USD000s	—		14,430		75,731		125,051		109,010		123,675		98,476		89,347		69,781		83,314		99,577		76,905		13,571
Present Value	5.0	%	—		USD000s	764,690
Post-Tax Cash Flow	—		—		USD000s	830,653		13,255		73,152		121,901		87,132		96,341		82,273		73,935		60,994		67,416		79,920		64,311		10,023
Cumulative	—		—		USD000s	—		13,255		73,152		121,901		87,132		96,341		82,273		73,935		60,994		67,416		79,920		64,311		10,023
Present Value	5.0	%	—		USD000s	653,166

Figure 22-5: Technical Economic Model

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23.0  ADJACENT PROPERTIES

The only exploration program in the immediate area surrounding the Los Gatos project was conducted by VVC Exploration on the Santa Valeria project.  The Santa Valeria project is now owned by SSMRC, which is a different entity than MPR.  The Santa Valeria project located adjacent to the Southeast side of the Los Gatos project concession block and is considered as separate project.  Previously published accounts by VVC Exploration indicate that they have conducted mapping, sampling and have completed a 7-hole drill program.  The highlight of the drilling was a 1.5 m apparent thickness interval containing 145 grams of silver per tonne.  Since the sale of the project to SSMRC, the details of VVC Exploration’s work at the Santa Valeria project are no longer publicly available.

As the current owner, SSMRC is the source of this data.

The author has been unable to verify the information relating to the Santa Valeria project and the information provided is not necessarily indicative of the mineralization on the property that is the subject of this technical report.

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24.0  OTHER RELEVANT DATA AND INFORMATION

24.1  Hydrogeology and Mine Dewatering

Because much of the proposed underground mine would be developed at elevations below the regional water table elevation, a regional numerical groundwater flow model was developed to evaluate mine dewatering scenarios and potential impacts to local water resources.  The numerical model was based on a compilation of regional, local and project-specific geologic and hydrologic data.  Primary data sources included regional geologic mapping by the Mexican Geological Service (Servicio Geológico Mexicano [SGM] 2000; SGM 2014); exploration borehole lithologic logs and core photographs by MPR; project-specific geologic maps, cross sections, and three-dimensional structural geologic interpretations generated by Tetra Tech (Elson 2016) for resource estimation, by Rowearth (2016) for geologic evaluation of the project area, and by MPR during mineral exploration activities; on-site testing of wells, piezometers and boreholes (Tetra Tech 2015; Tetra Tech 2016); and geologic and hydrologic data and reports prepared for the Los Gatos Project (Hernández Bedolla 2015, López Ortega 2010, Schlumberger Water Services undated).  Tetra Tech updated the model in 2019 with additional site data collected since the 2016 model.

24.1.1  Hydrogeology

Groundwater in the Los Gatos model area is believed to originate from precipitation, underflow into and out of the model area and, potentially, heat-driven upward flow from beneath the model.

·Infiltration providing recharge to the groundwater system has been estimated at 5.94% of annual precipitation (Hernández Bedolla 2015).  Average annual precipitation totals about 498 mm, and average annual recharge totals about 30 mm.  The proportion of total recharge infiltrating into the subsurface is a function of soil and rock permeability, with more infiltration occurring in the areas with higher permeability.  Although some precipitation infiltrates into the subsurface in the upland areas, most precipitation reportedly moves as runoff into the valleys and infiltrates into the alluvium.  The drainages in the area only have surface flow immediately after a rain event, when they carry the runoff.  However, groundwater is normally found in the alluvium along the major drainages.  Recharge to the groundwater system was therefore conceptualized as occurring primarily in the lower areas underlain by alluvium along the arroyo Santo Toribio and to a lesser degree in the uplands.  Recharge sources other than precipitation include infiltration of a portion of water flowing from the Cieneguitas wells into the arroyo Santo Toribio about three kilometers upstream of the mine portal and water pumped from the dewatering wells and the mine that is discharged to the arroyo Santo Toribio about 500 meters east of the mine portal.  This infiltration likely adds water to the perched alluvial aquifer and contributes recharge to the deeper groundwater system.

·Underflow into and out of the model area is expected, since the model area represents part of a larger drainage basin.  The drainage basin extends approximately 10 km to the northwest of the model boundary and 2 km to the southeast of the model boundary.

·The relatively high groundwater temperatures in the project area, which increase with depth, are very likely due to intrusive rock bodies far beneath the model.  For purposes of modeling, areas identified as rhyolite domes were assumed to represent areas where heat-driven flow would be more prevalent, since the cooling intrusive bodies would presumably be closer to the surface there.

262

Groundwater flow in the project area is generally from northwest to southeast, with a strong downward gradient. However, faults appear to play a key role in the hydrogeology of the Los Gatos model area.  Some faults act as barriers, some as conduits, and some as both barriers and conduits.  The following summarizes primary faults and their hydraulic properties, based on borehole log descriptions, anecdotal evidence provided by site geologists, observations of groundwater inflow to the mine, and hydraulic head data.  While not all faults could be included in the numerical model, the major faults and lineaments listed below were represented as realistically as possible in the model, given the model grid spacing and limitations governing mathematical convergence of the model.

·Los Gatos fault and Lineament 4 — The Los Gatos fault acts as both a barrier and a conduit to groundwater flow.  Infilling and gouge along the fault result in the hanging wall side (north side) having a lower hydraulic conductivity (K).  A damage zone observed on the footwall side (south side) has a higher K.  Lineament 4 appears to be related to the Los Gatos fault, so it was assigned a similar configuration of damage zone and gouge/infilling.

·Anti-Gatos and East faults — These faults are believed to serve primarily as conduits due to the presence of incohesive breccia but little to no gouge (Rowearth 2016).  However, north of the Los Gatos fault, the East fault may also serve as a barrier.

·West fault — The West fault may serve as both a barrier and a conduit, based on observations of both incohesive and cohesive breccia (Rowearth 2016).

·Cieneguitas fault — Hydraulic data indicate that the cross-fault near the Cieneguitas area is a barrier to flow, since hydraulic heads just west of it are above ground surface and have an upward gradient, whereas hydraulic heads east of it are below ground surface and have a downward gradient.

·Southeast fault and Lineaments 1, 2, 3, and 5 — Because the groundwater system in the Los Gatos area appears to be highly compartmentalized, these faults were represented as barriers to groundwater flow in a similar manner to the Cieneguitas fault.  Anecdotal evidence from drilling in the vicinity of Lineament 3 indicates that it is gouge-filled, and borehole lithologic logs for the Southeast fault area similarly indicate the presence of clayey gouge.

In general, the Los Gatos area has a strong downward hydraulic gradient east of the Cieneguitas fault. Just west of the Cieneguitas fault, an upward gradient is observed; this suggests that the Cieneguitas fault forms a competent barrier to west-east flow.  Interestingly, the downward gradient near the Los Gatos mine site appears to be diminishing over the period of observation.  In other words, during the period of record for manual and transducer water level monitoring, the heads appear to be equalizing somewhat between the depth intervals monitored.  This phenomenon is believed to be due to the practice in past years of leaving exploration boreholes open.  Thus, the open boreholes can form conduits that provide vertical connection between shallow and deep units.  In the model, the areas with large numbers of older exploration borings were given a higher vertical K and called “drilled areas”.

Numerous short-term single-boring or single-well pumping or slug tests and one long-term (93-day) aquifer test have been performed at the project site between 2010 and 2016.  Although many of the tests were conducted on borings or wells that were open to more than one lithologic unit, the majority could be associated with one lithology or a small number of similar lithologies.  Figure 24-1 provides a statistical summary of the results and shows the 90% upper and lower confidence limits (UCL and LCL, respectively), the geometric mean and the number of tests (n) for the four primary rock types that were subjected to testing.

263

Figure 24-1. Hydraulic Conductivity Summary

Because the majority of these tests were single-well or single-boring tests and a primary influence on drawdown during the long-term pumping test was drainage through historical exploration borings, no site-specific estimates of storage coefficients could be made in 2016.  Since mine inflow measurements are now available and dewatering wells have been extracting groundwater since 2016, the storage parameters were varied during the 2019 model calibration to match longer-term drawdown and flow data.

The 93-day aquifer test conducted at Pozo 1 (formerly PW-15-01) has been retained for the 2019 model update as a pumping stress and a recovery period of 22 days.  In the 2016 model calibration, it was evident that two monitoring wells did not respond during this 93-day test (OW15-01-245m and GAGT-06) and continued to exhibit decreasing water-level elevations after the test was completed.  This trend of decreasing water levels continued at the same rate since 2016, showing little to no change in trend after dewatering commenced.  Both wells are located in very low permeability formations, based on slug testing, so very likely the water levels are still recovering from initial drilling and well installation activities.  Consistent with the 2016 calibration, the observed water levels at these two wells were not included in the 2019 model calibration.

Between the completion of the 93-day aquifer test in 2016 and early 2019, nine additional dewatering wells (Pozo 2 through Pozo 9 and Pozo 8A) have been installed at the site and pumped to remove groundwater from entering the underground mine workings.  Measurement of pumping rates from the dewatering wells and from the mine, combined with measurement of water levels in monitoring wells in the area, provided a substantial database for recalibration of the groundwater model.

264

Table 24-1 summarizes hydraulic conductivity (K) values for the hydrogeologic units included in the groundwater model, including the faults represented as having a conductive damage zone.  Table 24-2 provides details on the simulated hydraulic conductivities of various faults in the model domain.  The hydraulic conductivity values in both tables were determined during calibration of the groundwater model.

Table 24-1:  Hydrogeologic Units Represented in the Model

Hydrogeologic Unit	Model K zone	Horizontal K (Kh, meters/day)	Vertical K (Kv, meters/day)
Alluvium	8	4.0E+00	4.0E-01
Terrace Deposits	5	2.1E+00	2.1E-01
Epiclastics	2	4.8E-01	5.0E-04
Lapilli Tuff	4	3.1E-01	3.1E-03
Ignimbrite	6	5.4E-02	5.4E-04
Rhyolite Dikes/Domes	7	1.2E-02	2.4E-04
Dacite	3	1.0E+00	7.0E-03
Andesite	1	5.0E-01	2.3E-03
Anti-Gatos Fault	9	1.0E+00	8.0E-01
East Fault	10	1.0E+00	8.0E-01
West Fault	11	1.0E+00	8.0E-01
Los Gatos Fault and Lineament 1
Epiclastics - Low Kv	15, 16	1.3E-01	7.8E-03
Epiclastics - High Kh and Kv	17	1.2E+00	5.0E-04
Rhyolite - Low Kv	18	1.2E-02	7.8E-03
Rhyolite - High Kh and Kv	19	1.2E+00	2.4E-04
Dacite - High Kh and Kv	20	2.0E+00	7.0E-03
Andesite - Low Kv	13	1.2E-01	7.8E-03
Andesite - High Kh and Kv	14	1.2E+00	2.3E-03
Drilled Areas
Epiclastics	22	4.8E-01	1.1E-03
Rhyolite	24	1.2E-02	1.5E-01
Dacite	23	1.0E+00	6.2E-01
Andesite	21	5.0E-01	1.6E-02
Deep Bedrock (Layers 28-30)	25	1.0E-04	1.0E-06
Deep Bedrock (Layers 31-33)	26	1.0E-05	1.0E-07

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Table 24-2:  Faults Represented as Barriers to Flow in the Model

Horizontal Flow Barriers (HFBs)	Model HFB Zone	Hydraulic Conductivity (meters/day)
Los Gatos Fault	0	7.8E-03
Southeast Fault	2	7.8E-03
East Fault	3	1.1E-01
West Fault - north side	7	8.4E-03
West Fault - south side	8	7.4E-03
Cieneguitas Fault - north side	13	8.9E-06
Cieneguitas Fault - south side	14	3.8E-05
Lineament 1 - north side	11	1.6E-01
Lineament 1 - south side	12	1.6E-01
Lineament 2 - north side	9	1.6E-01
Lineament 2 - south side	10	1.0E-04
Lineament 3 - north side	5	1.6E-01
Lineament 3 - south side	6	1.6E-03
Lineament 4	1	8.6E-05
Lineament 5	20	1.0E-01

24.1.2  Groundwater Modeling

Tetra Tech simulated the groundwater flow system in the vicinity of the Los Gatos mine by updating the three-dimensional, finite-difference groundwater model (model) constructed in 2016.  The model was constructed using the MODFLOW-SURFACT code and calibrated to groundwater elevation data, drawdown data from aquifer testing and mine dewatering, and mine inflow data obtained from measurements and tests conducted at the project site.  Additional details can be found in Tetra Tech’s Los Gatos Groundwater Flow and Dewatering Model 2019 Revision.

266

24.1.2.1  Groundwater Model Setup and Calibration

The groundwater flow model was constructed using the Groundwater Vistas pre- and post-processing software, version 7.20 (Environmental Simulations Incorporated 2017).  The model domain is approximately 10 kilometers long by approximately six kilometers wide and covers approximately 60 square kilometers (Figure 24-2).  The horizontal extent of the domain was selected to incorporate enough of the drainage basin surrounding the proposed mine area that mine dewatering flows and drawdown from mine dewatering were unlikely to be significantly impacted by the boundaries at the model perimeter.  The extent of the model domain was not changed in the 2019 model.

Figure 24-2:  Groundwater Model Location

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The hydraulic conductivity distribution in the model was based on the three-dimensional geologic models of the mine area and on regional geologic mapping.  Initial values for the various lithologies were assigned based on results of on-site hydraulic tests.  Various MODFLOW and MODFLOW-SURFACT packages were used to represent boundary conditions and stresses, including general head boundaries to represent groundwater flow into our out of the model on the Northwest and Southeast sides and at the bottom of the model, no-flow boundaries to represent groundwater divides on the Southwest and Northeast sides, recharge to represent precipitation-derived, aerially-distributed recharge to the groundwater system, horizontal flow barrier boundaries to represent low-permeability faults, fracture wells to represent pumping or flowing wells, and drains to represent dewatering of the mine during predictive simulations.

The model was calibrated in steady-state mode to water-level elevations in 17 wells and in transient mode to drawdowns observed during a long-term aquifer test in the mine area.  During calibration, adjustments were made to horizontal and vertical hydraulic conductivities, specific storage and specific yield, the spatial distribution of recharge, heads and hydraulic conductivities in the general head boundary cells, and the hydraulic conductivities of the horizontal flow barriers.  Although the dewatering well pumping rates were incorporated into the model calibration, the measured water levels in the dewatering wells were not included and cannot be relied upon for model calibration.  Water levels in pumped wells do not represent water levels in the aquifer outside the well, because well efficiency and other head losses affect the measured water levels in pumped wells, usually resulting in lower water levels in the well than in the aquifer outside the well.  The well water levels could not be modeled because the dewatering wells have not been tested to determine well efficiencies and head loss parameter values.

