Exhibit 99.1

ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

DATE AND SIGNATURES PAGE

The effective date of this report is 5 November 2014. The effective date of the Escobal Mineral Resource Estimate is January 23, 2014. The effective date of the Escobal Mineral Reserve Estimate is July 1, 2014. See Appendix A, Feasibility Study Contributors and Professional Qualifications, for certificates of qualified persons. These certificates are considered the date and signature of this report in accordance with Form 43-101F1.

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ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT

FEASIBILITY STUDY

TABLE OF CONTENTS

SECTION				PAGE
DATE AND SIGNATURES PAGE				I
TABLE OF CONTENTS				II
LIST OF FIGURES AND ILLUSTRATIONS				X
LIST OF TABLES				XIII
1	SUMMARY			1
	1.1	PRINCIPAL FINDINGS		1
	1.2	PROPERTY DESCRIPTION AND LOCATION		1
	1.3	MINERAL TENURE, SURFACE RIGHTS, AND ROYALTIES		1
	1.4	PERMITS		2
	1.5	ENVIRONMENT		2
	1.6	GEOLOGY AND MINERALIZATION		3
	1.7	EXPLORATION STATUS		3
	1.8	DRILLING		3
	1.9	SAMPLE PREPARATION AND ANALYSIS		3
	1.10	DATA VERIFICATION		4
	1.11	MINERAL PROCESSING AND METALLURGICAL TESTING		4
	1.12	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES		4
		1.12.1	Mineral Resources	4
		1.12.2	Mineral Reserves	5
	1.13	MINING		6
	1.14	PROCESSING		7
	1.15	TAILINGS		9
	1.16	INFRASTRUCTURE		9
	1.17	TRANSPORTATION AND LOGISTICS		10
	1.18	RECLAMATION		10
	1.19	OPERATING COST ESTIMATE		10
	1.20	CAPITAL COST ESTIMATE		10
	1.21	FINANCIAL ANALYSIS		12
	1.22	CONCLUSIONS AND RECOMMENDATIONS		14

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

2	INTRODUCTION			15
	2.1	PURPOSE AND BASIS OF REPORT		15
	2.2	SOURCES OF INFORMATION		15
	2.3	QUALIFIED PERSONS AND SITE VISITS		15
	2.4	EFFECTIVE DATES		16
	2.5	UNITS AND ABBREVIATIONS		16
3	RELIANCE ON OTHER EXPERTS			19
	3.1	MINERAL TENURE		19
	3.2	SURFACE RIGHTS, ACCESS, AND PERMITTING		19
4	PROPERTY DESCRIPTION AND LOCATION			20
	4.1	LOCATION		20
	4.2	MINERAL TENURE AND TITLE		20
	4.3	SURFACE RIGHTS		22
	4.4	PROPERTY AGREEMENTS		24
	4.5	PERMITS		24
	4.6	ENVIRONMENTAL LIABILITIES AND MANAGEMENT		24
		4.6.1	Primary Watershed	25
		4.6.2	Dry Stack Tailings	25
		4.6.3	Lined Stormwater and Waste Facilities	25
		4.6.4	Concurrent Reclamation	25
		4.6.5	Process Water Recovery and Recycling	25
		4.6.6	Paste Backfill	25
		4.6.7	Geochemical Characterization	26
		4.6.8	Environmental Management Program	26
	4.7	RISKS TO ACCESS, TITLE OR OPERATIONS		26
	4.8	RECLAMATION		26
5	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			27
	5.1	ACCESSIBILITY		27
	5.2	CLIMATE		27
	5.3	LOCAL RESOURCES AND INFRASTRUCTURE		27
	5.4	EXISTING INFRASTRUCTURE		27
	5.5	PHYSIOGRAPHY		28
6	HISTORY			29
	6.1	INTRODUCTION		29
	6.2	GOLDCORP / ENTRE MARES 1996-2010		29
	6.3	TAHOE RESOURCES		29

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7	GEOLOGICAL SETTING AND MINERALIZATION			32
	7.1	REGIONAL GEOLOGY		32
	7.2	LOCAL AND PROPERTY GEOLOGY		33
	7.3	LITHOLOGY		34
	7.4	STRUCTURE		35
	7.5	MINERALIZATION		37
	7.6	ALTERATION		39
	7.7	ESCOBAL VEIN ZONES		40
	7.8	VEIN MODEL		42
8	DEPOSIT TYPES			44
9	EXPLORATION			46
	9.1	GEOPHYSICS		46
	9.2	GEOCHEMISTRY		46
10	DRILLING			50
	10.1	SURFACE DRILLING		54
		10.1.1	Drill Campaigns	54
	10.2	UNDERGROUND DRILLING		55
	10.3	DATA COLLECTION		56
		10.3.1	Drill Core Handling	56
		10.3.2	Drill Collar Surveys	56
		10.3.3	Downhole Surveys	57
		10.3.4	Geological Logging	57
		10.3.5	Geotechnical Logging	57
	10.4	DRILLING SUMMARY AND RESULTS		57
	10.5	CORE RECOVERY — METAL GRADE ANALYSES		58
		10.5.1	2010 Analyses — Surface Core Recovery versus Metal Grades	58
		10.5.2	2014 Analyses — Underground Core Recovery versus Silver Grades	60
11	SAMPLE PREPARATION, ANALYSES AND SECURITY			61
	11.1	SAMPLE METHOD AND APPROACH		61
	11.2	SAMPLE SECURITY		61
	11.3	LABORATORY SAMPLE PREPARATION		62
	11.4	LABORATORY ANALYSES		62
	11.5	QUALITY ASSURANCE / QUALITY CONTROL PROCEDURES		62
		11.5.1	Standard Reference Materials	62
		11.5.2	Blanks	63
		11.5.3	Check Assays	63

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		11.5.4	Duplicates	63
	11.6	CONCLUSIONS		64
12	DATA VERIFICATION			65
	12.1	DATABASE AUDIT		65
		12.1.1	Drill Collar Database	65
		12.1.2	Down-Hole Survey Database	65
		12.1.3	Assay Database	66
	12.2	SITE VISITS		66
		12.2.1	Drilling Operations	66
		12.2.2	Sampling and Logging Procedures	67
		12.2.3	Drill Collar Verification	67
		12.2.4	Verification Sampling	67
	12.3	QUALITY ASSURANCE AND QUALITY CONTROL		67
		12.3.1	Assay Standards	68
		12.3.2	Sample Blanks	70
		12.3.3	Duplicate Samples	71
		12.3.4	Re-Analyses of QA/QC Failures	73
	12.4	CONCLUSIONS AND RECOMMENDATIONS		74
13	MINERAL PROCESSING AND METALLURGICAL TESTING			76
	13.1	SAMPLING		77
	13.2	GRINDING TESTS		77
	13.3	GRINDABILITY TESTS		78
	13.4	REAGENT SCREENING TESTS		78
	13.5	DETERMINATION OF RECOVERIES AND REAGENT CONSUMPTIONS		81
	13.6	ESTIMATED METALLURGICAL RECOVERIES FOR PROCESS DESIGN		81
	13.7	POST-PLANT DESIGN METALLURGICAL TESTS		82
		13.7.1	Pilot Plant	82
		13.7.2	Flotation Tests	83
14	MINERAL RESOURCE ESTIMATES			84
	14.1	INTRODUCTION		84
	14.2	DATA		84
	14.3	DEPOSIT GEOLOGY PERTINENT TO RESOURCE MODELING		84
	14.4	GEOLOGIC MODEL		85
	14.5	MINERAL-DOMAIN GRADE MODELS		86
	14.6	DENSITY		91
	14.7	SAMPLE CODING AND COMPOSITING		92
	14.8	RESOURCE MODEL AND ESTIMATION		94