Based on the recent drawdown and mine inflow data, adjustments were made to parameters that were most sensitive to drawdown and mine inflows.  During calibration, adjustments were made to the following model parameters:

·Horizontal and vertical hydraulic conductivity;

·Specific storage;

·Specific yield;

·Recharge;

·Drain conductance; and

·HFB hydraulic characteristics.

The hydraulic conductivity (K) values in the 2016 and 2019 models were determined during calibration of the models using the data available at the time.  The 2016 model calibration relied on K estimates from short-term, single-well tests and water-level change data from monitoring of observation wells during the test pumping of PW15-01.  Some of the important limitations of the data from the PW15-01 test were that, in terms of the groundwater model, the test well represented a single point in only one lithology (whereas the mine penetrates many different lithologies), and the test data spanned only about three months.  The 2019 model calibration used the data from the 2016 model calibration but, more importantly, incorporated the inflows to the mine that were measured during more than two years of actual mine development, as well as water levels measured in monitoring wells during that time.  Inclusion of the measured mine inflows was the main factor that resulted in the increases in the K values in the 2019 model compared to the 2016 model.

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24.1.2.2  Mine Plan and Dewatering

The Los Gatos mine employs stoping and drifting methods to remove ore from zones in four areas along the south (footwall) side of the Los Gatos fault over a period of approximately 12 years under a mine plan and schedule developed by MPR (2018).  The mine layout is shown in plan view on Figure 24-3.

Figure 24-3. Simulated Mine Drains and Dewatering Wells

The mine is divided into three areas called, for the purposes of this report, the Northwest, Central, and Southeast areas. Proposed mining elevations are between approximately 1190 masl and 1450 masl.  A ramp (decline) provides primary entry to the mine and access to the upper part of the ore in the Central zone.  A branch ramp to the Northwest area, with access currently to the 1370 and 1390 levels, a spiral ramp with access to the 1390 level, and a segment of the spiral ramp down to the 1350 level, have been constructed.  A branch ramp to the eastern part of the Central area also has been constructed and will eventually extend to the Southeast area.  In each area, horizontal tunnels will be constructed to provide access to the drifts and stopes from which ore will be mined.  The Central and Northwest areas will be mined earliest in the mine life, followed by the Southeast area.  The existing portions of the Northwest area and Central area ramps below the water table are experiencing inflows of groundwater.  Dewatering is by a combination of pumping from internal sumps within the mine and from a series of ten (as of this report date) dewatering wells outside the mine.  It is expected that all the mine areas below the water table will continue to experience groundwater inflow and require dewatering.

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Tetra Tech used the groundwater model to evaluate the current and three potential mine dewatering scenarios as part of efforts to identify an effective water management strategy for the existing and future mine operations.  Four scenarios were simulated; in each of the four scenarios inflow to the mine would continue to be removed by pumping from strategically located sumps within the mine:

·Base-Case — This scenario simulated groundwater inflow to the mine under the operational scenario in which the existing dewatering wells except Pozo 1 would continue to operate at their current rates without the addition of any other wells.

·Scenario 1 — Simulated the operation of four currently planned dewatering wells provided by mine site personnel in addition to the existing dewatering network.  The four new wells were each simulated as pumping at 14 l/s, which was the average rate achieved in the existing dewatering wells.

·Scenario 2 — The same network of existing and proposed dewatering wells from Scenario 1 was simulated.  However, five of the wells were simulated with increased pumping rates of 30 l/s to represent the use of larger pumps.

·Scenario 3 — The network of wells from Scenario 2 was supplemented by 9 new wells located as close as possible to mine workings.  All new wells were assumed to pump at only 14 l/s to distinguish the effects of new dewatering wells from the effects of higher pumping rates tested by Scenario 2.

The intent of the dewatering wells is to intercept groundwater before it reaches the mine and thereby reduce the amount of water that enters the mine and needs to be pumped out, treated, and either reused or disposed.  The setup of the groundwater flow model for simulating mining and mine dewatering is described below.  Figure 24-3 shows the locations of wells and drains used for each of the four scenarios. The additional wells simulated by Tetra Tech are labeled Pozo 21 through Pozo 29.

24.1.2.3  Groundwater Modeling Results

The reduction of predicted cumulative inflow to the mine ranged from 11% - 29% percent when compared to the base-case scenario with the currently existing, planned, and proposed dewatering wells (Table 24-3 and Figure 24-4).

Table 24-3:  Summary of Dewatering Simulation Results

Scenario Number	Wells Included	Cumulative  Mine Inflow  (Mm3)	Cumulative  Well  Pumpage  (Mm3)	Total  Cumulative Water Removed  (Mm3)	Reduction  in Mine Inflow Over  Base Case
Base Case	Pozos 2 through 8, 8A, and 9	104.2	55.5	159.7	—
1	Pozos 2 through 8, 8A, 9, 11, 12, 17, and 18	92.7	71.6	164.3	11	%
2	Pozos 2 through 8, 8A, 9, 11, 12, 17, and 18	78.9	96.1	175.0	24	%
3	Pozos 2 through 8, 8A, 9, 11, 12, 17, 18, 21 through 29	73.6	100.3	174.0	29	%

Mm3 = Millions of cubic meters

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Figure 24-4. Predicted Mine Inflow and Dewatering Well Pumpage for All Simulated Scenarios

Figure 24-5 graphically summarizes the predicted rates of groundwater inflow to the mine for the base case and three dewatering scenarios, and the modeled aggregate pumping rates for the dewatering wells for the dewatering scenarios.  The peak rates of inflow to the mine ranged from about 560 l/s for the Base Case with only the current dewatering wells operating, to about 400 l/s for Scenario 3, with an additional 9 dewatering wells each pumping at 14 l/s.  The reduction of predicted rates of groundwater inflow to the mine varied throughout the mine life in all scenarios, depending on the number of dewatering wells and simulated pumping rates.  The modeled aggregate pumping rates for the dewatering wells (Figure 24-5) increased until approximately 2029 as more dewatering wells were activated.  No wells were assumed to shut down during the life of the mine.

Predicted water-level drawdown was monitored at the locations of the dewatering wells during the simulation. Figure 24-6 and Figure 24-7 present a time series of predicted drawdowns at four of the dewatering wells under Scenario 2.  The graphs cover the modeled time period, starting February 1, 2016 and ending at the end of the mine life.  The drawdown displayed for the various model layers is limited by the layer-bottom elevations at the well because, although the model cell represented in the graph may become dry, the drawdown cannot go below the layer bottom.  Consequently, the drawdown in the aquifer near the wells is best represented by the drawdown for the deepest layer shown on the graph.

The maximum predicted drawdown at each well location occurred about two-thirds to three-quarters of the way through the mine life.  Maximum drawdown was approximately 200 meters at Pozo 5 and about 160 meters at Pozos 6, 11, and 12.  The predicted water-level elevations at the time of maximum drawdown were about 20 meters above the lowest nearby mining level at Pozo 5, about 55 to 60 meters above the lowest nearby mining level at Pozos 6 and 11, and about 5 meters above the lowest nearby mining level at Pozo 12.

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Figure 24-8 and Figure 24-9 show the horizontal extent of drawdowns predicted, respectively, at the time of maximum extent approximately two-thirds through the life of mine and at the end of mining.  At the time of maximum extent, predicted drawdown was about 200 meters near the Northwest spiral ramp and about 180 m near the deepest levels in the north part of the Central Zone.  By the end of mining, water levels had begun to recover.

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Figure 24-5:  Predicted Mine Inflows and Dewatering Well Pumping Rates for All Dewatering Scenarios

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Figure 24-6:  Predicted Drawdown at Pozos 5 and 6

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Figure 24-7:  Predicted Drawdown at Pozos 11 and 12

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Figure 24-8:  Maximum Predicted Drawdown

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Figure 24-9. Drawdown at End of Mining

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24.1.2.4  Groundwater Model Limitations

The Los Gatos numerical groundwater flow model was developed as a tool for quantitative estimation of potential inflows to the proposed underground mine and to provide data for mine dewatering designs.  The mine inflow estimates represent conceptual dewatering scenarios based on currently available geologic and hydrogeologic data for the project area and surrounding region, as well as the mine plans and schedule provided to Tetra Tech in September 2018.

The regional-scale flow model used to simulate the groundwater system has inherent limitations due to the simplifications necessary to represent complex natural systems. Model grid size and available data constrain the resolution and accuracy of the predictions.  Nonetheless, the predictions are useful for assessing the potential range of mine inflows, dewatering well locations, and pumping rates.

In general, monitoring of groundwater inflow to mines in fairly low-permeability rocks with relatively low fracture density and interconnectivity has shown that fractures can initially yield substantial rates of inflow that decrease rapidly over time.  The degree to which this occurs depends on how well-connected the fracture network is over large areas.  The equivalent porous media conceptual flow model used for the Los Gatos numerical model assumes that the fracture network is sufficiently connected to be simulated as a porous medium at the regional scale.  This conceptual model has been shown to be applicable on a regional scale.  However, as the scale becomes more local, small-scale fracturing and geologic structures play an increasingly significant role in groundwater inflow to the mine, and the equivalent porous media assumption can become less appropriate.  Difficulties encountered during the model calibration indicated that the degree of natural contrast in hydraulic conductivity between the faults and the surrounding geologic units may be greater than what can be simulated as an equivalent porous media.  It appears that the Los Gatos hydrogeological system is stretching the limits of the conditions that can be efficiently simulated using this type of model (finite-difference method, with the assumption of an equivalent porous medium).  The inflows predicted by this model are therefore averages that do not account for extreme high or low flows due to faults, fractures, other local-scale geologic features, or potential flow pathways such as unsealed or incompletely sealed boreholes.

The regional scale and equivalent porous media assumptions inherent in the numerical groundwater model, as well as unknowns regarding geologic structures and their hydraulic properties, may contribute to over- or under-estimation of the long-term groundwater inflow to the mine and dewatering wells and may result in predictions of drawdown different from those that will actually occur.  Uncertainty remains in quantitative estimation of the mine inflows and resulting drawdown.  As with many proposed mine operations, this uncertainty will remain until additional observations made during mine development are compiled, analyzed and incorporated into the model.

The model was constructed based on present-day conditions, but natural and anthropogenic changes may occur over the simulation period.  No attempt has been made to simulate possible future changes that could alter the groundwater system.  As predictive simulations extend further in time, the potential error associated with the predictions increases.

The factors described above limit the precision and accuracy of the model predictions.  However, the results presented here represent Tetra Tech’s current best estimates of groundwater inflows to the proposed mine and groundwater dewatering well operations associated with the project.  A limited sensitivity and uncertainty analysis were performed as part of this scope.  The uncertainty in these predictions could be further evaluated as part of a more detailed sensitivity analysis.  The uncertainty could be reduced by incorporating observations made during future mine development into the model, just as the uncertainty with mine inflows was decreased by incorporating inflow data collected since 2016.

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24.1.3  Production Dewatering

Development and mining underground started at the site in 2018.  Ground water removal is being tracked and this will allow for additional future planning for dewatering.  As planned, there are two components to the mine dewatering, which include pumping wells, and removal of water that has infiltrated the mine.  There are currently 13 dewatering wells, pumping from the surface, to try to prevent mine inflows.  The wells are shown in Figure 24-10 in relation to the mine plan.

Figure 24-10:  Pumping wells in relation to the overall mine plan.

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Water that infiltrates the mine is removed through a series of mobile units, which pump the water from the working faces, to a main sump for the mine.  Water is pumped from the main sump to the surface.  Water pumping data is being recorded, and more water has been encountered than estimated in the previous models.  Average rates pumped from the mine by area, are 98.4 l/s for the NWZ, 20 l/s for the Central Zone. Table 24-4 in March was due to the mining encountering a fissure, which lead to the temporary flooding of the mining level, which was then pumped out.

Table 24-4:  Water pumped from dewatering wells January-May 2020

	January L/S	February L/S	March L/S	April L/S	May L/S
1	0.0	0.0	0.0	0.0	0.0
2	0.0	0.0	0.0	0.0	0.0
3	20.0	0.0	20.0	0.0	63.1
4	15.3	9.5	15.2	15.2	0.0
5	78.2	78.8	79.6	79.1	76.1
6	20.0	18.3	20.0	16.7	75.3
7	63.1	56.8	56.8	56.8	53.6
8	14.2	13.1	13.9	14.0	13.8
8A	75.7	75.7	69.4	78.0	75.0
9	17.7	17.2	15.8	14.3	16.7
11	21.0	21.7	21.7	21.9	21.0
12	20.0	20.0	20.0	20.0	18.3
17	20.2	20.8	19.8	19.5	17.9
Total	365.4	331.9	352.1	335.4	430.8

The current pumping rates will need to be increased to give enough drawn down to mine the lower levels of the mine.  Currently, as can be seen in Figure 24-11, the pumping rates are approximately 550 l/s, which combines the dewatering wells, and mine infiltration.

Figure 24-11:  Water pumped at the Los Gatos site, blue line represents mine infiltration and orange represents the dewatering wells.

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24.1.4  Conclusions and Recommendations

24.1.4.1  Conclusions

The modeling results suggest that adding dewatering wells at the modeled locations and pumping rates will intercept a portion of groundwater inflow to the mine before it reaches the underground workings.  Three scenarios were considered.

Production has started at the mine and groundwater data is being collected.  To date, more water has been extracted than was planned based on 2019 ground water model.  Water temperatures also exceed the predicated temperatures.

Significant flows into the mine are expected to continue and will require continued in-mine dewatering.  However, adding wells or increasing the pumping rates at existing and already-planned wells will result in decreased flow into the mine.

Water pumped from the dewatering wells is used to meet on-site demand such as consumptive use in the mine, man camp domestic use, and process water makeup.  The excess water from the dewatering wells requires only cooling prior to discharge to the surface.  Water pumped directly from the mine requires treatment to reduce the suspended solids and possible hydrocarbons content prior to discharge. Consequently, limiting flow into the mine reduces treatment requirements.

24.1.4.2  Recommendations

Tetra Tech recommends an active dewatering strategy to reduce groundwater inflows into the underground workings.  This strategy is composed of the following depressurization/dewatering methods:

·Install future dewatering wells as close as possible to the underground workings.  The farther the well is away from the workings, the less effective it will be in reducing inflows.

·Install future dewatering wells to depths approximately 20 meters below the deepest mine depth, to allow pump placement below the mine floor elevations.

·Perform capacity tests on selected wells to better identify potential pumping rates.  A standard protocol for such capacity testing should be developed.

·Install larger capacity pumps in dewatering wells to reduce mine inflows.

·Continue to use dewatering well design similar to the recently installed wells.

·When discrete water-bearing zones are intersected during mining or during drilling inside the mine, install horizontal or inclined coreholes for depressurization.  To limit the volume of water requiring treatment before discharge, water from depressurization coreholes could be handled separately from the general mine inflow.  The cost of separate handling should be reviewed to determine whether it would provide an economic benefit.

Because cooling and discharge would be technologically simpler, more reliable, and potentially less expensive than treating to remove suspended solids and possibly other components, a combination of increasing pumping rates and adding new wells could be used to further reduce mine inflows.

The yields of the proposed dewatering wells are based on the typical pumping rates achieved at existing dewatering wells.  Tetra Tech recommends that the finally selected locations of the dewatering wells first be core-drilled from the shallowest practical depth to the total depth.  The core should be photographed and logged using methods consistent with the standard protocol for exploration core holes drilled for this

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project.  Detailed logging would allow the dewatering wells to be designed with total depth and screen placement based on the location-specific geology.

A new hydrological study is recommended to detail a plan for the remainder of the mine life.  This study will be essential to further water management, and the management of water in the lower levels of the mine.

24.2  Geochemistry

24.2.1  Waste Rock Characterization

A total of 21 waste rock samples and one ore sample were analyzed by static and kinetic testing methods.  Total elemental concentrations exceeded Maximum Permissible Limits (MPLs) for antimony, arsenic, barium, cadmium, and lead.  However, these were not leachable.  Acid Base Accounting (ABA) indicates that a portion of the waste rock is Potentially Acid Generating (PAG), leading to kinetic testing with the Kinetic Net Acid Generation (NAG) test.  This testing indicated a low capacity to generate acid.

The material currently being stockpiled at the site has low capacity to generate an acidic leachate, although some material within this storage facility can potentially generate acid.  The proportion of material that can generate acid is low and there is a general lag time before acid is generated.  As the material is being stored in a climate with distinct wet and dry rainfall periods, the armoring of pyrite is common in these environments, reducing oxidation potential.  There is also some, if limited, acid neutralizing potential, that will reduce the generation of acid leachate.

The potential to generate acid from any given waste unit does appear to have a finite length of acid generation and is not considered to be a long-term closure issue.  However, as there is a small percentage of material that can generate acid, it is recommended that a surface water monitoring station be established below the waste rock storage facility.  The production of leachate should be monitored, and water samples be collected at the beginning and end of the wet season to evaluate leachate generation at the field scale, and to determine whether management is required.  Given the location of the waste rock within a drainage, and proximity to the tailings storage facility, collection and disposal of any acidic drainage via pumping to the tailings storage facility, will likely be the most appropriate management strategy, if required.

From a closure perspective, the results indicate that long-term acid generation is unlikely.  Collection of surface water samples, as described above, will be used to evaluate this condition.  Given the low likelihood of acid generation, a soil and vegetative cover should suffice at closure.

24.2.2  Tailings Characterization

Tailings material to be placed within the designated tailings storage facility (TSF) are processed via the INCO process that involves the addition of sulfide dioxide and air, with a copper catalyst, for cyanide destruction.  A variety of samples of pre- and post-cyanide destruction material was analyzed by static and kinetic testing methods.

Total elemental concentrations exceeded MPLs for antimony, arsenic, barium, and lead.  However, these were not leachable.  ABA indicates that a portion of the waste rock is Potentially Acid Generating (PAG), leading to kinetic testing with the Kinetic Net Acid Generation (NAG) test.  This testing indicated a low capacity to generate acid.

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Kinetic testing supports the low capacity to generate acid, with leachate pH circum-neutral to alkaline.  Static and kinetic data indicates that metal leachate generation also appears to be low.  Although TCLP testing indicates the potential for some lead leachate generation, the kinetic data indicates that lead leachate above MPL will not occur during operation and closure.  Kinetic data does indicate the potential for cadmium discharge at closure once draindown has been completed.  However, on balance with all the available data, this is considered to be low probability and risk.

24.3  Surface Water Hydrology

The objective of this report is to provide the surface water hydrology and water management basis of design for the Tailings Storage Facility (TSF) at MPR’s Los Gatos Project.  This study was developed using site geotechnical investigations and weather data from adjacent meteorological station 8057(1).  Water management has been designed in accordance with NOM-141-SEMARNAT-2003 (Procedure for Characterizing Mine Tailings and Criteria for the Characterization and Preparation of the Site, Project, Construction, Operation, and Post Operation of Mine Tailings Dams), NOM-011-CNA-2000 (Water Resource Conservation-Que) and the dam safety standards set forth by CONAGUA.  Additional guidance was provided by Joel Hernandez-Bedolla, who authored Estudio Hidrológico Superficial, Subterráneo e Hidráulico “Los Gatos” Satevó, Chihuahua.

(1)  Located 5-kilometers East of Los Gatos, 32-year period of record

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24.3.1  Methodology

Design hydrology for the Los Gatos site was evaluated using the SCS curve number method to model losses to soils and specify unit hydrograph transformation.  The SCS method relies on basin characteristics, design storm rainfall depths and temporal distribution to calculate volumetric flow rates.  The Los Gatos Project site is situated in the Humid Hydrologic Zone, exhibiting sloped to mountainous topography.  The Probable Maximum Precipitation (PMP) event was used as a basis of design for the TSF spillway and diversions and the 1,000-year event was used as a design basis for the Mill channel(2).  The TSF ultimate embankment height would measure up to 39-meters, resulting in a categorization as High (“Mayor” — reference Table 24-5, below) dam size classification(3).  As such, the spillway has been designed based on a storm equal to or greater than the 10,000-year storm.  For the purposes of this study, the PMP has been used.  Hydrology for the site was computed using HEC-HMS software and applying input parameters discussed herein.

Table 24-5:  CONAGUA Design Storm Requirements for Impoundments

Source:  CONAGUA Manual para el Control de Inundaciones Table 4

(2)  Personal communication with Joel Hernandez-Bedolla September 15, 2016.

(3)  CONAGUA Manual para el Control de Inundaciones Table 4

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24.3.1.1  Basin Characteristics

Soils

Site soils were categorized using the Soil Conservation Service (SCS) and NOM-011-CNA-2000 methods.  The SCS method relies on a curve number, whereas NOM-011-CNA-2000 relies on a runoff coefficient, Ce.  Each parameter is indicative of infiltration capability and resulting runoff but is applied differently.

Geotechnical samples collected within and above TSF indicate exposed, weathered bedrock mixed with higher infiltrative (higher porosity) soil groups.  Dominant soil types are shown in Figure 24-12.

The dominant soil type and resulting hydrologic soil group as described for use in the SCS runoff method is shown for each region below.

·TSF North Region.  Well graded sands with exposed and weathered bedrock.  Overall Hydrologic Soil Group D.

·TSF Paleo Channels Region.  Sand with silt and gravel (Hydrologic Soil Group A/B)

·TSF South Region.  Silty gravel with sand and bedrock (Hydrologic Soil Group C/D)

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Figure 24-12:  Soils Map of the Project Area

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The hydrologic soil group describes the runoff potential for a particular soil type and is defined below by the SCS in USDA Technical Release 55.

Table 24-6:  TR-55 Hydrologic Soil Groups

Hydrologic Soil Groups

Group A is sand, loamy sand or sandy loam types of soils.  It has low runoff potential and high infiltration rates even when thoroughly wetted.  They consist chiefly of deep, well to excessively drained sands or gravels and have a high rate of water transmission.

Group B is silt loam or loam.  It has a moderate infiltration rate when thoroughly wetted and consists chiefly or moderately deep to deep, moderately well to well drained soils with moderately fine to moderately coarse textures.

Group C soils are sandy clay loam.  They have low infiltration rates when thoroughly wetted and consist chiefly of soils with a layer that impedes downward movement of water and soils with moderately fine to fine structure.

Group D soils are clay loam, silty clay loam, sandy clay, silty clay or clay.  This HSG has the highest runoff potential.  They have very low infiltration rates when thoroughly wetted and consist chiefly of clay soils with a high swelling potential, soils with a permanent high-water table, soils with a claypan or clay layer at or near the surface and shallow soils over nearly impervious material.

Source:  USDA Technical Release 55

Upland soils of the TSF are assumed to be similar in character to the samples collected within the TSF footprint.  Given the dominant C/D and D soils types, and considering the Los Gatos site is a desert landscape, an SCS curve number of 86 was assigned to the TSF uplands.  These curve numbers have been used within calculations discussed in Section 24.3.1.3 .

NOM-011-CNA-2000 sub-divides soils into one of three hydrologic types, similar to the table presented above.  The NOM-011-CNA-2000 table is presented below.

Table 24-7:  NOM-011-CNA-2000 Hydrologic Soil Type and K Values as a Function of Soil Use

Soil Type	Characteristics
A	Permeable soils such as sands or loess
B	Moderately permeable soils, more compact than Soil Type A
C	Nearly impermeable soils, such as a thin layer of sand or loess underlain by an impermeable layer or clay

Soil	Soil Type
Use	A	B	C
Fallow, uncultivated or bare	0.26	0.28	0.3
Cultivated
Row crops	0.24	0.27	0.3
Rotated crops	0.24	0.27	0.3
Grains	0.24	0.27	0.3
Pastures
Minor	0.14	0.2	0.28
Regular	0.2	0.24	0.3
Excessive	0.24	0.28	0.3

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Soil	Soil Type
Use	A	B	C
Forest
Greater than 75 Percent Cover	0.07	0.16	0.24
50 to 75 Percent Cover	0.12	0.22	0.26
25 to 50 Percent Cover	0.17	0.26	0.28
Less than 25 Percent Cover	0.22	0.28	0.3
Urban	0.26	0.29	0.32
Roads	0.27	0.3	0.33
Permanent Pasture	0.18	0.24	0.3

Source:  NOM-011-CNA-2000 Table 1

Overall, the upland areas above the TSF has been classified as soil type C.  Given the land use has been categorized as fallow, uncultivated and/or bare, the resulting K value would be 0.30.  For K values greater than 0.15, the rainfall-runoff coefficient, Ce, is calculated as follows:

Ce = K*(P-250)/2000 + (K-0.15)/1.5 Equation 1

Where

Ce = runoff coefficient

K = soil coefficient

P = annual precipitation (mm)

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Vegetative Cover

Vegetation is characterized by a semi-desert landscape, with typical low brush vegetation in the slopes including lechuguilla, sotol, ocotillo, yucca, sage, bear grass, and other types of indigenous grasses.  Larger brush and trees are common along the main watercourses, with the presence of oak, cypress, poplar, cottonwood, huizache, and mesquite, among others.  Vegetative cover is low to moderate, with sparser occurrence in the vicinity of the TSF, as shown in Figure 24-13.  The presence of vegetation serves to reduce the curve number by interception of rainfall, either by interception on the canopy (leaf surfaces) or plant litter on the ground.

Figure 24-13:  Site Photo Showing Vegetative Cover

Contributing Areas

Basin areas contributing to the TSF were delineated within AutoCAD Civil 3D and are summarized in Table 24-8.

Table 24-8:  Basin Areas

Basin	Contributing Area (km2)
TSF Diverted Uplands	1.15
TSF Un-diverted Uplands	0.38
TSF	0.65
Mill Diverted Uplands	0.21

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Lag Time

Time of concentration(4) was calculated for the uplands above, and within, TSF.  Transformation of excess precipitation to direct runoff using the SCS unit hydrograph method requires calculation of lag time(5).  For ungauged basins, lag time (tlag) is computed as a function of the time of concentration (tc), as follows:

tlag = 0.6tc Equation 2

Lag times are summarized below.  Short times of concentration and associated lag times result in more rapid conveyance of storm runoff to diversion channels and spillways.  Given steep terrain and short flow paths, lag times associated with and natural areas occurring above the Mill were all set to 3.6 minutes (the minimum recommended value).

Table 24-9:  Lag Times

Basin	Lag Time (minutes)
TSF Diversion	6.1
Mill Diversion	3.6
TSF Un-diverted	9.9

24.3.1.2  Design Storms, Typical Rainfall and General Meteorology

Design storms have been calculated previously within the Hidrológico Botadero report and by statistical evaluation of the 8057 meteorological gauge.  The 8057 gauge is maintained by CONAGUA and was selected due to its favorable proximity to the site; it is located 5-kilometers to the East.  Additionally, the period of record for the gauge spans 32 years and the data are of good quality.  The design storm rainfall depths are summarized in Table 24-10.

Table 24-10:  Design Storms

Return Period (year)	Precipitation Duration (hour)	Total Precipitation (mm)
2	24	40
5	24	49
10	24	55
25(6)	24	61
50	24	67
100	24	72
500	24	83
1,000	24	88
PMP(7)	24	216

(4)  The time required for rainfall occurrence at the most hydrologically remote point in the basin to report to the outlet

(5)  Defined as the delay between peak precipitation and peak discharge

(6)  Calculated by Tetra Tech using Gamma, Log Pearson III and Extreme Value I methods

(7)  Calculated by Tetra Tech using Hershfield method

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Additionally, the Probable Maximum Precipitation (PMP) of 216 mm was calculated using the Hershfield (1965) method.

The minimum, average and maximum monthly total rainfall(8) is presented in Table 24-11, for reference.  The greatest observed rainfall occurring over a 24-hour period during the period of record was 61 mm, occurring August 23, 1965.

Table 24-11:  Minimum, Average and Maximum Total Monthly Rainfall for Los Gatos Site

Month	Minimum Monthly Precipitation (mm)	Average Monthly Precipitation (mm)	Maximum Monthly Precipitation (mm)
Jan	0	9	38
Feb	0	9	38
Mar	0	5	45
Apr	0	5	33
May	0	11	66
Jun	0	42	171
Jul	30	110	230
Aug	20	123	237
Sep	8	94	246
Oct	0	25	106
Nov	0	10	42
Dec	0	12	79
Total	58	454	1331

With the exception of precipitation, meteorological data have been collected on site from January 2012 to present.  Throughout the year, the general wind direction is West to Southwest.  Wind velocities are highest in April and May, at 2.0 and 2.1 m/s, respectively.  Given the nearest populated area is located approximately 6 kilometers to the Southeast; fugitive dust is not expected.

(8)  Pivot table on data from Estaciones Climatologicas meteorological station 8057

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24.3.1.3  Design Storm Runoff

The SCS curve number method was used to determine the runoff that would occur from the 1,000-year, 24-hour design storm of 88 mm and from the PMP storm of 216 mm.  A curve number of 86 was selected for the TSF uplands, assuming fair coverage of desert scrub overlaying soil hydrologic groups C and D and for consistency with the Hidrológico Botadero report.

A curve number of 86 was selected to model the natural upland areas.

The SCS method uses the curve number to compute the initial soil retention capability when the storm begins, and then calculates direct runoff (Q) as follows:

S = 1000/CN – 10 Equation 3

Where

S = maximum soil retention after runoff begins (in)

CN = curve number (unitless)

Equation 4

Where

Q = direct runoff (in)

P = precipitation (in)

0.2S = the initial abstraction, or amount of initial storage in the soil prior to runoff being initiated.

Temporal Distribution

The 1,000-year and PMP storms were modeled using an SCS Type II temporal distribution, wherein the greatest rainfall intensity occurs during the middle of the storm, as shown in Figure 24-14.

Figure 24-14:  Type II Temporal Distribution
(excerpt from Urban Hydrology of Small Watersheds)

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HEC-HMS software was used to account for losses to soil and transform the design storms by applying the appropriate unit hydrographs and temporal distributions described herein.  Excess runoff was then applied to basin areas to determine total storm volume and flow rate.