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	14.9	RESOURCE CLASSIFICATION		96
	14.10	MINERAL RESOURCES		97
	14.11	DISCUSSION, QUANTIFICATIONS, RISK, AND RECOMMENDATIONS		105
15	MINERAL RESERVE ESTIMATES			106
	15.1	MINERAL RESERVE CLASSIFICATION		106
		15.1.1	Mineral Reserve	106
		15.1.2	Probable Mineral Reserve	106
		15.1.3	Proven Mineral Reserve	106
	15.2	MINERAL RESERVE STATEMENT		107
	15.3	CUTOFF GRADE		108
	15.4	MINING SHAPES		110
	15.5	DILUTION AND RECOVERY ESTIMATES		111
16	MINING METHODS			116
	16.1	DEVELOPMENT AND PRODUCTION		116
	16.2	CURRENT STATUS		118
	16.3	GEOTECHNICAL CONSIDERATIONS		124
		16.3.1	Rock Mass Rating	124
		16.3.2	Geologic Structure	125
		16.3.3	Rock Strength Testing	125
		16.3.4	Stress Regime	125
		16.3.5	Probabilistic Stope Stability Analysis	126
		16.3.6	Installed Ground Support	126
		16.3.7	Ground Support QA/QC	127
		16.3.8	Backfill	127
	16.4	PUMPING		127
	16.5	VENTILATION		128
	16.6	STOPE DESIGN		129
	16.7	DILUTION AND RECOVERY ESTIMATES		132
	16.8	PASTE BACKFILL		139
	16.9	DEVELOPMENT DESIGN		140
	16.10	PRODUCTION SCHEDULE AND MINING RATES		142
	16.11	MINING EQUIPMENT AND INFRASTRUCTURE		149
	16.12	MINING WORK FORCE		151
17	RECOVERY METHODS			154
	17.1	INTRODUCTION		154
	17.2	PROCESS SUMMARY		154
	17.3	PROCESS PLANT DESCRIPTION		156

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		17.3.1	Primary Crushing	156
		17.3.2	Secondary and Tertiary Crushing	156
		17.3.3	Conveyance	157
		17.3.4	Grinding	158
		17.3.5	Flotation	160
		17.3.6	Concentrate Thickening, Filtering and Packaging	165
		17.3.7	Tailings Thickening, Filtration & Disposal	169
		17.3.8	Process Water	171
	17.4	CONCENTRATE PRODUCTION		172
	17.5	PLANT EXPANSION		175
18	PROJECT INFRASTRUCTURE			177
	18.1	SITE ACCESS		178
	18.2	TRANSPORTATION AND LOGISTICS		178
	18.3	ELECTRICAL POWER		179
	18.4	WATER AVAILABILITY		179
	18.5	SITE DRAINAGE		179
	18.6	TAILINGS AND WASTE ROCK FACILITY		179
		18.6.1	Overview	179
		18.6.2	Layout	181
		18.6.3	Ongoing Monitoring and Testing	183
19	MARKET STUDIES AND CONTRACTS			184
20	ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT			185
	20.1	ENVIRONMENTAL STUDIES		185
		20.1.1	Baseline Environmental Study	185
		20.1.2	Geochemical Characterization	185
	20.2	SITE MONITORING		188
		20.2.1	Air Quality Monitoring	189
		20.2.2	Water Quality	189
		20.2.3	Surface Water	189
		20.2.4	Ground Water	189
		20.2.5	Process and Other Industrial Use Water	189
		20.2.6	Sediment Sampling	189
		20.2.7	Vibration Monitoring	190
		20.2.8	Noise (Sound Pressure) Monitoring	190
		20.2.9	Rock Geochemistry (Acid Rock Drainage Monitoring)	190
		20.2.10	Waste Disposal	190
		20.2.11	Reagent Storage	190
	20.3	WASTE ROCK AND TAILINGS DISPOSAL		190
	20.4	PERMITS AND APPROVALS		190
		20.4.1	Exploration	190
		20.4.2	Mine Operations	191

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

	20.5	SOCIAL OR COMMUNITY IMPACTS		191
	20.6	RECLAMATION		193
		20.6.1	Concurrent Reclamation	193
		20.6.2	Final Reclamation and Closure	193
21	CAPITAL AND OPERATING COSTS			195
	21.1	CAPITAL COSTS		195
		21.1.1	Expansion Capital	195
	21.2	SUSTAINING CAPITAL		195
	21.3	DEVELOPMENT COSTS		197
	21.4	OPERATING COSTS		198
		21.4.1	Mining Costs	198
		21.4.2	Backfill Costs	200
		21.4.3	Processing Costs	200
		21.4.4	Surface Operation Costs	201
		21.4.5	General and Administrative	201
	21.5	ELECTRICAL POWER		202
22	ECONOMIC ANALYSIS			203
	22.1	INTRODUCTION		203
	22.2	MINE PRODUCTION STATISTICS		203
	22.3	PLANT PRODUCTION STATISTICS		203
		22.3.1	Smelter Return Factors	204
	22.4	CAPITAL EXPENDITURE		204
		22.4.1	Expansion and Sustaining Capital	204
		22.4.2	Working Capital	205
		22.4.3	Salvage Value	205
	22.5	REVENUE		205
	22.6	OPERATING COST		206
		22.6.1	Total Cash Cost	206
		22.6.2	Reclamation & Closure	206
	22.7	TAXATION		206
	22.8	PROJECT FINANCING		206
	22.9	NET INCOME AFTER TAX		207
	22.10	NPV		207
	22.11	IRR		207
23	ADJACENT PROPERTIES			212
24	OTHER RELEVANT DATA AND INFORMATION			215

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25	INTERPRETATION AND CONCLUSIONS	216
26	RECOMMENDATIONS	217
27	REFERENCES	219
APPENDIX A: FEASIBILITY STUDY CONTRIBUTORS AND PROFESSIONAL QUALIFICATIONS		221

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LIST OF FIGURES AND ILLUSTRATIONS

FIGURE	DESCRIPTION	PAGE
Figure 4-1:	Escobal mine Location Map	20
Figure 4-2:	Minera San Rafael Concessions — Active and Pending	21
Figure 4-3:	Surface Ownership Map	23
Figure 7-1:	Regional Geology	33
Figure 7-2:	Local Geology	34
Figure 7-3:	Interpretation of Central Escobal Vein (North-South Section — Looking East)	37
Figure 7-4:	Vein Episodes and Generalized Relationships	38
Figure 7-5:	Escobal Central Zone — Breccia with Red Proustite Bands (Drill Hole E08-110)	39
Figure 7-6:	Escobal Long Section (East-West Section — Looking North)	42
Figure 8-1:	Spatial Relationship of Intermediate Sulfidation Deposits (after Corbett, 2002)	45
Figure 9-1:	Soil and Rockchip Chemistry — Silver	47
Figure 9-2:	Soil and Rockchip Chemistry — Zinc	48
Figure 10-1:	Escobal Drill Hole Location Map	51
Figure 10-2:	Escobal Central Zone Cross Section 806800E	52
Figure 10-3:	Escobal East Zone Cross Section 807500E	53
Figure 10-4:	Surface Exploration Drilling	55
Figure 10-5:	Underground Infill Drilling	56
Figure 10-6:	Core Recovery — Silver Grade Comparison	59
Figure 10-7:	Underground Core Recovery — Silver Grade Comparison	60
Figure 12-1:	Control Chart for Silver in CDN-ME-7	68
Figure 12-2:	Control Chart, Silver Z Score All Standards	69
Figure 12-3:	Blanks in Inspectorate Gold Analyses	70
Figure 12-4:	Scatterplot, 2012 and 2013 Silver Check Assays vs. Originals	72
Figure 12-5:	Silver Relative Percent Difference (2012 and 2013 Check Assay vs. Original)	72
Figure 12-6:	Silver Absolute Relative Percent Difference (2012 and 2013 Check Assay vs. Original)	73
Figure 14-1:	Section 806300 - Escobal Central Zone Silver Geologic Model	88
Figure 14-2:	Section 806800 - Escobal Central Zone Silver Geologic Model	89
Figure 14-3:	Section 807500 - Escobal East Zone Silver Geologic Model	90
Figure 14-4:	Section 806300 - Escobal Central Zone Block Model: AgEq Block Grades	102
Figure 14-5:	Section 806800 - Escobal Central Zone Block Model: AgEq Block Grades	103
Figure 14-6:	Section 807500 - Escobal East Zone Block Model: AgEq Block Grades	104