Average Annual Runoff Volume

In addition to 24-hour design storm runoff rates and volumes, which were calculated using the SCS curve number method, the average annual runoff volume was calculated using the equation from NOM-011-CNA-2000:

Ve = Ce*Vp Equation 5

Where

Ve = average annual runoff volume

Ce = runoff coefficient (defined in section 0)

Vp = average annual rainfall volume

An annual rainfall depth of 454 mm would result in a runoff coefficient of 0.14, using Equation 1.  The rainfall and coefficient were applied to the basin areas summarized in Table 24-12 resulting in the following annual runoff volumes:

Table 24-12:  Average Annual Runoff Volume

Basin	Average Annual Runoff Volume (1,000 m3)
TSF Diverted Uplands	73
TSF Un-diverted Uplands	23
TSF	41
Mill Diverted Uplands	13

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24.3.1.4  TSF Stage-Storage-Area

Stage-storage relationships for the TSF is presented in Figure 24-15 graphically; all surface area within each dam is assumed to be water.  The PMP was modeled as a storm event occurring onto diverted and un-diverted uplands, and onto the TSF pond surface(9).  At the time of the storm, a freeboard of 2-meters was assumed between the TSF embankment crest and the pool water surface elevation.  An emergency spillway for the TSF was modeled with a sill elevation equal to the pool water surface.  The majority of the PMP was accommodated by the 2-meter freeboard, prior to activation of the emergency spillway.  Given the large pond surface areas, depth of flow within the emergency spillway was noted to be minor.  Results of the hydrologic analysis are presented in greater detail within the following section.

Figure 24-15:  TSF Stage Storage

24.3.1.5  Stormwater Conveyance Structures Basis of Design

HEC-HMS results are summarized in Table 24-13.

Table 24-13:  Source of Basis of Design

Water Management  Structure	Design Storm	Storm Volume (1,000 m3)	Storm Discharge (m3/s)	Guidance for Design Storm Selection
TSF South Diversion	PMP	199.2	76	Personal communication with Asesores en Impacto Ambiental y Seguridad (ASI); NOM-141-SEMARNAT-2003
Mill Diversion	1000-year	11.1	4.9	Personal communication with ASI
TSF Emergency Spillway	PMP	28.6	0.5	NOM-141-SEMARNAT-2003, (Comisión Nacional del Agua (CONAGUA)
TSF East Diversion	PMP	210.3	80.9	Personal communication with ASI; NOM-141-SEMARNAT-2003

(9)  Conservatively, the ultimate condition was modeled

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Additionally, 2 meters of freeboard must be maintained from the TSF embankment crest over water(10).

24.3.1.6  Results

The peak flow rates were input into FlowMaster, a hydraulic calculation package, to estimate channel dimensions (refer to Table 24-14) and evaluate channel stability (refer to Table 24-15).  Diversion channels and spillways are assumed to be earthen lined and armored as needed using riprap or similar lining, with the addition of energy dissipaters or other erosion controls in steeper segments exhibiting velocities greater than 1 m/s.  For the capacity and stability calculations, a robust 12-inch D50 riprap lining (roughness equal to 0.078) was assumed throughout.

Table 24-14:  Estimated Stormwater Conveyance Structure Dimensions, Capacity Analysis

Water Management Structure	Description	Bottom Width (m)	Depth(11) (m)	Length (m)	Excavation Volume (1,000 m3)
TSF South Diversion	trapezoidal channel	4	3.5	2,206	135.4
Mill	trapezoidal channel	2	1.4	678	9.8
TSF Emergency Spillway	trapezoidal channel	2	0.4	176	0.8
TSF East Diversion	trapezoidal channel	4	2.6	646	28.2

Table 24-15:  Channel Slopes and Velocities, Stability Analysis

Water Management Structure	Minimum Slope (%)	Minimum Velocity (m/s)	Maximum Slope (%)	Maximum Velocity (m/s)
TSF South Diversion	1	2.0	1	2.0
Mill Diversion	2	1.68	17	2.9
TSF Emergency Spillway	2	0.83	18.5	1.8
TSF East Canal de Desvío	4	3.42	22	6.4

(10)  Per table 5.6.13 of NOM-141-SEMARNAT-2003

(11)  Does not include 30-centimeters of freeboard

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The freeboard assigned to each channel is a function of the discharge it conveys; freeboard criteria are shown below.

Table 24-16:  Freeboard Criteria

Discharge (m3/s)	Freeboard (cm)
0 - 0.5	15
0.5 - 1	20
1 - 3	25
3 - 10	30
10 - 20	35
20 - 40	40
40 - 60	50
60 - 100	60

The diversion channels were constructed for the ultimate TSF footprint.  Preliminary modeling shows the TSF emergency spillway conveys minimal flows during the PMP event for a starting water surface elevation of 1638, or 2-meters below the crest of the dam.  This elevation was also assigned to the spillway sill. However, a 2-meter wide emergency spillway is assumed for conservatism.  Channels have been constructed by excavation and windrowing and include energy dissipaters as needed to mitigate erosion and discourage high velocities.

The longitudinal profiles for each channel were evaluated for stability.  For channels constructed in sandy regions, segments exhibiting a maximum velocity greater than 1 m/s have been designed to include drop structures, riprap lining or other energy dissipation to reduce erosion potential.  Channels constructed in weathered bedrock dominated zones have been evaluated on a case-by-case basis.

Refer to Figure 24-16 for typical channel section views.  The freeboard shown for each channel is consistent with freeboard criteria presented in Table 24-16; dimensions are based on the capacity and stability evaluation.

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Figure 24-16:  Channel Section Views

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24.4  Site-Wide Water Balance

24.4.1  Introduction and Discussion

Tetra Tech prepared a site wide water balance model for the Minera Plata Real (MPR) Los Gatos project for a life of mine (LOM) of approximately 11 years.

Model parameters were developed from information and data obtained from project mine planners, process engineers, geotechnical engineers, hydrologists and hydrogeologists.  Source values used for model parameters and their sources are summarized below in Table 24-17.

Table 24-17:  Model Parameters and Source

Parameter	Value	Source
Production (tonne/day)	2,500	Stantec
Tailings to TSF/paste	60/40	Minera Plata Real
Specific gravity of tailings solids	2.69	SGS: process mass balance dated 19Sept2016
Percent slurry solids by weight	55%	TGI: Design Criteria
Deposited tailings specific weight (kN/m(3))	15.4	rheology test results
Deposited tailings dry density (tonne/m(3))	1.5	rheology test results
Pan evaporation coefficient	0.7	typical value for region
Ultimate Embankment crest elevation (m, MSL)	1639	Tetra Tech
Embankment Freeboard Over Water (m)	2	NOM-141-SEMARNAT-2003 (humid hydrologic zone)
TSF Ultimate Footprint (m(2))	651,360	Tetra Tech
Plant Availability	92%	SGS: Design Criteria
Target Mill Feed (tonne/hr)	113.2	SGS: process mass balance dated 19Sept2016
Target Reclaim from TSF (m3/hr)	27.47	SGS: process mass balance dated 19Sept2016
Target Raw Water Demand (m3/hr)	50.32	SGS: process mass balance dated 19Sept2016
Mine Water and Well Water (L/s)	Time series	Tetra Tech groundwater model with amplified stress periods
Potable/domestic demand (L/person/day)	80	International Finance Corporation paper on Mine Worker Accommodation
Personnel Construction Phase; Personnel Mining Operations Phase	600; 400	Estimate provided by SGS
Precipitation and Evaporation Data	Time series	CONAGUA Gauge 8057
Runoff Coefficient	0.138	Calculated per NOM-011-CNA-2000

The updated water balance model used the production start date of July 1, 2019 with respect to mining and processing operations.  The assumed production end date is August 15, 2029 for a duration of 10.2 years.  After August 15, 2029, the water balance assumes an additional 1.3 years to account for mine dewatering required during the mine closure period, allowing for equipment and infrastructure associated with the underground workings to be dismantled and removed.  Including this closure requirement, the

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total model simulation period is 11.5 years.  The model uses stochastic processes to simulate multiple variations of project variables on a daily time step for 300 realizations, or equally likely scenarios.  This process produces statistically valid results of model predictions.

24.4.2  Model Components

The water balance tracks inputs and outputs from reservoirs, to calculate change in storage.  The reservoir elements within the Los Gatos model include the TSF pond, the process facility, and the mine workings dewatering settling pond.  The interceptor well field is not modeled as a reservoir; however, inflows and outflows are tracked over time.

Water and solids are tracked separately within the TSF.  Solids and pore water associated with the slurry accumulated over time and are assigned no pathway for removal during operations.  Water is allowed to enter and exit the TSF, such that the following simple water balance equation would be satisfied:

Equation 1.

Inflows, outflows and storage are presented in greater detail below.

24.4.2.1  Meteorology

CONAGUA gauge 8057 was used as a basis for model rainfall and evaporation.  Daily rainfall was determined by a random, year-based lookup from the 8057 gauge time series.  Given that the 8057 gauge only collected evaporation data for one-year, daily evaporation rates were identical for every year modeled.  Pan evaporation data were converted for use by applying a factor of 0.7, which is an accepted correction factor.  Net positive rainfall is calculated prior to runoff calculations as follows:

Equation 2.

Net rainfall (and resulting runoff) is set to zero for days where evaporation exceeded rainfall.

24.4.2.2  Tailings Storage Facility

For the 2016 Feasibility Study, the TSF was designed as a water dam with the intention to hold large amounts of water along with tailings.  As a result, the 2016 water balance model treated the TSF as a water dam with solids and liquids being accounted for together.  The design and construction have since changed to resemble a conventional TSF with plans to directly discharge water offsite which originally was intended to be stored in the TSF.  The 2019 water balance model was updated to reflect this change and the TSF supernatant pond is now modeled separately from the deposited solids and pore water.

24.4.2.2.1  TSF Supernatant Pond

The TSF supernatant pond is modeled so that its geometry, stage and capacity were constantly evolving as more tailings are deposited in the TSF and as the embankment lifts are constructed.  Seventy-five percent of the slurry water was assumed to report to the supernatant pond upon tailings deposition.  The remaining 25% of slurry water is assumed to be locked in the tailings for the entirety of the model period.

24.4.2.2.2  TSF Supernatant Pond Inflows

Water inflows to the TSF include precipitation directly onto pond surfaces, minor un-diverted runoff from upland basins, and 75% of the water associated with the slurry.

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The percentage of rainfall that becomes runoff was calculated using NOM-011-CNA-2000 by applying a runoff coefficient of 0.138.  Development of the runoff coefficient for the Los Gatos site is discussed in greater detail within the Hydrology Report.  Runoff from un-diverted upland basins and precipitation occurring directly onto the TSF pond surface contribute to inflows.

60 percent of the total slurry is delivered to the TSF while the remaining 40 percent is assumed in the model to be placed underground as paste backfill.  Slurry water enters the TSF at a rate of 13.5 l/s, and slurry solids enter at a volumetric rate of 11 l/s.

24.4.2.2.3  TSF Supernatant Pond Outflows

Water exits the TSF as evaporation and reclaim.  Water is reclaimed to the mill at an approximate rate of 7.6 l/s and is delivered to the zinc circuit only.

The TSF is modeled as a lined facility.  Seepage loss from the TSF into the overdrains is not modeled as part of this study.

24.4.2.3  Well and Mine Dewatering

The well and mine workings dewatering flowrates are shown in Figure 24-17.  The mine working dewatering daily outputs were averaged over three months to smooth large fluctuations between steps of the 2019 Tetra Tech Groundwater Model discussed in Section 24.1.  The extracted water typically exhibits a temperature between 50 and 65 degrees Celsius and must be cooled to 40 degrees Celsius prior to use or discharging to the Santo Toribio River.

Figure 24-17:  Well and Mine Working Dewatering

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24.4.2.4  Well and Domestic Water

Well water has exhibited acceptable water quality(12) and would only require cooling prior to use.  It would therefore serve as a good source for process makeup water (14.6 l/s, typical), underground usage (15 l/s, typical) and domestic water demand (0.4 l/s, typical).  Any water in excess of these demands would be released once it has cooled to within discharge limits.

Domestic wastewater is assumed to be treated by a package plant prior to release.  Ninety percent of the domestic water demand is assumed to be converted to wastewater that must be treated prior to discharge to the environment.

(12)  Compiled lab reports provided by Tetra Tech Groundwater Team on September 27, 2016.

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24.4.3  Results

24.4.3.1  TSF Results

The embankment raise schedule is presented in Figure 24-18, assuming a starter phase 1 dam crest of 1,618.5 masl.  The schedule is also shown below in Table 24-18.  The TSF supernatant pond reaches an elevation of 1632 masl at the end of the LOM.

Table 24-18:  TSF Embankment Raise Schedule

Date	Height (m, MASL)
Starter Phase 1	1618.6
8/1/2020	1625.6
12/31/2021	1629.0
06/21/2024	1632.5
08/06/2026	1639.0

Figure 24-18:  TSF Embankment Raise Schedule

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24.4.3.2  Dewatering and Discharge Results

Over the life of the mine 9.9 Mm(3) of cooled well and workings dewatering water would be used on site and 145 Mm(3) would be discharged.  Refer to Figure 24-19 for well extraction, use (includes process freshwater makeup, underground use and domestic) and discharge rates over the life of the mine.  No freshwater deficiency is expected for the site.

Figure 24-19:  Well Water, Use and Discharge

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24.5  Tailings Management

This report utilizes the conventional slurry tailings storage facility (TSF) method, which involves the deposition of slurry tailings from the process plant.  The slurry is deposited using a pump and tailings distribution pipeline along the crest of the TSF.  The pipelines are pressure rated high-density polyethylene (HDPE).  The TSF facility has been designed considering the following:

·The TSF facility site provides the best location to hold the required storage capacity.

·The use of rockfill during the construction of the facility.

·The storage of additional water from the underground mine operations.

The design of the TSF allows for a life of mine (LOM) of 11 years, and a storage capacity of approximately 9.3 Mm(3), while maintaining a 2-meter freeboard on the facility.

The ultimate TSF would be constructed in four stages using downstream construction methods as shown in Figure 24-20 and Figure 24-21.  These stages will be constructed with upstream slopes of 2H:1V and downstream of 2.5H:1V (horizontal: vertical).  A maximum crest width of 15-m with a maximum crest elevation of 1,639 masl.

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Figure 24-20:  Tailing Storage Facility Plan View

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Figure 24-21: Tailing Storage Facility Section View

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Tailings containment has been achieved using a combination of factors and engineering controls such as the incorporation of a 1.5 mm (60mil) high-density polyethylene (HDPE) geomembrane liner over the impoundment area.  This liner has been combined with a series of underdrains to help manage water under the liner system.  The water collected from these systems is collected in a series of collection ponds located downstream of the TSF.  Additional to the drain systems, water is managed with the incorporation of surface water diversion channels and a spillway, which are discussed in Sections 24.3 and 24.4 of this report.