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Figure 15-1:	Transverse Stope Shape Parameters	110
Figure 15-2:	Longitudinal Stope Shape Parameters	111
Figure 15-3:	Generalized Location of Resources and Reserves	113
Figure 16-1:	Generalized Longitudinal Longhole Stoping Method	117
Figure 16-2:	Generalized Transverse Longhole Stoping	118
Figure 16-3:	East Central Decline Portal Area	119
Figure 16-4:	West Central Decline Portal Area	119
Figure 16-5:	As-built Plan Map — 1365 to 1415 m Levels	120
Figure 16-6:	As-built Plan Map — 1340 m Level	121
Figure 16-7:	As-built Plan Map — 1315 m Level	121
Figure 16-8:	As-built Plan Map — 1290 m Level	122
Figure 16-9:	As-built Plan Map — 1265 m Level	122
Figure 16-10:	As-built Plan Map — 1240 m Level	123
Figure 16-11:	As-built Plan Map — 1215 m Level	123
Figure 16-12:	As-built Plan Map — 1190 m Level	124
Figure 16-13:	Generalized Surface Fan Installation	128
Figure 16-14:	Generalized Ventilation Network	129
Figure 16-15:	Transverse Stope Shape Parameters	130
Figure 16-16:	Longitudinal Stope Shape Parameters	130
Figure 16-17:	Escobal Reserve Stopes — Looking North	131
Figure 16-18:	Escobal Reserve Stopes — Oblique View	132
Figure 16-19:	Reserve Grade-Tonnage Curve	139
Figure 16-20:	Development Design — Looking North	141
Figure 16-21:	Development Design — Oblique View	142
Figure 16-22:	Complete Mine Design — Oblique View	142
Figure 16-23:	Development Quantities by Type and Period	143
Figure 16-24:	Production Quantities by Type and Period	145
Figure 16-25:	Planned Mine Status — 2015	147
Figure 16-26:	Planned Mine Status — 2020	147
Figure 16-27:	Planned Mine Status — 2025	148
Figure 16-28:	Planned Mine Status — 2030	148
Figure 16-29:	Planned Mine Status — End of Mine Life (2033)	149
Figure 17-1:	Escobal Process Flowsheet	155
Figure 17-2:	Escobal Mine — Primary Crushing Plant	156

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Figure 17-3:	Escobal Mine — Secondary & Tertiary Crushing Plant	158
Figure 17-4:	Escobal Mine — Ball Mill	159
Figure 17-5:	Escobal Mine — Flotation Plant	165
Figure 17-6:	Escobal Mine — Concentrate Pressure Filter	168
Figure 17-7:	Escobal Mine — Concentrate Packaging	168
Figure 17-8:	Escobal Mine — Concentrate Loading	169
Figure 17-9:	Escobal Mine — Tailings Filtration	170
Figure 17-10:	Escobal Mine — Tailings Conveyor to Dry Stack	170
Figure 17-11:	Escobal Mine — Dry Stack Radial Stacker	171
Figure 17-12:	Escobal Mine — Tailings Conveyor to Paste Plant	171
Figure 17-13:	Escobal Grade-Recovery Curves	174
Figure 18-1:	General Site Layout	177
Figure 18-2:	Site Map With Existing Facilities	178
Figure 18-3:	Cross Section of the Tailing Dry Stack (End of Mining)	181
Figure 18-4:	Dry Stack Layout (End of Construction)	182
Figure 18-5:	Ongoing Construction During Tailing Placement	183
Figure 18-6:	Example of Ongoing Reclamation Efforts	183
Figure 21-1:	Life of Mine Capital Expenditures	197
Figure 21-2:	Life of Mine Annual Operating Costs	198
Figure 23-1:	Regional Exploration Targets	213
Figure 23-2:	Escobal District Exploration Targets	214

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LIST OF TABLES

TABLE	DESCRIPTION	PAGE
Table 1-1:	Summary of Escobal Mineral Resources	5
Table 1-2:	Escobal Mineral Reserves	5
Table 1-3:	Escobal Mine Production — Life of Mine (Tonnes, Ounces and Pounds in 000s)	7
Table 1-4:	Mill Throughput and Concentrate Production through June 30, 2014	8
Table 1-5:	Payable Metal Recovery Comparison — Design Parameters vs. Q2 2014 Actuals	8
Table 1-6:	Life of Mine Payable Metal Recovery	8
Table 1-7:	Escobal Process Plant Operations — Life of Mine (Tonnes, Ounces and Pounds in 000s)	9
Table 1-8:	Total Operating Cost	10
Table 1-9:	Mine Expansion Capital Expenditures	11
Table 1-10:	Plant Expansion Capital Expenditures	11
Table 1-11:	Life of Mine Expansion and Sustaining Capital Summary	12
Table 1-12:	Sensitivity Analysis — NPV after Taxes	13
Table 2-1:	List of Qualified Persons	16
Table 2-2:	Terms and Abbreviations	16
Table 4-1:	San Rafael Concessions	21
Table 10-1:	Escobal Exploitation Concession Drilling (Project-to-Date through July 1, 2014)	50
Table 10-2:	Escobal Drilling Included in Resource Estimate (Drill Data Through January 23, 2014)	50
Table 12-1:	Summary Comparison of 2011 Original and Check Analyses	71
Table 13-1:	Master Composite Head Assay Results	77
Table 13-2:	P80 Versus Metal Recovery to the Lead Rougher Concentrate	77
Table 13-3:	Resulting Ball Mill Work Indices from Ball Mill Grindability Tests	78
Table 13-4:	Typical Metal Recovery to Lead Rougher Concentrate	78
Table 13-5:	Typical Metal Distribution in Rougher Flotation Products	79
Table 13-6:	Typical Metal Distribution in Rougher Flotation Products	80
Table 13-7:	Escobal Concentrator Operational Parameters	81
Table 13-8:	Reagent Consumptions	81
Table 13-9:	Overall Pilot Plant Results	82
Table 13-10:	Final Cleaned Lead and Zinc Concentrate Grades	82
Table 13-11:	Metal Distribution in Rougher Flotation Products (2014)	83
Table 14-1:	Mineral Domain Grade Populations	87
Table 14-2:	Lithology Density Values Used in Model	91

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Table 14-3:	Mineral Domain Density Values Used in Model	92
Table 14-4:	Escobal mineral Domain Assay Statistics	93
Table 14-5:	Escobal mineral Domain Composite Statistics	94
Table 14-6:	Escobal Estimation Parameters for Mineral Resources	95
Table 14-7:	Escobal Deposit Reported Resource	98
Table 14-8:	Escobal Deposit AgEq Resource Tabulation	99
Table 15-1:	Mineral Reserve Estimate	107
Table 15-2:	Cutoff Value Assumptions — Metal Prices	109
Table 15-3:	Cutoff Value Assumptions — Process Recoveries	109
Table 15-4:	Cutoff Value Assumptions —% Payable Metals in Concentrate	109
Table 15-5:	Cutoff Value Assumptions — Initial Operating Costs	109
Table 15-6:	Cutoff Value Assumptions — Transport & Treatment Charges	109
Table 15-7:	Stope Shape Parameters	110
Table 15-8:	Mineral Resources in Relation to the Mine Design	114
Table 15-9:	Mineral Reserve Estimate by Elevation	115
Table 16-1:	Summary of Development Completed (as of July 1, 2014)	119
Table 16-2:	Summary of Ore Production (as of July 1st, 2014)	120
Table 16-3:	Summary of Drill Core RMR Values	125
Table 16-4:	Stope Shape Parameters	129
Table 16-5:	Mineral Resources in Relation to the Mine Design	133
Table 16-6:	Mineral Resources in Relation to the Mine Design (Transverse Stopes Only)	134
Table 16-7:	Mineral Resources in Relation to the Mine Design (Longitudinal Stopes Only)	135
Table 16-8:	Mineral Reserve Estimate by Elevation	136
Table 16-9:	Transverse Stope Reserves by Elevation	137
Table 16-10:	Longitudinal Stope Reserves by Elevation	138
Table 16-11:	Required Development Quantities	141
Table 16-12:	Development Advance Rates	143
Table 16-13:	Total Development Quantities Scheduled	144
Table 16-14:	Development Quantities Scheduled — 2014 through 2023	144
Table 16-15:	Development Quantities Scheduled — 2024 through 2033	144
Table 16-16:	Total Production Quantities Scheduled	145
Table 16-17:	Production Quantities Scheduled — 2014 through 2023	146
Table 16-18:	Production Quantities Scheduled — 2024 through 2033	146
Table 16-19:	Mine Mobile Equipment List	150

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Table 16-20:	Summary of Mine Personnel	151
Table 16-21:	Summary of Mine Operations Personnel	151
Table 16-22:	Summary of Mine Services Personnel	151
Table 16-23:	Summary of Training Personnel	152
Table 16-24:	Summary of Backfill Services Personnel	152
Table 16-25:	Summary of Shotcrete Services Personnel	152
Table 16-26:	Summary of Mine & Surface Equipment Shop Personnel	152
Table 16-27:	Summary of Mine Electrical Services Personnel	152
Table 16-28:	Summary of Mine Engineering Personnel	153
Table 16-29:	Summary of Geology Personnel	153
Table 16-30:	Summary of Underground Drilling Personnel	153
Table 17-1:	Mill Throughput and Concentrate Production (Life of Mine Through June 30, 2014)	172
Table 17-2:	Mill Throughput and Concentrate Production Apr 2014 - Jun 2014	173
Table 17-3:	Payable Metal Recovery Comparison — Design Parameters vs. Q2 2014 Actuals	173
Table 17-4:	Plant Expansion Capital Expenditures	175
Table 17-5:	Tailings Pressure Filter Expansion Capital Expenditures	175
Table 19-1:	Average Payable Metal Percentages and Refining Charges	184
Table 20-1:	Acid-Base Accounting Results	186
Table 20-2:	Meteoric Water Mobility Procedure Results	186
Table 20-3:	Humidity Cell Effluent pH	187
Table 20-4:	Humidity Cell Effluent Analysis — Waste Rock (mg/L)	187
Table 20-5:	Humidity Cell Effluent Analysis — Tailings (mg/L)	188
Table 21-1:	Expansion Capital Costs	195
Table 21-2:	Life of Mine Sustaining Capital Costs	196
Table 21-3:	Development Costs (Materials & Consumables Only)	197
Table 21-4:	Escobal Mine Operating Costs	198
Table 21-5:	Mine Operating Costs by Work Area	199
Table 21-6:	Production Costs (Materials & Consumables Only)	199
Table 21-7:	Mine Operations Costs	200
Table 21-8:	Backfill Operating Costs	200
Table 21-9:	Process Operating Costs by Area	201
Table 21-10:	Surface Operations Operating Costs	201
Table 21-11:	General and Administrative Costs by Area	202
Table 21-12:	Labor Distribution	202

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Table 22-1:	Life of Mine Ore, Waste and Metal Grades	203
Table 22-2:	Metal Recovery Factors	203
Table 22-3:	Life of Mine Concentrate Summary	203
Table 22-4:	Smelter Return Factors	204
Table 22-5:	Expansion and Sustaining Capital Summary	205
Table 22-6:	Operating Cost	206
Table 22-7:	Sensitivity Analysis After Taxes (in Thousands of $)	208
Table 22-8:	Detail Financial Model	209

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LIST OF APPENDICES

APPENDIX	DESCRIPTION
A	Feasibility Study Contributors and Professional Qualifications
	·Certificate of Qualified Person (“QP”)

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1SUMMARY

1.1PRINCIPAL FINDINGS

Tahoe Resources Inc. (Tahoe or the Company), through its wholly owned subsidiary, Minera San Rafael, S.A. (MSR), owns and operates the Escobal mine in Guatemala. The Escobal deposit is an intermediate-sulfidation silver-gold-lead-zinc vein deposit which the Company is mining by underground longhole stoping methods. Processing by differential flotation produces precious metal-rich lead and zinc concentrates for sale to international smelter customers. The mine is currently operating at the initial design rate of 3,500 tonnes per day (t/d) and preparing to increase the production rate to 4,500 t/d.

The Escobal mine commenced production in the fourth quarter of 2013 and reached commercial production in January 2014. Through June 2014, the Escobal mine produced 12 million ounces of silver, 7,600 ounces of gold, 6,600 tonnes of lead and 7,300 tonnes of zinc in lead and zinc concentrates from mill feed averaging 581 grams per tonne (g/t) silver, 0.47 g/t gold, 1.0% lead and 1.4% zinc.

The Feasibility Study (Study) demonstrates the feasibility of the Escobal mine and supports the declaration of Proven and Probable reserves. The Study provides economic parameters for the Escobal mine from July 1, 2014 forward.

Highlights of this Study include:

·Measured and Indicated mineral resources of 434 million silver ounces at an average grade of 346 g/t.

·Inferred mineral resources of 9.3 million silver ounces at an average grade of 224 g/t.

·Proven and Probable mineral reserves of 31.4 million tonnes at an average silver grade of 347 g/t, containing 350.5 million silver ounces in the life of mine plan.

·Average annual production of 19.1 million silver ounces and 22.4 million silver equivalent ounces over the first 10 years of the mine life.

·As of July 1, 2014, capital costs remaining to expand the production rate from 3,500 to 4,500 t/d are estimated at $24.3 million. All expansion capital, and the life of mine sustaining capital, is expected to be funded by the Company’s existing cash balance and projected future cash flow from the Escobal mine.

·After tax net present value at a 5% discount rate of $1.52 billion at the base case metal prices of $18.00/oz silver, $1300.00/oz gold, $0.95/lb lead and $0.90/lb zinc.

1.2PROPERTY DESCRIPTION AND LOCATION

The Escobal mine is located in southeast Guatemala, approximately 40 km east-southeast of Guatemala City and three kilometers east of the town of San Rafael Las Flores in the Department of Santa Rosa. San Rafael Las Flores has a population of 3,500 people and is 70 km from Guatemala City by paved road. Access to the area is also possible from the northeast on a paved highway via the town of Mataquescuintla. San Rafael Las Flores’ population is 99.6% “Ladino”, i.e., of Hispanic origin and non-indigenous.

The local climate consists of two major seasons; a “rainy” season between May and November and a “dry” season between November and May. Annual precipitation averages 1,689 mm. Average temperatures vary between 14°C and 33.1°C.

1.3MINERAL TENURE, SURFACE RIGHTS, AND ROYALTIES

Tahoe acquired the Escobal deposit from Goldcorp through a transaction completed in June, 2010. Through MSR, the Company holds mineral rights to 150.7 km2 in Guatemala through exploration and exploitation concessions. MSR’s concession package includes the Escobal exploitation concession which covers the Escobal deposit and mine operation facilities, and is the subject of this Study.

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The Company purchased approximately 281 ha of surface rights which are sufficient for the area required for mining, processing plant and ancillary facilities, surface operations, and tailings and waste rock disposal. No additional surface area is required to expand the mine operations to 4,500 t/d as discussed in this Study. Land surrounding the mine area is privately owned by local farmers and used for growing coffee in the higher elevations and vegetables and other crops in the flatter low lying areas.

A significant portion of the production royalties paid by the Company are directed to the local communities. In addition to the one percent NSR royalty mandated by the Guatemalan mining law, the Company entered into a voluntary royalty agreement which commits the Company to pay an additional four percent net smelter return (NSR) royalty on the concentrates sold from the Escobal mine. Of the total royalty of five percent, two percent will benefit communities in the San Rafael municipality and one percent will benefit certain outlying municipalities in Santa Rosa and Jalapa departments. The remaining two percent is paid to the Guatemalan government. The Company also established a profit sharing program that provides a 0.5% NSR payment to an association of the former land owners of the Escobal mine property.

1.4PERMITS

The Escobal mine operations are conducted under an Environmental Impact Assessment (EIA) and an Exploitation License.