In general, the liner shall extend along the base of the impoundment as well as the upstream slope of the embankment and shall be anchored along the edges of the current construction stage.  The impoundment configuration of the TSF consists of (from top to bottom):

·Tailings and water;

·1.5 mm HDPE geomembrane liner;

·Geosynthetic clay line (GCL);

·On the embankments, 3-m thick filter zone (zone A);

·1.5-m thick transition zone (Zone B),

·Rock fill (Zone C);

·Underdrain system and blanket drain.

24.5.1  Topography

Per NOM-141-SERMANAT-2003, “Table 2” any slope grade steeper than approximately 18% is considered Mountainous Land.  The topography of the site can be classified as “Mountainous Land” as some of the steepest slopes for the TSF were 18 % and 23%.

24.5.2  Site Seismicity

The Mexican Servicio Sismológico Nacional has divided the republic into four seismic zones (http://www.ssn.unam.mx/website/jsp/region_sismica_mx.jsp) based on the catalog of earthquakes that occurred during the last century, great earthquakes mentioned in historical records, and ground motion records of some of the larger events of this century.  The resulting seismic hazard map for Mexico places the Project site within Zone B, an intermediate zone where infrequent earthquakes have been recorded but peak ground accelerations do not exceed 70% of the acceleration of the soil.  Zone B is called the “Penesísmica” region, with a peak ground acceleration (PGA) range of 0.8 to 1.6 m/s2 (approximately 0.08g to 0.16g) for 10 percent probability of exceedance in 50 years.  The site is located on the border of Seismic Zones A and B; in order to be conservative for the purposes of this design, it is assumed that the site is located in Seismic Zone B (Figure 24-22).

For structures, such as the TSF, which present a significant hazard for damage to the environment, the structure should be capable of tolerating displacements from a maximum credible earthquake (MCE) without catastrophic loss of tailings or supernatant, although limited damage to the structure may take place.  This approach is consistent with current International Commission of Large Dams (ICOLD) (Wieland, 2005) and Federal Emergency Management Agency (FEMA) (2005) guidelines for seismic stability, which indicates that “significant structural damage is accepted” for maximum earthquake ground motions although no uncontrolled release from the reservoir shall occur.  The deterministic seismic hazard evaluation for the project indicates a maximum site PGA of 0.18 acceleration due to the earth’s gravity (g) for stiff soil/soft rock conditions at the site for the assumed background event MCE of moment magnitude

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(Mw) of 6.5 at a source-to-site distance of 12 km.  Relative to published results of probabilistic ground motion estimates, this represents a conservative level of ground motion with a recurrence interval much greater than 10,000 years.

Figure 24-22:  Seismic Regions of Mexico

24.5.3  Surface Water Hydrology

Based on Per NOM-141-SERMANAT-2003, the Project site is situated in the Humid Hydrologic Zone (Figure 24-23) and exhibits mountainous topography.  For the design, the hydrology of the TSF area was evaluated using the (Soil Conservation Service) SCS curve number method to model losses to soils and specify unit hydrograph transformation.  The SCS method relies on basin characteristics, design storm rainfall depths and temporal distribution to calculate volumetric flow rates.  Additional information on surface water hydrology calculations can be found in Sections 24.3 and 24.4 .

Figure 24-23:  Hydrologic Zones of Mexico

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24.5.4  Site Investigation & Site Conditions

Two geotechnical investigation programs took place for the project.  The first geotechnical investigation performed by Tetra Tech for the TSF location was based on a siting study performed between June 24 and 26, 2016.  Upon completion of the siting study, the investigation was geared towards two main sites and the identification of potential borrow areas.  The second investigation was performed by Tierra Group International (TGI) between August 13 to August 18, 2017 and October 30 to December 7, 2017.

24.5.4.1  Surface and Subsurface Conditions

In general, the TSF soil profile can be described as shallow.  In areas where soil is present, it typically consisted of 0.5 to 2.0 meters of silty sand (SM) and clayey sand with gravel (SC).  Upstream of the TSF, silty gravel with sand (GM) and sand with clayey sand (SP-SC & SW-SP) was encountered.  At the North end of the TSF, approximately 100 meters from the embankment toe, silty sand (SM) with gravel was encountered up to 19 meters in depth.  This deeper soil deposit may be attributed to the proximity of the Los Gatos fault line.  In areas along the foundation of the TSF, areas of exposed slightly weathered bedrock were visible with little or no topsoil.  Below the soil horizon, slightly weathered and highly fractured rock was encountered.  Typically, this layer is approximately one meter to 10 meters thick, depending on the location area.  Below this layer, fresh bedrock was encountered.  This rock is generally described as Andesite and Rhyolite.  Per TGI’s investigation, the permeability of the bedrock ranges from 1.24x10-6 to 1.53x10-8.  In addition, some boreholes Northwest and Southeast of the TSF encountered a potential shear zone consisting of moderately weathered rock with clay infilling.

24.5.5  Borrow Material

The borrow areas for investigation were identified by MPR.  They are located North of the proposed TSF embankment across the Santo Toribio River in areas of epiclastic sedimentary deposits.  Test pits in the potential borrow areas encountered silty gravel to clayey gravel to clayey sand with gravel (GC-GM & SC-GM).  The typical depth of soils in the test pits were approximately two meters.  Weathered bedrock was encountered below the soils.  The material is suitable for the purposes of construction, but additional investigation to confirm is required.  The location of the investigated areas can be found in Los Gatos Project Tailings Storage Facility & Waste Rock Facility Feasibility Design Report.

24.5.6  Tailings Dam Design

The TSF is designed to accommodate tailings, mine water, and tailings slurry water.  The capacity of the TSF is presented in Table 24-19.  The average dry density of the tailings is 1.55 tonnes/cubic meter (t/m3).  Based on this design, the TSF is estimated to hold approximately 7.8 Mt of dry tailings.

Table 24-19:  TSF Capacity

Description	Crest Elevation (masl)	Total Capacity (Mm3)
TSF	1,639	9.3

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Note: Volumes in the figure represent storage capacity to the indicated elevations (Tailings + Water + IDF + Freeboard).

These values should be updated during each stage of operation using a water balance model.

Figure 24-24:  TSF capacity curve

24.5.7  Staged Construction

The footprint of the TSF was stripped, grubbed and graded to the to the design grades and lines as shown in Figure 24-25, Figure 24-26, and Figure 24-27.  The ultimate TSF will be built in four stages using downstream construction methods.  These stages will be constructed with upstream slopes of 2H:1V and downstream of 2.5H:1V (horizontal: vertical).  Stage 1 Phase 1 has been completed.  A 13.8-m wide bench was constructed on the crest of Stage 1 and a final crest width of 15-m and a maximum crest elevation of 1,639 m, and a maximum capacity of 9.3 Mm3, considering a 2-meter freeboard.  Stage 1 (starter) was constructed to an elevation of 1618.6-m primarily from rock fill (Zone C), 1.5-m thick transition zone (Zone B), and 3-m thick filter zone (zone A).

24.5.8  Liner Design

The TSF impoundment has been designed with a 1.5 mm (60mil) high-density polyethylene (HDPE) geomembrane liner.  The liner shall extend along the base of the impoundment as well as the upstream slope of the embankment and shall be anchored along the edges of the current construction stage.  The liner is underlined by Geosynthetic Clay Line (GCL).  From top to bottom, the impoundment of the TSF consists of:

·HDPE geomembrane liner; and

·GCL

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24.5.9  Tailings Delivery System

Tailings from the process plant is pumped to a tailings distribution pipeline along the crest of the TSF through pressure rated high-density polyethylene (HDPE) pipelines ranging from 15 to 20 cm, at a nominal solids content of 54.7% by weight.  The tailings distribution system consists of a header and manifold system with controlled discharge of tailings through multiple spigots connected to a peripheral tailings distribution header pipeline.  The extent and location of the supernatant pond within the impoundment shall be controlled by selective operation of the spigots such that the pond location is constrained to the South-central area of the impoundment away from the embankment faces at all times during operation.

24.5.10  TSF Water Collection System

The collection system for the TSF impoundment consists of a network of underdrains for collecting subsurface water within the TSF footprint.  The various drainage features comprising the collection system are discussed further in the following sections.

24.5.11  Underdrain System

A network of drains was installed underneath the geomembrane liner to collect and convey subsurface water emanating from seeps and springs within the TSF footprint.  The underdrains were installed along predetermined channels generated during the subgrade grading of the TSF footprint.  The drainage network consists of 8-inch [200 mm] nominal diameter solid and or perforated pipe with drain gravel and geotextile wrap.  Flows are collected in the underdrain collection pond located on the East side of TSF.  The water quality of the flow collected in the underdrain collection pond is monitored regularly and the water is discharged directly to natural drainages.

24.5.12  Blanket Drain System

The incorporation of the blanket drain helped improve the stability of the embankment in the event of liner failure.  The blanket drain consists of a 1.0 m thick layer of drain material.

24.6  Surface Water Management

Based on the hydrology data and analysis described in the NI 43-101 report sections 24.3 and 24.4, HEC-HMS results are summarized below.  The peak flow rates were used to estimate channel dimensions, which are also presented below.  The diversion channels were designed as earthen lined for the TSF, with localized armoring of steep segments exhibiting high velocities.  The spillway structure is armored with rip rap.  Dimensions are provided in Table 24-20.

Table 24-20:  TSF Spillway & Channel Dimensions

Water Management Structure	Description	Bottom Width (m)	Depth (m)	Length (m)
TSF South Diversion	trapezoidal channel	4	3.5	2206
Mill Diversion	trapezoidal channel	2	1.4	678
TSF Emergency Spillway	trapezoidal channel	2	4	176
TSF East Diversion	trapezoidal channel	4	2.6	646

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24.6.1  Monitoring

A tailings facility operating manual and a monitoring and surveillance plan should be developed.  The monitoring plan should include measurements to confirm the condition of the embankment and foundation and the performance of the containment system.  Standpipe piezometers and survey monuments will be installed at each stage for ongoing monitoring of the stability and hydraulic conditions of the TSF.  Inclinometers and/or other monitoring instrumentation can be installed if severe movement occurs, as visually observed or by survey monument detection.

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Figure 24-25:  Underdrain for Stage 1, Phase 1

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Figure 24-26:  Tailing Deposition Plan — Stage 1, Phase 2 (Part 1)

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Figure 24-27:  Tailing Deposition Plan — Stage 1, Phase 2 (Part 2)

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25.0  INTERPRETATION AND CONCLUSIONS

25.1  Geology and Resources

Project geologic and drill hole data has been collected and analyzed by MPR using industry standard methods and practices and is sufficient to characterize grade and thicknesses of the deposit and to support the estimation of Measured, Indicated and Inferred Mineral Resources.  Although the deposit has been densely drilled, Resource expansion potential and project upside exist in the immediate deposit area as well as at other identified prospects such as the Amapola and Esther, which have been preliminarily investigated with drilling showing Indicated and Inferred Mineral Resources, and other prospects throughout the land package.

25.1.1  Data Verification

The current QA/QC program is adequate and supports the conclusion that data collected, and the monitoring of QC data is reliable for the purposes of estimating Resources; however, additional improvements are required for the QA/QC program to align with industry best practice and facilitate more meaningful QC.

Clerical corrections and optimization of standard reference material is necessary to assess laboratory analytical performance.  Standard performance for Ag, Pb, Zn, Au, and Cu are good, with few results outside of +/-2 standard deviations of the certified value.  To leverage the use of standards, material should be sourced closer to the range of the deposits average grade for Ag, Pb and Zn.  Most of the standards used are too low-grade Ag, Pb, and Zn.

Field duplicate testing has shown good reproducibility; however, current protocols do not adequately test the variability of the deposit within the likely mining areas.  Field duplicates contain too few ore-grade samples chosen from within the vein.  The field duplicates that have been analyzed and are above 100 g/t Ag show a similar range of variability as the sample pairs below 100 g/t Ag.  Collecting more field duplicates from within the expected mining area will help to evaluate the variability that could be encountered.

In-stream blank material analysis for Ag has demonstrated acceptable sample preparation and laboratory performance for Ag; the performance for Pb, Zn, and Cu show many samples with values several times the detection limit, and exceedances are not significant in relation to the average deposit grade.

Exceedances could be a result of background levels of Pb, Zn, and Cu in the blank material or a result of contamination from sample preparation or analysis.  Because the Ag blank testing has shown few failures, it is possible that the blank material contains base amounts of Pb, Zn, and Cu; however, the blank failures often visually correlate with preceding samples of higher-grade.

Umpire (third party) sampling should be conducted to meet industry standard practices to confirm the analyses performed at ALS Chemex.

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25.2  Mineral Reserve and Mine Plan

25.2.1  Mineral Reserve

The Mineral Reserve for Los Gatos is the economically mineable portion of the Measured and Indicated Mineral Resource that can justify economic extraction.  The Mineral Reserve defined herein includes the application of mining factors defined in this report for stope design criteria and dilution and recovery factors.  The Mineral Reserve is then supported by a mine plan that is based on detailed stope layouts.

25.2.2  Geotechnical Conclusion

The footwall rock and the ore being of fair to good qualities indicate little issue from a geomechanical point of view.  Proper planning and tactical measure such as perimeter blasting, respecting recommended stope size, minimizing capital development excavation size, and installing proposed ground support mitigates any issues encountered during the development and extraction of the Mineral Reserves.  The higher geomechanical risk comes from the presence of the Los Gatos fault in the hanging wall, creating a weak rock mass.  The weak hanging wall implies potential for dilution or uncontrolled failure of the stopes.  This can be mitigated with proper planning, control of the opening size, timely backfilling, and leaving protective pillars between the fault zone and the open stope (as required).

Operational data has indicated that ground stability is reached with ore pillars of approximately 2 meters.  In some places the geomechanical conditions are favorable enough to mine all the ore of the vein until the contact with the fault.  Some areas where the ground is less stable, a pillar of 1.0 m is left to avoid unstable conditions during operations.

25.2.3  Mine Plan

Based on the deposit geometry and anticipated geomechanical conditions, economic extraction of the Los Gatos Resource incorporates both longhole mining and drift-and-fill mining methods.  Modern trackless mobile equipment is being used for all development and mining activities.  The exploration decline from surface was extended to provide primary access and delivery of services.  The ramp is also used for haulage of the Mineral Reserves and waste from the underground operations.

Preproduction development has been completed and production has started underground.  Ramp up of mining continues to support the 2,500 tpd production rate.  Ongoing waste development to sustain the 2,500 tpd production rate averages approximately 211 m/month during the production period.

The life of mine has been scheduled at approximately 2,500 tpd for a total of 11 years.  Along with the Inferred Resources, there are indications of additional Resources along strike that, with additional drilling, may increase the Mineral Resources.

25.3  Mineral Processing and Metallurgy

The Cerro Los Gatos deposit is a silver, lead, and zinc resource concentration.  Lead and zinc occur primarily as galena and sphalerite, respectively.  Significant amount of willemite is identified in all tested samples.  Lead oxide minerals are also identified in some of the samples, especially from South East zone samples.  The existence of lead and zinc oxide minerals impacted their flotation performance.