The Escobal EIA was approved by the Guatemalan Ministry of Environment and Natural Resources (MARN) in October 2011. Public disclosure and involvement was required and developed throughout each stage of the permitting process. MARN Resolution 3061-2011 allowed the Company to begin full construction of the mine, process plant and all other surface facilities. The Company files quarterly environmental monitoring reports with MARN as required by the Resolution.

Mineral production is licensed through the Guatemalan Ministry of Energy and Mines (MEM). The Escobal Exploitation License was granted by MEM in April 2013. The Company files annual reports with MEM as required by the license stipulations.

Environmental requirements for surface exploration activities are specified Resolution 4590-2008/ELER/CG, which was issued by MARN to Entre Mares in December 2008. These requirements and approvals were transferred from Entre Mares to MSR in September 2010 as specified in MARN Resolution 1918-2010/ECM/GB.

All other approvals, permits and licenses necessary to conduct the Company’s business in Guatemala are in place and current.

1.5ENVIRONMENT

The mandate from Tahoe is to meet or exceed the standards of sustainability and environmental management based on North American practice and regulation. No impacted waters and materials are directly discharged from the site. Impacted water is held in lined containment and treated, if necessary, prior to being released to the environment. The environmental management design includes the following:

·Dry stack tailings

·Lined storm water and waste facilities

·A concurrent reclamation program

·Process water recovery and recycling

·Process/contact water containment and treatment systems

·Underground paste backfill

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These environmental controls represent the state of the art in sustainable design. The Company has implemented a comprehensive environmental management plan to regularly and systematically monitor air quality, surface water and groundwater quality, stream sediment geochemistry, blast vibration, noise levels, waste rock and tailings geochemistry (ARD monitoring), waste disposal practices, reagent handling and storage, and reclamation and reforestation progress.

1.6GEOLOGY AND MINERALIZATION

The Escobal deposit is an intermediate-sulfidation fault-related vein system formed within Tertiary sedimentary and volcanic rocks within the Caribbean tectonic plate. The Escobal vein system hosts silver, gold, lead and zinc, with an associated epithermal suite of elements, within quartz and quartz-carbonate veins. Quartz veins and stockwork up to 50 m wide, with up to 10% sulfides, form at the core of the Escobal deposit and grade outward through silicification, quartz-sericite, argillic and propylitic alteration zones.

Drilling to date has identified continuous precious and base metal mineralization at Escobal over 2,400 m laterally and 1,200 m vertically in four zones; the East, Central, West/Margarito and East Extension zones. The vein system is oriented generally east-west, with variable dips. The East and East Extension zones dip to the south from 60° to 75° with recent drilling showing a change to a more vertical dip at depth. The majority of the mineralized structure(s) in the Central and Margarito zones dip from 60° to 75° to the north, steepening to near-vertical at depth. The upper eastern portion of the Central Zone dips 60° to 70° to the south as in the East Zone.

1.7EXPLORATION STATUS

Tahoe continues to explore the Escobal Exploitation Concession for the continuation of mineralization along strike of the Escobal vein structure, with exploration drilling being carried out principally from the surface. Drilling is also exploring for ancillary veins or vein splays associated with the Escobal vein. Drilling to test below the lower extents of the existing resource/reserve will be carried out from underground drill platforms once mine development reaches the necessary elevation.

Near-term exploration focus on the exploitation concession includes recently interpreted targets that may represent a fault offset of the western extent of the Escobal vein to the south-southwest. Work to develop regional drill targets within the Company’s extensive exploration concession package also continues.

1.8DRILLING

Drill targeting of the Escobal and ancillary veins has been conducted by Entre Mares and Tahoe from 2007 to the present, with 946 exploration and infill/definition drill holes totaling 249,392 m completed within the boundaries of the Escobal exploitation concession through July 1, 2014. Surface drilling was done using both contractor- and company-owned drill rigs. Beginning in 2012, infill and definition drilling from underground drill stations has been conducted using drills owned and operated by the Company.

Nearly all drilling at Escobal has been done by diamond drill core methods. Core recovery from surface and underground drilling averages 95% and 89%, respectively.

1.9SAMPLE PREPARATION AND ANALYSIS

BSI Inspectorate is the primary analytical laboratory for all of the Escobal drill sample preparation and analysis, with only minor exceptions. All samples have been prepared and analyzed using industry-standard practices suitable for the mineralization at Escobal. Entre Mares and Tahoe conducted quality assurance and quality control (QA/QC) programs throughout all of the drill campaigns at Escobal, which included check assaying, duplicate sample assaying at other laboratories, and the use of blind assay standards and assay blanks.

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The core sampling procedures, sample analyses, QA/QC procedures, and sample security have provided sample data that are of sufficient quality for use in the resource estimation.

1.10DATA VERIFICATION

Data verification was supervised by Paul Tietz, CPG, of Mine Development Associates (MDA, Reno, Nevada USA). Mr. Tietz conducted site visits to the Escobal property in 2010, 2012 and 2014, which included verifying drill locations in the field, reviewing sample handling and data collection procedures, verifying downhole survey data, and independent verification sampling of drill core. MDA also completed full audits of the project database, analysis of the QA/QC data, and study of core recovery and its relationship to metal grades. The results of these verification programs support the estimation of the Escobal resource and the assignment of Measured, Indicated and Inferred classifications.

1.11MINERAL PROCESSING AND METALLURGICAL TESTING

McClelland Laboratories (Sparks, Nevada, USA) conducted the initial metallurgical tests in 2009 on three drill core samples from the Escobal deposit. It was concluded from the results of the tests that differential lead/zinc flotation producing a high value lead concentrate containing most of the silver and gold and a salable lower value zinc concentrate was the optimum processing route.

In June 2010, FLSmidth Dawson Metallurgical Laboratories was contracted to conduct metallurgical testing on drill cores representative of the mineralization from the Escobal Project. The primary objective of the test program was to determine process design criteria for crushing, grinding and flotation for the Escobal sulfide deposit. Results of the differential flotation tests indicate that the Escobal sulfide mineralization will respond to widely used and proven mineral processing techniques. The test programs conducted to date show that good recoveries of gold, silver, lead and zinc and acceptable reagent consumptions can be obtained by using conventional lead-zinc differential flotation processes.

The design flotation feed consisted of a primary grind size of 80% passing 105 µm in the rougher flotation circuits and regrind size of 80% passing 37 µm in the cleaner flotation circuits. Expected recoveries from the sulfide mineral processing were 86.8% for silver, 75.1% for gold, 82.5% for lead, and 82.6% for zinc. These recoveries were used for plant design.

In 2013, Tahoe commissioned a metallurgical pilot plant to validate the flotation process selected for the Escobal project, evaluate reagent type and usage, and to produce samples of lead and zinc concentrates for marketing to potential smelter customers. Overall, the pilot plant validated the results of the prior flotation tests and confirmed the process design selected for the Escobal mine would produce highly marketable concentrates.

Additional flotation tests were conducted in 2014 on drill samples collected from new areas in the East Extension Zone, deep Central zone, and West/Margarito Zone. Although mineralization in each of these areas appeared to be same mineralogically as samples previously tested from the Escobal deposit, rougher flotation tests were completed as a check of metallurgical compatibility with the Escobal process plant design. Results from the flotation tests demonstrated that metal recovery from ore in each of the three of new areas is similar to previous flotation test results and alterations to the process design is not required.

1.12MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

1.12.1Mineral Resources

The Mineral Resource Estimate for the Escobal deposit contains 107.4 million ounces of silver classified as Measured resources, 326.5 million ounces of silver classified as Indicated resources and 9.3 million ounces of silver classified as Inferred resources, with significant amounts of gold, lead, and zinc reported in all resource categories. Table 1-1 is a summary of the Escobal mineral resources, using a cutoff grade of 130 g/t silver-equivalent.