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JK-drop weight and SMC test results show the samples are softer when compared to JKTech database while the SPI test results categorize the samples from moderately soft to moderately hard per SGS database.  The Ai and BWI tests results describe the samples as from abrasive to very abrasive and from moderately hard to hard.

Very good lead and silver flotation results have been achieved.  The final lead cleaner concentrates of the LCTs averaged 60.9% Pb at 5,404 g/t Ag at average recoveries 89% lead and 68.7% silver.  The zinc cleaner concentrates averaged 54.2% Zn at an average recovery of 66% due to the high willemite content.

The production plant has been constructed and adjustments will continue to optimize performance.  Additional floatation cells have been installed to remove fluorine from the concentrates.

25.4  Infrastructure

The power line installation to the Cerro Los Gatos Mining project site has been completed.  Power is supplied via a 115-kV utility transmission line.  This originates from the ‘San Francisco de Borja’ substation in Satevó (Chihuahua), where a 115-kV connection has recently been installed.  The generators used for power before the completion of the line remain onsite to provide backup power.

Site access roads were upgraded, and the river crossing has also been completed.  Buildings and facilities to support the mine operations have been completed, including offices, change areas, maintenance shops, and camp facilities to feed and house the staff.  The processing facility and its associated buildings are constructed and being utilized to produce concentrate.

The water supply is being sourced from the groundwater being pumped from the dewatering wells, which is then cooled and used on site.  Additional water is available from wells onsite.

25.5  Environmental and Social Impacts

The impact assessment conducted for the Project is compliant with the approach requested by the LGEEPA and the REIA as detailed below:

·Qualify the effect of impacts on the ecosystems, with regards to the relevance of possible effects on their functionality (Article 44, section II of the REIA).

·Develop this qualification in context of an ES and an area of influence of the project (Article 12, section IV of the REIA), so that the evaluation refers to both the ES and the PA where the project is intended to be located.

It was found, by the regulations established by the REIA, the ES integrated by the micro basin of the San José River will generate a non-relevant impact for the removal of 390.37 ha of vegetation.  In comparison to the surface area of the ES, these areas represent only 1.93% of the total area, showing the impact is not significant.  This ensures the function or continuity of ecosystem processes in the environmental system is not affected.

The conclusions of the environmental impact assessment indicate the functional integrity of the ecosystems is respected, since the relevant environmental components will not be significantly affected.  In the case of species under some category of risk, their areas of distribution are greater than the ES.  For water pollution, considered as a relevant impact, it is not planned to discharge process or mining water

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to natural effluents, even though it is considered to implement water treatment practices to be discharged to the tailings dam, which was built with liner and geo membranes; plus, recirculation of process water.

The level of conservation of regional biodiversity demonstrates the project will not cause any species to be declared as threatened or endangered, and the habitat of individuals of flora and fauna will not affect the species, according to Article 35-III, subsection b) of the LGEEPA.

Finally, as a result of the above conclusions the project will not generate significant and/or relevant effects to the ES, such as:

·Ecological imbalances;

·Damage to public health; and

·Irreversible effects to the ecosystems of the ES.

In addition, a set of measures designed for the prevention and mitigation of environmental impacts is included within the Environmental Management System (EMS) as part of the company’s commitments to prevent and mitigate the environmental impacts of the Project.

The following Federal, State and Municipal Permits issued for the Los Gatos Mining Operation have been applied for and issued by Federal, State and Municipal authorities as required for exploration, preparation, development and operations during the 2018 and 2019, have been obtained by Minera Plata Real:

·SEMARNAT — Environmental Impact Statement (Manifestación de Impacto Ambiental — Modalidad Regional) (MIA-R) issued in 2017, No. SGPA/DGIRA/DG/05121-2017.

·SEMARNAT 1st.  Modification to increase Impacted area from 211.0841 ha to 268.8450 ha and from 95 to 99 mine workings, No. SGPA/DGIRA/DG/01914, March 15, 2018.

·SEMARNAT 2nd.  Modification to increase Impacted area from 268.8450 ha to 325.0860 ha and from 99 to 133 mine workings, No. SGPA/DGIRA/DG/09272, November 28, 2018.

·SEMARNAT — Exemption to present MIA for Extension of Road from San José to Los Gatos mine.  No. SG.IR.08-2018/097, May 4, 2018.

·SEMARNAT — Modification of trajectory for “Power Line 115 KV Los Gatos”.  No. SG.IR.08-2018/093, May 4, 2018.

·SEMARNAT — Authorization of Preventive Report for Direct Mining Exploration, diamond Drilling in Los Gatos NW-CE-SE, Cascabel Fault and El Valle Vein”.  No. SG.IR.08-2019/070, May 7, 2019.

·CONAGUA — Residual waters discharge from Los Gatos into Santo Toribio Creek, 8.0 l/s. No. 06CHI141265/24FADL16, August 31, 2018.

·CONAGUA — Authorization for residual waters discharge from Los Gatos into Santo Toribio Creek, 8.0 l/s.  No. 06CHI141265/24FADL16, August 31, 2018.

·CONAGUA — Authorization for increment of residual waters discharge and change of point from Cerro Los Gatos into Santo Toribio Creek, 8.0 l/s to 120 l./s.  No. 06CHI141265/ 24FADL16, July 16, 2019.

·CONAGUA — GASIR — Authorization for construction and operation of tailings storage facilities No. 1 with capacity for 7.6 Mm3 to be built in four stages and a period of 9 years for construction.  No. 4494, January 18, 2019.

·SEMARNAT — Approval of Environmental Unique License (Licencia Ambiental Unica, LAU), for production of 2,500 tpd for MPR, No.  LAU-CHIH-001-2019, May 27, 2019.

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·DGGIMAR — SEMARNAT — Approval of Registry for Management of Dangerous Residues for production of 144.642 tonnes/year, No. 08-PMG-I-3405-2019, March 26, 2019.

·Secretaría de Desarrollo Urbano y Ecología (SEDUE) Chihuahua — Approval of Registry and Generating Corporation with Plan of Management for Residues with Special Handling, approval for production of 468.747 tonnes/year, April 9 and 11, 2019.

·Permits and Approvals by the Municipality of Satevó, Chihuahua, during the months of June and July 2018, of the following permits:

·Land use permit;

·Authorization and approval for the initiation of construction of mining workings and infrastructure; and

·Official alignment and number.

25.6  Mine Reclamation

A Closure and Reclamation Plan (CRP) for the Project was developed for mining operations.  The CRP includes approaches and plans to address the closure and reclamation of Project-related disturbance in accordance with regulatory requirements as discussed in Section 20, with sound scientific and engineering practices, and industry-standard practices for mine closure and reclamation.  The CRP also serves as the basis for estimating the cost of closing and reclaiming Project facilities and disturbed areas.

25.7  Economic Analysis

The economic model is presented on an unlevered, post-tax, present value (PV) basis.  Valuation estimates presented in this technical report should be adjusted for existing LGJV current liabilities, receivables and long-term indebtedness.  Economic results are summarized in Table 22-2.  The analysis suggests the following conclusions, assuming no gearing:

·Mine Life: 11 years;

·Pre-tax present value (PV5.0%): $764 million;

·Post-tax present value (PV5.0%): $653 million;

·Taxes Paid: $148 million;

·Sustaining project capital of $267 million; and

·Initial project capital of $316 million (completed 2019)..

25.8  Groundwater Hydrology/Dewatering

The modeling results suggest that adding dewatering wells at the modeled locations and pumping rates will intercept a portion of groundwater inflow to the mine before it reaches the underground workings.

Significant flows into the mine are expected to continue and will require continued in-mine dewatering.  However, adding wells or increasing the pumping rates at existing and already-planned wells will result in decreased flow into the mine.

Water pumped from the dewatering wells is used to meet on-site demand such as consumptive use in the mine, man camp domestic use, and process water makeup.  The excess water from the dewatering wells

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requires only cooling prior to discharge to the surface.  Water pumped directly from the mine requires treatment to reduce the suspended solids and possible hydrocarbons content prior to discharge.  Consequently, limiting flow into the mine reduces treatment requirements.

25.9  Surface Hydrology

Average annual runoff volumes were estimated for the TSF by applying a runoff coefficient of 0.14 to the average annual rainfall for each basin.

25.10  Tailings Management

The TSF design has been operated following the guidelines of the host country Mexico and other guidelines widely accepted by the mining industry such as the Canadian Dam Association and the International Large Dam Committee.

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26.0  RECOMMENDATIONS

26.1  Engineering, Procurement, and Construction Management

Detailed engineering has been completed.  Production has started at the mine and facilities have been completed to support the mine.  The plant is now constructed and in start-up mode.  Ramp up of mining and processing will continue until the target of 2,500 tpd is reached in Q2 2021.

26.2  Geology and Resources

Tetra Tech recommends the following to continue exploration and possible expansion of the Los Gatos Mineral Resource base within the property, as follows:

·Recognizance and in-fill drilling to test the following areas:

·Down-dip and along strike to the Northeast of the NW Block;

·Up-dip of the Central Block extending the high-grade mineralized zone along strike of the bulk sample area;

·At the hanging-wall detached blocks currently classified as Inferred Resources of the NW block;

·At the plunging mineralized concentration of the SE3 Block with additional potential of the Block’s extension; and

·Down-dip of the Central Block along the high-grade zone intercepted by drill holes GA-55, GA-66, and GA-243 to determine possible continuity;

·Additional infill drilling is recommended at the Amapola and Esther zones to delineate the possible extension of the identified Mineral Resources and assess full Resource potential.  If the results are positive, the identified Mineral Resources should be updated with a Scoping Study to determine if they may contribute to the Los Gatos economics;

·Complete detailed surface mapping and sampling in the area to define and prioritize other probable prospects within the project’s area;

·Geophysical surveys should complement the prospects prioritization prior to perform drilling exploration.

26.2.1  Standards

The following recommendations are to improve the current QA/QC protocols:

·Check sample standards relative to Ag, Pb, Zn, Cu, and Au. Correct assay records and database when errors are recognized;

·Clerical errors rather than actual analysis failure are the most frequent issue.  This can be corrected by comparing all five payable elements.  Spikes are visually observed when standard assays issues are encountered; and

·Keep records of sample results in the database for a quick review to confirm possible clerical errors.

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·Submit with more frequency ore-grade standards from the mineralized zone:

·Identify on the sample submittal sheet the ore-grade standards to be tested directly using the assay methods OG62 or GRA21.  This action will avoid falling short of the 30-grams required for re-runs by the ME-ICP61 and 50 grams required to complete the assays; and

·Include additional sulfide standards that are as close as possible to grades of 250 g/t Ag, 2.5% Pb and 6% Zn, while assays for Au and Cu are of secondary importance and should only be evaluated after the economic metals when selecting ideal standards.

·Generate several custom standards from material collected from the area of the bulk sample, 500, 50-gram standards could be made from 25 kilograms of material. Internationally certified laboratories offer custom standard services.

26.2.2  Blanks

Tetra Tech recommends improve blank samples use by:

·Inserting blanks more frequently after assaying high-grade samples to ensure the sample preparation equipment has been cleaned properly;

·Review geologic cross-sections to optimize the insertion of blanks within the vein intersections;

·Currently, the intersected veins show approximately 1,400 2 m intervals; while the insertion of blank samples represents about 20% (280 samples) of the sampling population.

·Investigate the use of a coarser certified blank so that crushing, splitting, and pulverizing equipment at the preparation facility are operating in a similar condition to a normal interval core sample;

·Submit blank samples directly for re-run testing to test OG62 or GRA21 equipment.  Present protocols do not test potential contamination of OG62 or GRA21 equipment because the first test never triggers re-runs;

·Record the sample ID of the sample tested before each of the blanks.  This will enable assessment of blanks in the context of possible contamination from the sample preparation and the sample analyzed before the blanks.  Any poor blank performance following high-grade samples should trigger re-runs of several samples following the high-grade sample;

·Source certified or self-certify blank material;

·Blanks obtained from core are best because the lab is blind to the control sample and both the laboratory preparation and analysis are checked.  Blanks could be sourced from splits of andesite that have been tested and returned detection limit results for Ag, Pb, Zn, Au, and Cu.  If blank core cannot be sourced andesite outcroppings should be considered;

·Untested “barren” full core should not be used as blank material because of the risk that it contains low background levels of Pb, Zn, and Cu as seen in the current blanks; and

·If certifiable blank core is limited, blanks can be submitted as reduced weight (not the equivalent of a 2 m core split), with only enough material to produce a coarse reject and pulp.

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·Ensure the blank material is stored away from the core preparation area and the blank material sample bagging is completed in a clean environment; and

·Request the laboratory’s internal blank control sample to potentially evaluate or rule out contamination.

26.2.3  Duplicates

To improve duplicate use for QA/QC Tetra Tech recommends:

·Insert instream field duplicates in areas that visually appear to be high-grade.  Doing so will test the prep lab’s ability to homogenize and provide data to evaluate the nugget effect; and

·Insert an empty bag with a sample ID tag and instruct the lab to generate a pulp split (pulp duplicate).  Analyzing pulp duplicates will provide data to evaluate the reproducibility of the sample analysis and suggest an error range of analysis.

26.2.4  Umpire Sampling

It is industry best practice to have 5% of sample pulps tested at a third-party lab.  Aside from limited re-sampling completed as part of past NI 43-101 reports, it is not apparent that umpire sampling has been performed.

As discussed, a small percentage of the total drill hole database encounters the area of interest, because of this umpire sampling should be focused on resampling most of the pulps within the expected mining areas.  Re-analyzing half of the pulps from all vein samples would account for approximately 750 samples and provide additional confidence in the database.

26.3  Mineral Reserve and Mine Planning

26.3.1  Mineral Reserve Estimate

The Los Gatos deposit, particularly in the CZ, consists of multiple sub-parallel veins, sometimes quite close to each other, resulting in both mine scheduling and geomechanical complexity.  Definition diamond drilling and close attention to extraction scheduling will help minimize dilution and the loss of Resources.

There is potential to increase mined grades through careful attention to scheduling, blasting, and backfilling practices to reduce dilution.

Increases to the Reserve, with resulting increases to the life of mine, may be achieved by the following.

·The conversion of Inferred Resources into Reserves through definition diamond drilling.

·The addition of new Resources through exploration diamond drilling.

·Mine production from areas not included in Reserves.

The Mineral Reserve estimate is based on assumptions concerning ground conditions, mining methods and recoveries, and economic parameters, (e.g., capital costs, operating costs, and metal prices).  Changes in any of these assumptions will have an impact on the Mineral Reserve.  It is recommended to revise the Mineral Reserves with the recent exploration results and throughout the mine life.

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26.3.2  Mining

The following key mining risks and mining opportunities have been noted.

·NSR values were assigned only to Measured and Indicated Resource blocks.  All other blocks in the model were assigned a zero-dollar value.  Stopes that include mining of Inferred Resources may see an increase in grade.

·Proximity to the Los Gatos fault, particularly in the CZ, may negatively impact ore recoveries or dilution.  A conservative approach has been taken, with no mining occurring within a 5 m offset from the fault zone.  Definition diamond drilling during level development will demonstrate if this approach is valid.  Additional geomechanical studies are recommended to better define ground control requirements.

·When more than two drifts are planned, consideration must be given to installing additional ground support, such as shotcrete posts in the first drift, to support the effective span created by driving the adjacent drifts.

·The recommended stope size is 20 m high (floor to floor) × 12 m wide (along strike) and 21 m long (hanging wall to footwall).  Hanging wall equivalent linear slough dilution is expected to be 2 m.  Additional measures could be used to minimize dilution, such as cable bolting the hanging wall, using pre-splitting blasting techniques along the hanging wall, backfilling stopes within 3 weeks from the first blast, and reducing the stope width (along strike).  If additional information from bulk sampling and exploration drilling proves positive, consideration may be given to increasing stope height to 25 m.

26.3.3  Mine Planning

The following recommendations are for mine planning:

·Complete a detailed mine design and schedule for the first 5 years of the project.  Look for opportunities to improve the average grade by selectively targeting higher-grade areas.  Include the recent exploration results into the mine plan.

·Review the exploration development necessary to increase the mineral inventory and incorporate it into the mine plan.

26.4  Metallurgy and Recovery Methods

The expected grades and recoveries for lead, zinc, and silver to individual flotation concentrates utilized in this report were further investigated by a pilot plant program using a sample composed of a bulk sample accessed by an underground decline into the orebody.  The full process plant is now in operation/startup mode.  Any adjustments to the plant design and operations conditions will be monitored and adjusted as required.  Additional floatation cells have been added to remove fluorine, and this aspect should continue to be carefully monitored and adjusted as needed.

In the current operations, no re-grinding is required (probably due to the lower than design thru put/ ramp-up operations) as the plant ramps up re-grind of rougher concentrates will occur.  The need or percentage of rougher concentrates requiring re-grind will be monitored.

It is recommended that a tradeoff study be completed to evaluate the feasibility of expanding production to 3,000 tpd.

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26.5  Environmental, Permitting, and Reclamation

Tetra Tech recommends updating the CRP and develop concurrent reclamation programs throughout the project life.

26.6  Geochemistry

A surface water monitoring station is recommended to be established below the waste rock storage facility.  The production of leachate should be monitored, and water samples be collected at the beginning and end of the wet season to evaluate leachate generation at the field scale, and to determine whether management is required.  The samples should be analyzed for pH, electrical conductivity, dissolved metals (antimony, arsenic, barium, beryllium, cadmium, chrome, copper, lead, mercury, nickel, selenium, silver).  Results should be compared to Maximum Permissible Limits for Contaminants in the Discharge of Wastewaters into National Waters and Resources, NOM-001-SEMARNAT-1996.

During operations, regular sampling of the tailings is recommended to be at a rate of 1 sample for every 1,000,000 tonnes.  Samples are to be analyzed for total and extractable metals (antimony, arsenic, barium, beryllium, cadmium, chrome, mercury, silver, lead, selenium) and for acid base accounting and results compared to Maximum Permissible Limits for Toxic Constituents as Total Elemental and Extractable Elemental Concentration, §5.4.2.3 and §5.4.2.4, NOM-157-SEMARNAT-2009, and Limits for Determining the Hazard Potential for Generating Acid, §5.4.2.6, NOM-157-SEMARNAT-2009.

Additionally, samples from the underdrains are recommended to be sampled quarterly for pH, electrical conductivity, dissolved metals (antimony, arsenic, barium, beryllium, cadmium, chrome, copper, lead, mercury, nickel, selenium, silver), with results to be compared to Maximum Permissible Limits for Contaminants in the Discharge of Wastewaters into National Waters and Resources, NOM-001-SEMARNAT-1996.

26.7  Tailings Management

Tetra Tech recommends the following for the next stage of the TSF construction:

·A seismic refraction investigation on the footprint of the TSF.

·Ongoing monitoring of wells for groundwater contamination;

·Development of a TSF operating manual and a monitoring and surveillance plan; and

·Consideration for additional monitoring evaluation such as vibrating wire piezometers and inclinometers.

·The performance of the TSF facility should be closely monitored during the initial production years.  The monitoring plan includes measurement of the quantity and quality of seepage emanating from the TSF facility.  If the results of the monitoring program are evaluated and the design requirements are met, no further modifications shall be taken.  However, if the monitoring produces information that may affect the intent of the design, the design shall be evaluated and modified accordingly.

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26.8  Surface Water Hydrology

MPR maintains a meteorological station at the Los Gatos site.  However, precipitation was not a measured parameter.  Precipitation monitoring should occur on site by placement of a tipping bucket or similar rain gauge.  Installation of a totalizer coupled with a temperature probe is recommended at the cooling tower outfall to ensure compliance throughout the through the operational and closure phases of the project.  Placing a staff gauge fitted with a high-water indicator within the Rio Santo Toribio is also recommended.

Minera Plata Real should maintain a database of the meteorological, stream flow and groundwater data collected on site.  Project models should be periodically updated as measured data accrue.

Given the level of uncertainty associated with the runoff coefficients, it is recommended that the runoff volumes presented in Table 24-13 be verified by installation and evaluation of a site streamflow gauge.  A greater emphasis should be placed on the design storm criteria, presented below.

26.9  Groundwater Hydrology/Dewatering

Tetra Tech recommends an active dewatering strategy to reduce groundwater inflows into the underground workings.  This strategy is composed of the following depressurization/dewatering methods:

·Install future dewatering wells as close as possible to the underground workings.  The farther the well is away from the workings, the less effective it will be in reducing inflows.

·Install future dewatering wells to depths approximately 20 meters below the deepest mine depth, to allow pump placement below the mine floor elevations.

·Perform capacity tests on selected wells to better identify potential pumping rates.  A standard protocol for such capacity testing should be developed.

·Install larger capacity pumps in dewatering wells to reduce mine inflows.

·Continue to use dewatering well design similar to the recently installed wells.

·When discrete water-bearing zones are intersected during mining or during drilling inside the mine, install horizontal or inclined coreholes for depressurization.  To limit the volume of water requiring treatment before discharge, water from depressurization coreholes could be handled separately from the general mine inflow.  The cost of separate handling should be reviewed to determine whether it would provide an economic benefit.

Because cooling and discharge would be technologically simpler, more reliable, and potentially less expensive than treating to remove suspended solids and possibly other components, a combination of increasing pumping rates and adding new wells could be used to further reduce mine inflows.  A detailed study of the groundwater should be completed for the life of mine.

26.10  Water Balance

It is recommended that the site-wide water balance be periodically updated as hydrologic data becomes available during mine development and initial production.  Actual mine inflows and pumping rates from the mine should be measured, as well as water consumption components.  Water quality data should be collected from water pumped from the mine for dewatering purposes.

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26.11  Reclamation and Closure

Recommendations to address information gaps and advance closure and reclamation planning, design, budgeting and implementation are as follows:

·Regularly collect mine rock samples (i.e., mine rock gobbed underground and brought to the surface) that are representative of the lithologic unit mined during advanced Project design and the production period and subject the samples to static and kinetic geochemical testing to assess the potential to degrade the quality of water in contact with waste rock (contact water) and potential mitigation measures (if needed).

·Regularly monitor TSF seepage and mine water discharge rates during advanced Project design and the production period.

·Regularly collect samples of TSF seepage and supernatant pond water and mine water during advanced Project design and the production period and analyze the samples for and compare the analytical results to the regulated contaminants in Tables 2 and 3 of NOM-001 to assess compliance the wastewater discharge standards.

·Predict TSF draindown quantity and quality following termination of tailings and mine water deposition.

·Identify, characterize and quantify of PGM borrow areas or substitute PGM sources to address the anticipated PGM deficit of approximately 27,000 BCM.

·Investigate the feasibility, efficacy and cost of installing an enhance evaporation system on or adjacent to the impounded surface of the TSF to hasten supernatant pond water elimination, tailings consolidation, and equipment access to the impounded surface of the TSF at closure.

·Investigate the type and quality of salt accumulation on the impounded surface of the TSF tailings surface following evaporation of the supernatant pond and the cover type and thickness that will be needed to successfully reclaim the impounded surface of the TSF given the potential for the capillary rise of soluble salt into the PGM.

·Revegetation test plots should be installed on cut or fill slopes where practical.  The test plots and concurrently reclaimed areas (if any) of the Project should be monitored to evaluate and confirm the performance of grading, stormwater drainage and erosion controls, and revegetation treatments.

·Conduct stability analyses of the TSF for the purposes of closure.

·The soil unit identified as Unit 22 in the soil survey may also be potential borrow material.  Two areas have already been identified, although more may be present.  A more detailed mapping of the upper portion of the Southern portion of the site may identify more such areas.

Results from the bulleted items immediately above should be used to update and calibrate the Feasibility Study GoldSim water balance model, estimates of waste rock and tailings contact, seepage and mine water quantity and quality, and the closure and reclamation strategies, plans and cost estimated provided in the CRP.

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26.12  Additional Work

It is recommended to continue exploration of the properties for additional Resources through drilling, surface mapping, and geophysical surveys.  Infill drilling should also be completed for the conversion of Resource to higher classification level.  If drilling results are favorable, it is recommended to complete a Resource update on the Amapola and Esther deposits.

It is recommended that an updated FS be completed for the Los Gatos site.  This study would include a tradeoff study of an expansion of production to 3,000 tpd.  The crushing and grinding circuits were designed with the capacity of 3,000 tpd, and additional Resources have been identified.  This new study should evaluate the additional cost for mine and flotation expansion.

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27.0  REFERENCES

Abrahamson, N.A. and W.J. Silva, 2008, Summary of the Abrahamson and Silva NGA Ground-Motion Relations, Earthquake Spectra, 24: 67-97.

Aguirre-Díaz, G.J. and McDowell, F.W., 1993; “Nature and Timing of Faulting and Synextensional Magmatism in the Southern Basin and Range, Central-Eastern Durango, Mexico.”  Geological Society of America Bulletin, Vol. 105, pp 1435-1444.

Arizona Department of Environmental Quality, Publication # TB 04-01 Arizona Guidance Manual, Best Available Demonstrated Control Technology (BADCT).

Baca Carreón, Julio César (1964), Informe Preliminar de Reconocimiento de los Lotes Mineros El Tren y Margarita, Servicio Geológico Mexicano Public archives report.

Barfield, B.J., Haan, C.T. and Hayes, J. C. 1994. Design Hydrology and Sedimentology for Small Catchments.

Barton, N., R. Lien, and J. Lunde. “Engineering Classifications of Rock Masses for the Design of Tunnel Support.”  Rock Mechanics, vol. 6, no. 6, 1974, pp. 189-236.

Behre Dolbear, (2011), (NI) 43-101 Independent Technical Report of the Los Gatos Exploration Project

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Boore, David M. and Atkinson, Gail M., 2008, Ground-Motion Prediction Equations for the Average Horizontal Component of PGA, PGV, and 5%-Damped PSA at Spectral Periods between 0.01 s and 10.0 s, Earthquake Spectra, 24: 99-138.

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Ferrari, Luca, Valencia-Moreno, Martin and Bryan, Scott (2005), Magmatismo y tectónica en la Sierra Madre Occidental y su relación con la evolución de la margen occidental de Norteamérica, Boletín de la Sociedad Geológica Mexicana Volumen Conmemorativo del Centenario Temas Selectos de la Geología Mexicana Tomo LVII, Núm. 3, 2005, P. 343-378

Goff, J.A., Bergman, E.A. and Solomon, S.C., 1987; “Earthquake Source Mechanisms and Transform Fault Tectonics in the Gulf of California.”  Journal of Geophysical Research, Vol. 92, No. B10, pp 10,485-10,510.

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Hernández Bedolla, Joel. 2015. Estudio hidrológico superficial, subterráneo e hidráulico “Los Gatos” Satevó, Chihuahua. September.

Hernández-Bedolla, M.C. Joel. 2015. Surface and Groundwater Hydrology and Hydraulic Study for “Los Gatos” Satevó , Chihuahua. September 2015.

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M3 Engineering & Technology Corporation, “Los Gatos Scoping Study, Chihuahua, Mexico, M3-PN145015 Los Gatos Final Report 2014-08-08,” prepared for Sunshine Silver Mines Corporation, August 8, 2008, 512 pp.

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28.0  DATE AND SIGNATURE PAGE

CERTIFICATE OF AUTHOR

Guillermo Dante Ramírez-Rodríguez, PhD, MMSAQP

Principal Mining Engineer of Tetra Tech

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

I, Guillermo Dante Ramírez-Rodríguez, PhD, MMSAQP, of Golden, Colorado do hereby certify:

a)I am a Principle Mining Engineer with Tetra Tech, Inc. with a business address of 350 Indiana St., Suite 500, Golden, CO 80401.

b)This certificate applies to the Technical Report titled “Los Gatos Project — Chihuahua, Mexico”, effective July 1, 2020 and issued July 2020.

c)I have a bachelor’s degree in Mining and Metallurgical Engineering from the University of Zacatecas School of Mines in Mexico, and a Master and Doctorate degrees in Mining and Earth Systems Engineering from the Colorado School of Mines, in the United States of America.  I am a QP member for the Mining and Metallurgical Society of America (Member No. 01372QP).  I have over 32 years of professional experience since my graduation in 1987 working for the mining industry in underground and opencast operations, and as a consultant.  During these years I have worked for major and mid-tier mining companies in different positions as supervisor, mine planning chief, and manager in hard rock mining operations.  As a consultant I have also provided consulting services to all varieties of mining operations including hard rock mining, ferrous and non-ferrous operations, precious metals, base metals and industrial minerals.  I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

d)I visited the property September 30-October 1, 2015, January 17, 2017 and August 20-21, 2019.

e)I am responsible for Sections 16, 19, and 22, as well as portions of Sections 1, 2, 15, 21, 25, 26, and 27.

f)I satisfy all the requirements of independence according to NI 43-101.

g)I have read NI 43-101, Form 43-101 F1, and the Companion Policy to NI 43-101 (43-101 CP) and this Technical Report has been prepared in compliance with NI 43-101, Form 43-101 F1, and 43-101 CP.

h)As of the effective date of the 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.

i)I consent to the filing of the Technical Report with any stock exchanges or other regulatory authority and any publication by them, including electronic publication in the public company files on the websites accessible by the public, of the Technical Report.