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Table 1-1: Summary of Escobal Mineral Resources

	Tonnes	Silver	Gold	Lead		Silver	Gold	Lead
Classification	(M)	(g/t)	(g/t)	(%)	Zinc (%)	(Moz)	(koz)	(kt)	Zinc (kt)
Measured	6.5	511	0.40	0.91	1.59	107.4	85	59	104
Indicated	32.5	313	0.32	0.68	1.12	326.5	333	221	364
Meas + Ind	39.0	346	0.33	0.72	1.20	433.9	418	281	467
Inferred	1.4	224	1.24	0.25	0.47	9.3	50	3	6

The dataset used for resource estimation is comprised of 842 drill holes totaling 231,326 m. MDA modeled and estimated the Escobal deposit resources by refining the geologic model, evaluating the drill data statistically, interpreting mineral domains on cross sections and level plans, analyzing the modeled mineralization statistically to establish estimation parameters, and estimating silver, lead, gold, and zinc grades into a three-dimensional block model using inverse distance cubed (ID3).

Silver-equivalent values for the resources were calculated using metal prices of $22.00/oz Ag, $1,325/oz Au, $1.00/lb Pb, and $0.95/lb Zn. The effective date of the mineral resource estimate is January 23, 2014.

1.12.2Mineral Reserves

The Escobal Proven and Probable reserves total 31.4 million tonnes at average grades of 347 g/t silver, 0.33 g/t gold, 0.74% lead and 1.21% zinc containing 350.5 million ounces of silver, 335,600 ounces of gold, 232,100 tonnes of lead and 381,600 tonnes of zinc. A summary of the Escobal mineral reserve estimate is provided in Table 1-2.

Table 1-2: Escobal Mineral Reserves

	Tonnes	Silver				Silver	Gold
Classification	(M)	(g/t)	Gold (g/t)	Lead (%)	Zinc (%)	(Moz)	(koz)	Lead (kt)	Zinc (kt)
Proven Reserves	6.0	457	0.37	0.86	1.51	87.8	70.8	51.7	90.2
Probable Reserves	25.4	321	0.32	0.71	1.14	262.7	265.2	180.4	291.3
Total Proven & Probable Reserves	31.4	347	0.33	0.74	1.21	350.5	335.6	232.1	381.6

The Escobal mineral reserves were estimated by Matthew Blattman of Blattman Brothers Consulting LLC (Cypress, Texas USA). Blattman completed a mine design and schedule from the Measured and Indicated resources based on actual production mining and development methods and rates used at the Escobal mine.

Cut-off grades to define the mineral reserves were calculated from the NSR value of the ore minus the production cost to account for variability in mining method and metallurgical response. NSR value was determined using metal prices of US$22.50 per ounce silver, US$1,300.00 per ounce gold, US$0.95 per pound lead and US$0.90 per pound zinc. By using a slightly optimistic value for silver, the continuity of the ore-grade mineralization along the edges of the mineral deposit improved which provided for more realistic stope design. Mining, processing and general and administrative (G&A) costs, metallurgical performance and smelter contract rates from the Escobal mine, and engineering first-principles were used to derive operating costs and revenue.

Proven and Probable reserves include 31% dilution that takes into account internal and external mining dilution and dilution from paste backfill where applicable. Subeconomic material internal to the stope designs and external mining dilution account for approximately 20% and 9% of the dilution total, respectively. Paste backfill dilution accounts for

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about 2%. Resources within the mine plan classified as Inferred have been given metal grades of zero. Blattman acknowledges that lower dilution rates may be attainable when mining.

The effective date of the Escobal Mineral Reserve Estimate is July 1, 2014. The reserve model has been depleted to account for mineral resources extracted prior to this date. Mineral Reserves are inclusive of Mineral Resources.

1.13MINING

The Escobal mine is accessed via two primary declines (East Central and West Central ramps) which provide access to the Central Zone of the deposit. A third internal primary ramp is being driven into the East Zone from the East Central ramp. Access ramps are driven from the main ramp system to establish sublevel footwall laterals driven parallel to the vein on 25 m vertical intervals. Primary and secondary development headings are mined 5 m wide by 6 m high with arched backs. The primary ramps are typically driven at a maximum inclination of -15%. Mining is currently being done by transverse longhole stoping with future mining by a combination of transverse and longitudinal longhole stoping. The stopes are accessed from the footwall laterals.

Ore is hauled to the surface by truck to the ore stockpile, located proximal to the primary crusher. Development waste rock is hauled by truck to the surface and used as construction material for the dry stack tailings buttress.

Filtered tails from the process plant are combined with cement and water to make a structural fill for use as backfill underground. A paste backfill plant located on the surface produces backfill for delivery via piping into the mine for placement in the mined out stopes.

Underground development of the Escobal mine commenced in May 2011, with construction of the East Central and West Central decline portals; after which ramp development began. Through June 2014, approximately 17,000 m of mine development and 230 m of vertical development (ventilation raises) had been completed. As of June 30, 2014, the mine produced 786,551 tonnes of ore grading 566 g/t silver, 0.46 g/t gold, 1.01% lead and 1.39% zinc.

The life of mine plan as of July 1, 2014 forecasts the Escobal mine to produce a total of 31.4 million tonnes of ore at average grades of 347 g/t silver, 0.33 g/t gold, 0.74% lead and 1.21% zinc. The life of mine production by year is summarized in Table 1-3.

Expanding mine production to meet the 4,500 t/d mill throughput rate requires the purchase of additional underground equipment over the next two years to increase development and production capabilities, and the rebuilding of the paste backfill plant to provide for additional paste backfill capacity to meet production goals. The paste backfill plant is scheduled for commissioning in the first quarter of 2015.

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Table 1-3: Escobal Mine Production — Life of Mine (Tonnes, Ounces and Pounds in 000s)

						Ag	Au	Pb	Zn
Year	Tonnes	Ag g/t	Au g/t	Pb %	Zn %	Ounces	Ounces	lbs	lbs
2014 (H2)	658	542	0.38	0.74	1.27	11,466	8	10,744	18,442
2015	1,529	482	0.36	0.68	1.19	23,707	18	23,033	40,153
2016	1,624	441	0.31	0.65	1.13	23,007	16	23,377	40,622
2017	1,658	442	0.32	0.67	1.12	23,555	17	24,668	40,803
2018	1,642	442	0.32	0.67	1.13	23,337	17	24,399	40,780
2019	1,633	442	0.40	0.66	1.11	23,200	21	23,816	40,107
2020	1,633	442	0.61	0.82	1.39	23,202	32	29,553	50,063
2021	1,631	442	0.27	0.46	0.78	23,184	14	16,436	27,991
2022	1,619	398	0.34	0.61	1.01	20,706	18	21,894	36,181
2023	1,643	266	0.27	0.56	1.03	14,056	15	20,431	37,138
2024	1,650	278	0.34	0.63	1.04	14,719	18	22,995	37,728
2025	1,635	283	0.33	0.63	1.11	14,904	17	22,588	39,922
2026	1,647	332	0.41	0.75	1.31	17,596	22	27,259	47,583
2027	1,639	312	0.44	0.81	1.50	16,416	23	29,372	54,239
2028	1,642	272	0.37	0.86	1.47	14,378	20	31,056	53,368
2029	1,654	266	0.32	0.75	1.27	14,130	17	27,429	46,209
2030	1,638	243	0.28	0.87	1.41	12,817	15	31,291	50,823
2031	1,637	251	0.17	0.80	1.02	13,188	9	28,850	36,675
2032	1,660	246	0.17	1.07	1.27	13,118	9	39,233	46,472
2033	1,362	224	0.25	1.08	1.80	9,797	11	32,453	53,956
Total	31,433	347	0.33	0.74	1.21	350,484	336	510,876	839,255

1.14PROCESSING

Ore from the Escobal mine is processed by differential flotation producing lead concentrates with high precious metal (silver+gold) grades and zinc concentrates with a lesser precious metal component. The original design basis for the processing facility is 3,500 t/d of ore or 1.28 million tonnes per year; though the installed crushing, grinding, flotation and concentrate processing components were sized for the contemplated increased throughput rate of 4,500 t/d discussed in this report. Expansion to 4,500 t/d requires additional tailing filtering capacity and modifications and/or upgrades to ancillary mill components. The additional tailings filtration unit is scheduled to be commissioned in the first quarter of 2015; plant modifications and upgrades are expected to be complete mid-2015.