Dated July 30, 2020

“Guillermo Dante Ramírez-Rodríguez PhD, MMSAQP” - Signed

Signature of Qualified Person

Guillermo Dante Ramírez-Rodríguez PhD, MMSAQP

Print name of Qualified Person

335

CERTIFICATE OF AUTHOR

Leonel López C.P.G., Associate Principal Geologist of Tetra Tech

Principal Mining Consultant of Tetra Tech

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

I, Leonel López, CPG, SME-RM, of Golden, Colorado do hereby certify:

a)I am currently employed as an Associate of Tetra Tech located at 350 Indiana Street, Suite 500, Golden, Colorado 80401.

b)This certificate applies to the Technical Report titled “Los Gatos Project — Chihuahua, Mexico”, effective July 1, 2020 and issued July 2020.

c)I am a Professional Geologist (PG-2407) in the State of Wyoming, USA, a Certified Professional Geologist (CPG-08359) in the American Institute of Professional Geologists, an SME Founding Registered Member (#1943910), a registered Geological Engineer (Cédula Professional #1191) in the Universidad Nacional Autónoma de México, a member of the Society of Economic Geologists, and a member of the Geological Society of America.  I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

d)I graduated from the Universidad Nacional Autónoma de México with the title of Ingeniero Geólogo in 1966 and have taken numerous short courses in Economic Evaluation and Investment Decision Methods at Colorado School of Mines, and other technical subjects in related professional seminars.  I have practiced my profession continuously for over 50 years. My related experience includes: Georges Ordoñez Consultants, (1965-72)  re-discovery of historical Silver/Gold mining districts, including Hostotipaquillo, Jalisco: Mololoa, Monte del Favor and La Trini) and Meztli, Sonora; explorations for Frisco Mining, Co. in the San Francisco del Oro and Parral mining districts, and Piedras Verdes, Chihuahua; as Division Manager N Zone Exploration for Peñoles (1972-85) including explorations and technical support in operations such as Fresnillo (largest silver mine in the World), La Encantada, La Ciénega, Topia, Naica, La Negra; as Independent, Minera Staleón (1986-1988) Soyopa, Sonora, Tahuehueto, Durango, Guadalupe Los Reyes, Sinaloa and San Pedro Corralitos, Chihuahua; in Pincock, Allen & Holt (PAH), consultant (1988-1993) COMIBOL, evaluated all Bolivia mining properties and projects including the Potosí silver mine for the World Bank; INMINE — Nicaragua evaluated La Libertad, El Limón, Bonanza and other silver/gold properties; Luismin — Tayoltita-San Dimas mining district; Western Silver, El Peñasquito, Zacatecas; for First Majestic Silver, Managed and prepared NI 43-101 Technical Reports for all operations during a period of 5 years including: La Parrilla, La Encantada, Del Toro, Minera El Pilón, Jalisco; and acted as Expert Witness and Project Manager in Vancouver Court for Legal suit regarding the Bolaños mine, Jalisco which was ruled in favor of First Majestic Silver; and I have participated in numerous other TR for most commodities around the World, including Fe, Cu, Mo, Mn, Co, Ni, Li, Mineral Sand Deposit in India and Africa; Geoambiente Mining, co-founder (1993 — 2003) consulting and developing properties, Oro Uno, Venezuela, other properties in Perú, Lucma, Río Chicama; PAH-Runge (2003 — 2014), participated in technical evaluations around the World including major deposits in Perú (Minera Poderosa, Cobriza), Argentina (Bajo de la Alumbrera, Pirquitas and other mines), Brazil (Most Vale’s Fe, Ni, Co, Mn deposits), Canada Arcellor Mittal Fe deposits), USA (Arcellor Mittal Fe deposits), Australia (Novo Resources, etc.; Cardno, consulting (2015- 2016) and Tetra Tech (2016 to date) has participated in TR NI 43-101 preparation including: Bacís, Durango; Lluvia de Oro, Durango; Novo Resources, Australia; Los Reyes, Sinaloa, and others.

e)I visited the property November 29-30, 2018 and August 20-21, 2019.

f)I am responsible for Sections 20 and 23, as well as portions of Sections 1-12, 25, 26, and 27.

g)I satisfy all the requirements of independence according to NI 43-101.

h)I have read NI 43-101, Form 43-101 F1, and the Companion Policy to NI 43-101 (43-101 CP) and this Technical Report has been prepared in compliance with NI 43-101, Form 43-101 F1, and 43-101 CP.

336

i)As of the effective date of the 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.

j)I consent to the filing of the Technical Report with any stock exchanges or other regulatory authority and any publication by them, including electronic publication in the public company files on the websites accessible by the public, of the Technical Report.

Dated July 30, 2020

“Leonel López, CPG, SME-RM” - Signed

Signature of Qualified Person

Leonel López, CPG, SME-RM

Print name of Qualified Person

337

CERTIFICATE OF AUTHOR

Kira Lyn Johnson, MMSAQP

Senior Geological Engineer of Tetra Tech

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

I, Kira Lyn Johnson, MMSAQP, of Golden, Colorado do hereby certify:

a)I am a Senior Geological Engineer with Tetra Tech, Inc. with a business address of 350 Indiana St., Suite 500, Golden, CO 80401.

b)This certificate applies to the Technical Report titled “Los Gatos Project — Chihuahua, Mexico”, effective July 1, 2020 and issued July 2020.

c)I have a bachelor’s degree in Geological Engineering from South Dakota School of Mines and Technology.  I am a QP member for the Mining and Metallurgical Society of America (Member No.  01539).  I have worked on Resource Estimations since my graduation from the South Dakota School of Mines in 2007.  This includes a variety of commodities, including gold, silver, nickel, taconite, oil sands, coal, potash, phosphates, aggregates, and other industrial minerals.  I have over 12 years of professional experience, including nearly 8 years of consulting in the mining industry for Tetra Tech.  I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

d)I have inspected the property July 17-18, 2012 and August 20-21, 2019.

e)I am responsible for Sections 14, as well as portions of Sections 1-12, 15, 25, 26, and 27.

f)I satisfy all the requirements of independence according to NI 43-101.

g)I have read NI 43-101, Form 43-101 F1, and the Companion Policy to NI 43-101 (43-101 CP) and this Technical Report has been prepared in compliance with NI 43-101, Form 43-101 F1, and 43-101 CP.

h)As of the effective date of the 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.

i)I consent to the filing of the Technical Report with any stock exchanges or other regulatory authority and any publication by them, including electronic publication in the public company files on the websites accessible by the public, of the Technical Report.

Dated July 30, 2020

“Kira Lyn Johnson, MMSAQP” - Signed

Signature of Qualified Person

Kira Lyn Johnson, MMSAQP

Print name of Qualified Person

338

CERTIFICATE OF AUTHOR

Keith Thompson, CPG

Professional Geologist of Tetra Tech

1100 McCaslin Boulevard, Suite 150

Superior, Colorado 80027

Telephone:  (303) 664-4630

I, Keith Thompson, C.P.G, of Greeley, Colorado do hereby certify:

a)I am an Senior Hydrologist with Tetra Tech, Inc. with a business address of 1100 McCaslin Boulevard, Suite 150, Superior, CO 80027 USA.

b)This certificate applies to the Technical Report titled “Los Gatos Project — Chihuahua, Mexico”, effective July 1, 2020 and issued July 2020.

c)I am a graduate of Youngstown State University, where I earned a Bachelor of Science degree in Geology in 1975, and of the University of Wyoming, where I earned a Master of Science degree in Geology, with specialization in Hydrogeology, in 1979. I have worked as a hydrogeologist for forty-one years since my graduation from university, in the capacity of consulting hydrogeologist with engineering firms and as a research associate professor with the Montana College of Mineral Science and Technology, Montana Bureau of Mines and Geology (now Montana Tech of the University of Montana). I have worked with Tetra Tech for more than 25 years. My work throughout my career has consistently included mining projects. It has covered characterization of baseline hydrologic conditions, mine dewatering and water management studies, design of dewatering systems, prediction and evaluation of hydrologic impacts of mining projects, mine-closure hydrology; and optimization reviews for historical mine-site remediation projects. I am an active member of the American Institute of Professional Geologists and am a certified professional geologist  (C.P.G. #6005).  I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (the “Instrument”) and certify that by reason of my education, affiliation with a professional association (as defined in the instrument), and past relevant work experience I am a “Qualified Person” for purposes of the Instrument.

d)I inspected the property on February 23-28, 2015; July 10-14, 2015, December 12-15, 2015; January 29 — February 2, 2016, and September 17-21, 2018.

e)I am responsible for portions of Sections 1, 24, 25, 26, and 27.

f)I satisfy all the requirements of independence according to NI 43-101.

g)I have read NI 43-101, Form 43-101 F1, and the Companion Policy to NI 43-101 (43-101 CP) and this Technical Report has been prepared in compliance with NI 43-101, Form 43-101 F1, and 43-101 CP.

h)As of the effective date of the 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.

i)I consent to the filing of the Technical Report with any stock exchanges or other regulatory authority and any publication by them, including electronic publication in the public company files on the websites accessible by the public, of the Technical Report.

Dated July 30, 2020.

“Keith Thompson, CPG” - Signed

Signature of Qualified Person

Keith Thompson, CPG

Print name of Qualified Person

339

CERTIFICATE OF AUTHOR

Kenneth E. Smith, RMQP

Mineral Processing Engineer/Consultant of Tetra Tech

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

I, Kenneth E. Smith, RMQP., of Golden, Colorado., do hereby certify:

a)I am a Mineral Processing Engineer / consultant with Tetra Tech with a business address of 350 Indiana Street, Suite 500, Golden, CO.80401.

b)This certificate applies to the Technical Report titled “Los Gatos Project — Chihuahua, Mexico”, effective July 1, 2020 and issued July 2020.

c)I graduated with a degree of Bachelor of Science in Environmental Engineering / with a Chemistry minor from Colorado State University in 1980.  I have worked as a metallurgical engineer for a total of 40 years since my graduation.  My work experience includes Goldstrike Mine, Western States Minerals for 8 years, Nevada Goldfields for 3 years, Canyon Resources for 4 years, while concurrently working on various International and North American consulting projects.  After 20 years of field experience, I began working for EPCM companies for an additional 20+ years, for all phases of project infrastructure, design, commissioning, and startup.  Including polymetallics, leaching operations, uranium projects and operations, as well as waste water treatment plants.  I am a certified MSHA/Hazmat instructor for 30+ years.

d)I am a registered member (Q.P) of the Society of Mining, Metallurgy, and Exploration, Inc. (SME #3004100).  I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

e)I visited the property on January 17, 2017 and August 20-21, 2019

f)I have read the Instrument, and the parts of the Technical Report that I am responsible for have been prepared in compliance with the Instrument.

g)I am responsible for Sections 13, 17, 18, and portions of 1, 11, 21, 25, 26, and 27 of the Technical Report and I am the Qualified Person for matters relating to ore crushing, process plant design, infrastructure requirements and the capital and operating costs associated with these project components.

h)I satisfy all the requirements of independence according to NI 43-101.

i)As of the date of this certificate, to the best of my knowledge, information and belief, the parts of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

j)I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them, including electronic publication in the public company files on their websites assessable by the public.

Dated July 30, 2020.

“Kenneth E. Smith, RMQP” - Signed

Signature of Qualified Person

Kenneth E. Smith, RMQP

Print name of Qualified Person

340

CERTIFICATE OF AUTHOR

Luis Quirindongo, QP

Geological Engineer of Global Resource Engineering

600 Grant St. Suite 975,

Denver, Colorado 80203

Telephone: (303) 547-6587

I, Luis Quirindongo of Denver, Colorado, do hereby certify:

a)I am a Geological Engineer with Global Resource Engineering, with a business address of 600 Grant St., Suite 975, Denver, Colorado, 80203.

b)This certificate applies to the technical report entitled “Los Gatos Project — Chihuahua, Mexico”, effective on July 1, 2020 and issued July 2020.

c)I graduated with from the University of Puerto Rico — Mayagüez, PR with a Bachelors in Geological Science (1999) and from the Missouri University of Science and Technology — Rolla, MO with a Master of Science in Geological Engineering (2004).  I have been working as a Geological Engineer in the area of geotechnical evaluations for the last 18 years, including many international projects.  During this time, I have worked on projects related to the evaluation and design of mine waste management such as tailings management, heap leach pads, and development rock storage areas as a consultant with large consulting firms such as Golder Associates and Tetra Tech.  I am a Registered Member of the Society for Mining, Metallurgy, and Exploration (SME) (#4208172RM).  Based on the above, I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

d)I inspected the property on June 25 and 26, 2015.

e)I am responsible for portions of Section 1, 24, 25, 26, and 27 of the Technical Report.

f)I satisfy all the requirements of independence according to NI 43-101.

g)I conducted and supervised the geotechnical investigation performed during the feasibility design work.  Additionally, I authored the Feasibility Level Design Report for the facility.

h)I have read NI 43-101, Form 43-101 F1, and the Companion Policy to NI 43-101 (43-101 CP) and this Technical Report has been prepared in compliance with NI 43-101, Form 43-101 F1, and 43-101 CP.

i)As of the date of this certificate, to the best of my knowledge, information and belief, the parts of the Technical Report that I am responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading.

j)I consent to the filing of the Technical Report with any stock exchanges or other regulatory authority and any publication by them, including electronic publication in the public company files on the websites accessible by the public, of the Technical Report.

Dated July 30, 2020

“Luis Quirindongo, QP” - Signed

Signature of Qualified Person

Luis Quirindongo, QP

Print name of Qualified Person

341

CERTIFICATE OF AUTHOR

Max Johnson, PE

Senior Civil Engineer of Tetra Tech

350 Indiana Street, Suite 500

Golden, Colorado 80401

Telephone: (303) 217-5700

I, Max Christian Johnson, PE, of Golden, Colorado do hereby certify:

a)I am a Senior Civil Engineer with Tetra Tech, Inc. with a business address of 350 Indiana St., Suite 500, Golden, CO 80401.

b)This certificate applies to the Technical Report titled “Los Gatos Project Chihuahua, Mexico”, effective July 1, 2020 and issued July 2020.

c)I graduated with a bachelor’s degree in Civil Engineering from Colorado State University in Fort Collins, Colorado. My work immediately after college was focused in mine water management including site-wide and facility-specific water balances, hydrology, meteorological monitoring, and mine reclamation. I received my water resource focused Professional Engineer license (PE.0051790) in 2016 from the State of Colorado in the United States of America. I have over 8 years of professional experience. I have reviewed the definition of a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”), and meet the requirements based on my education, work experience, and affiliation with a professional association.

d)I have not inspected the property.  I have not had any prior involvement with this property.

e)I am responsible for portions of Sections 1, 24, and 26.

f)I satisfy all the requirements of independence according to NI 43-101.

g)I have read NI 43-101, Form 43-101 F1, and the Companion Policy to NI 43-101 (43-101 CP) and this Technical Report has been prepared in compliance with NI 43-101, Form 43-101 F1, and 43-101 CP.

h)As of the effective date of the 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.

i)I consent to the filing of the Technical Report with any stock exchanges or other regulatory authority and any publication by them, including electronic publication in the public company files on the websites accessible by the public, of the Technical Report.

Dated July 30, 2020.

“Max Johnson, PE” - Signed

Signature of Qualified Person

Max Christian Johnson, PE

Print name of Qualified Person

342