Mill commissioning was initiated in the second half of 2013 with the first metal concentrates produced on September 30, 2013. The Company declared commercial production in January 2014 with the completion of mill commissioning and continued to ramp up the mill throughput rate through the first half of 2014. The Escobal ore processing facility is now operating at design levels and averaged 3,708 t/d in June 2014 with metal recoveries generally meeting or exceeding design expectations. Concentrate production through June 30, 2014 is summarized in Table 1-4.

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Table 1-4: Mill Throughput and Concentrate Production through June 30, 2014

Mill Production	Mill Feed	Pb Concentrate	Zn Concentrate
Tonnes	743,996	14,804	14,053
Ag Grade g/t	581	23,959	1,257
Au Grade g/t	0.47	14.70	1.27
Pb Grade %	1.0	44.7	1.4
Zn Grade %	1.4	13.5	51.9
Ag Ounces	13,888,635	11,403,376	568,014
Au Ounces	11,243	6,999	573
Pb Tonnes	7,528	6,611	202
Zn Tonnes	10,385	1,991	7,292

Table 1-5 below compares payable metal recoveries predicted by the process design testwork and the actual payable metal recoveries at the Escobal mine from the second quarter of 2014. Silver and lead recovery in the lead concentrate was 2.5% and 8% higher than indicated by the flotation testwork. Gold recovery to the lead concentrate was 9% lower than predicted by the testwork. Zinc recovery to the zinc concentrate was 9% less than predicted by the testwork.

Table 1-5: Payable Metal Recovery Comparison — Design Parameters vs. Q2 2014 Actuals

									Metal Recovery				Metal Recovery
	Head Grade				Total Recovery				Pb Concentrate				Zn Concentrate
Metal	Design		Actual		Design		Actual		Design		Actual		Design		Actual
Ag	415	g/t	657	g/t	86.7	%	88.1	%	82.5	%	85.0	%	4.2	%	3.1	%
Au	0.47	g/t	0.44	g/t	61.5	%	66.0	%	71.0	%	61.6	%	4.1	%	4.3	%
Pb	0.72	%	1.22	%	85.5	%	90.6	%	82.5	%	90.6	%	—		—
Zn	1.23	%	1.65	%	82.6	%	73.5	%	—		—		82.6	%	73.5	%

Grade-recovery curves for each metal and each concentrate were created from the current Escobal process recovery data to predict future metal recoveries in this Study. The average life of mine metal recoveries are summarized in Table 1-6.

Table 1-6: Life of Mine Payable Metal Recovery

		Lead	Zinc
Metal	Total	Concentrate	Concentrate
Ag	84.6	81.3	3.3
Au	60.9	58.3	2.5
Pb	91.4	89.8	—
Zn	87.9	—	75.7

*totals may not sum due to rounding

The Escobal mine is scheduled to produce a total of 495,000 tonnes of lead concentrate containing 285,265,000 ounces of silver, 196,000 ounces of gold and 459 million pounds of lead; and 551,000 tonnes of zinc concentrate containing 11,739,000 ounces of silver, 9,000 ounces of gold and 636 million pounds of zinc. Table 1-7 summarizes the life of mine process plant throughput schedule and metal production (metal recovered in concentrate).

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Table 1-7: Escobal Process Plant Operations — Life of Mine (Tonnes, Ounces and Pounds in 000s)

	Throughput						Metal Recovered in Concentrates
							Ag	Au	Pb	Zn
Year	Tonnes		Ag g/t	Au g/t	Pb %	Zn %	Ounces	Ounces	lbs	lbs
2014 (H2)	684	*	527	0.37	0.76	1.27	10,138	5	10,243	14,514
2015	1489		483	0.36	0.68	1.19	20,074	11	19,979	29,550
2016	1,643		442	0.31	0.65	1.14	20,133	10	20,978	30,953
2017	1,643		442	0.32	0.67	1.12	20,130	10	21,704	30,376
2018	1,643		442	0.32	0.67	1.13	20,131	10	21,682	30,644
2019	1,643		442	0.40	0.66	1.11	20,132	13	21,266	30,310
2020	1,643		442	0.60	0.82	1.38	20,133	21	26,770	38,330
2021	1,643		442	0.27	0.46	0.79	20,136	8	14,431	20,889
2022	1,643		399	0.34	0.61	1.01	17,987	11	19,486	27,247
2023	1,643		266	0.27	0.56	1.03	11,573	8	17,845	27,686
2024	1,643		278	0.34	0.63	1.04	12,113	11	20,221	28,034
2025	1,643		283	0.33	0.63	1.11	12,393	11	20,017	30,081
2026	1,643		332	0.41	0.75	1.31	14,745	14	24,406	36,135
2027	1,643		312	0.44	0.81	1.50	13,749	15	26,632	41,865
2028	1,643		272	0.37	0.86	1.47	11,870	12	28,253	41,074
2029	1,643		266	0.32	0.75	1.27	11,549	10	24,460	38,844
2030	1,643		244	0.28	0.87	1.41	10,518	9	28,534	39,033
2031	1,643		251	0.17	0.80	1.02	10,833	5	26,155	27,454
2032	1,643		246	0.17	1.07	1.27	10,661	5	35,924	34,922
2033	1,381		224	0.25	1.08	1.79	8,067	6	30,471	42,509
Total	31,476		347	0.33	0.74	1.21	297,004	204	459,459	636,449

*includes stockpiled ore as of July 1, 2014

1.15TAILINGS

The Escobal mine tailings facility is designed and operated as a dry-stack, in which dewatered tailings are spread, compacted and graded for erosion control and stability. Dry-stacking of tailings was selected and is implemented at the Escobal mine as it is an effective way to create a safe facility that will, upon closure, become a long-term stable geomorphic form in the landscape. Approximately 50% of the tailings will be dry stacked, with the remainder of the tailings returned underground as paste fill.

1.16INFRASTRUCTURE

The project is approximately 2 km from San Rafael Las Flores, a town of 3,500 people, and approximately 70 km by paved highway from Guatemala City. All year access to the area is good via paved highway from Guatemala City.

Power is provided by on-site diesel generation capable of sustaining approximately 21.5 MW of power. Normal operating requirement for mining, process and surface operations is 7.5 to 8.5 MW. The estimated peak power load required for the operation is approximately 12 MW during startup of high-consumption equipment such as the ball mill. The Company is actively investigating alternative lower-cost sources of power.

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All process and domestic water for the operation is supplied from mine dewatering and dedicated domestic water wells, respectively. Hydrological studies indicate sufficient water will be available to supply process and potable water requirements for the life of mine.

1.17TRANSPORTATION AND LOGISTICS

The major process and mining equipment was procured overseas and shipped to Guatemala. Guatemala has ports on both the Pacific and the Caribbean coasts. Access to the mine site from both ports is by paved highway.

Filtered concentrate is packaged in one- to two-tonne super-sacks, placed in sea-going containers, and carried on highway tractor trailer units along paved highways to port for shipment to international smelter customers.

1.18RECLAMATION

The entire facility was designed with closure in mind to the greatest extent practicable. The facilities are designed and operated to minimize the footprints and areas of disturbance and utilized the most advanced planning and reclamation techniques available including dry stack tailings, concurrent reclamation and geomorphic landform grading.

The disturbance footprint of the Escobal mine site is approximately 100 ha. Reclamation will commence as soon as practical during operations by placing salvaged topsoil on outslopes and encouraging vegetation. Final reclamation of the top surface will occur at final closure at the end of mine life. The front slope of the tailing dry stack is reclaimed concurrently as new lifts are added.

1.19OPERATING COST ESTIMATE

The operating costs for the Escobal mine were calculated for each year during the life of mine using the annual production tonnages and the actual mining, milling, and site General and Administration (G&A) costs from the mine’s initial year of production as the basis. Table 1-8 summarizes the Escobal mine average life of mine operating costs.

Table 1-8: Total Operating Cost

Operating Cost	$/ore tonne
Mine	$	37.23
Process Plant	$	22.83
General Administration	$	15.06
Production Cost	$	75.13
Smelting/Refining Treatment	$	26.64
Total Operating Cost	$	101.77

*Figures may not sum due to rounding

1.20CAPITAL COST ESTIMATE

Estimated capital expenditures for the Escobal mine are summarized in Table 1-9 and Table 1-10 for the mine and plant, respectively. Expansion capital accounts for approximately 30% of the total capital expenditures in the second half of 2014 and in 2015 and about 20% of the total capital expenditures in 2016.

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Table 1-9: Mine Expansion Capital Expenditures

Area	Cost Estimate
Underground Equipment	$	9,954,500
Paste Plant Expansion	$	6,900,000
Expended to Date*	$	(4,075,500	)
Total	$	12,779,000

*expended prior to July 1, 2014

Table 1-10: Plant Expansion Capital Expenditures

Area	Cost Estimate
Primary Crushing	$	649,900
Secondary/Tertiary Crushing	$	137,000
Grinding	$	473,300
Flotation	$	2,108,800
Concentrate Thickener	$	178,600
Tailings Filtration	$	9,205,200
Expended to Date*	$	(1,693,300	)
Tailings Dry Stack	$	475,000
Total	$	11,534,500

*expended prior to July 1, 2014

Life of mine sustaining capital totals approximately $301.2M expended over 19.5 years. About 50% of the sustaining capital is for primary underground development; the remaining 50% is divided between mine infrastructure and utilities, underground mobile equipment purchases and rebuilds, surface equipment, annual ball mill liner replacement, and miscellaneous site G&A. Total life of mine capital, including expansion capital, totals $325.5M.

The total capital carried in the financial model for the Escobal mine is shown in Table 1-11.

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Table 1-11: Life of Mine Expansion and Sustaining Capital Summary

Year	Cost Estimate (000s)
2014 Expansion*	$	5,549
Sustaining*	$	14,442
2015 Expansion	$	14,220
Sustaining	$	30,757
2016 Expansion	$	4,545
Sustaining	$	17,831
2017 Sustaining	$	19,117
2018 Sustaining	$	20,194
2019 Sustaining	$	23,258
2020 Sustaining	$	12,360
2021 Sustaining	$	12,151
2022 Sustaining	$	17,054
2023 Sustaining	$	18,301
2024 Sustaining	$	25,480
2025 Sustaining	$	15,476
2026 Sustaining	$	15,640
2027 Sustaining	$	12,851
2028 Sustaining	$	11,334
2029 Sustaining	$	13,828
2030 Sustaining	$	9,419
2031 Sustaining	$	5,205
2032 Sustaining	$	3,666
2033 Sustaining	$	2,814
Total	$	325,493

*July—December 2014

Expansion and sustaining capital is expected to be financed from cash flow from the Escobal mine production. Acquisition cost or expenditures prior to the July 2014 have been treated as “sunk” cost and are not included in the above summaries.

1.21FINANCIAL ANALYSIS

The Escobal mine economic analysis indicates the mine has an NPV5% of $1.52 billion using the base case metal prices of $18.00/oz silver, $1,300.00/oz gold, $0.95/lb lead and $0.90/lb zinc. The financial analysis does not present an internal rate of return (IRR), as it is a less meaningful metric for this study of the Escobal mine. In this study, all capital costs prior to July 1, 2014, are considered to be sunken. The calculated return on investment is magnified by the significant mine cash flows offset by the exclusion of the initial capital investment in the calculation and would exaggerate any IRR calculated. Net Present Value provides a more accurate and meaningful economic assessment of the mine.

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Sensitivity analyses were done using changes to all metal prices, silver prices only (no change to base case gold, lead and zinc prices), operating costs, capital costs and metallurgical recovery; the results of which are summarized in Table 1-12.

Table 1-12: Sensitivity Analysis — NPV after Taxes

Change in Metal Prices		NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case		$	2,012,550	$	1,515,689	$	1,214,041
	20%	$	3,100,197	$	2,288,349	$	1,805,971
	10%	$	2,556,373	$	1,902,019	$	1,510,006
	0%	$	2,012,550	$	1,515,689	$	1,214,041
	-10%	$	1,526,811	$	1,160,056	$	935,374
	-20%	$	1,029,418	$	800,057	$	654,775

Change in Silver Prices			NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case			$	2,012,550	$	1,515,689	$	1,214,041
	$	25.00	$	3,740,225	$	2,761,565	$	2,179,556
	$	22.00	$	2,999,793	$	2,227,618	$	1,765,764
	$	20.00	$	2,506,171	$	1,871,653	$	1,489,902
	$	18.00	$	2,012,550	$	1,515,689	$	1,214,041
	$	15.00	$	1,329,003	$	1,012,566	$	818,096

Change in Operating Cost		NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case		$	2,012,550	$	1,515,689	$	1,214,041
	20%	$	1,605,357	$	1,228,547	$	995,688
	10%	$	1,776,080	$	1,354,731	$	1,095,059
	0%	$	2,012,550	$	1,515,689	$	1,214,041
	-10%	$	2,249,020	$	1,676,647	$	1,333,023
	-20%	$	2,485,489	$	1,837,605	$	1,452,005

Change in Initial Capital		NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case		$	2,012,550	$	1,515,689	$	1,214,041
	20%	$	2,007,687	$	1,510,869	$	1,209,261
	10%	$	2,010,119	$	1,513,279	$	1,211,651
	0%	$	2,012,550	$	1,515,689	$	1,214,041
	-10%	$	2,014,981	$	1,518,099	$	1,216,431
	-20%	$	2,017,413	$	1,520,508	$	1,218,821

Change in Recovery		NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case		$	2,012,550	$	1,515,689	$	1,214,041
	2.0%	$	2,106,005	$	1,582,562	$	1,265,556
	1.0%	$	2,059,277	$	1,549,125	$	1,239,799
	0.0%	$	2,012,550	$	1,515,689	$	1,214,041
	-1.0%	$	1,965,823	$	1,482,253	$	1,188,284
	-2.0%	$	1,919,095	$	1,448,816	$	1,162,526

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1.22CONCLUSIONS AND RECOMMENDATIONS

The results of this Study demonstrate that:

1.Proven and Probable reserves are supported by the feasibility of the Escobal mine.

2.Operating results to date validate the mining method and process design for the Escobal mine.

3.Mine and mill expansion from the current rate to 4,500 t/d is achievable with minimal additional capital expenditure.

M3 recommends:

1.Tahoe continue to aggressively explore for extensions or offsets of the Escobal vein, or to accelerate its district exploration programs, to provide higher mining grades beyond 2022.

2.Concurrent with the prior recommendation, initiate studies to investigate increased mining and throughput rates to grow the annual silver production in the second half of the mine life.

3.Tahoe continue investigating lower cost alternatives to the current diesel generated power at the mine site. M3 believes lower power costs have the opportunity to provide significant operating cost savings in the near term.

4.Critical evaluation of operating and capital costs. Now that the Escobal mine has reached design operational parameters, Tahoe should begin to focus on optimizing mining and processing procedures in an effort to reduce operating costs. Reductions in primary development mining costs and increased equipment utilization would have a direct positive impact on the life of mine sustaining capital requirements.

5.Metallurgical studies to determine if silver and gold metallurgical recoveries can be improved as incremental increases in metal recovery, particularly for silver, may have a significant positive impact on the long-term cash flow from the mine.

6.Critical evaluation of mining dilution. M3 believes there are opportunities to incrementally increase mining grades without additional costs by reducing the dilution included in the mine plan. In addition, future resource modeling efforts should incorporate methods to better depict grade domain boundaries in the resulting model blocks to allow for a more accurate estimate of mining dilution.

7.Tahoe performs annual geotechnical and water management performance reviews of the tailings dry stack to ensure stability and reclamation success.

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2INTRODUCTION

2.1PURPOSE AND BASIS OF REPORT

M3 Engineering & Technology Corporation (“M3”) of Tucson, Arizona in the USA was commissioned by Tahoe Resources Inc. (“Tahoe”) to provide an independent Qualified Person’s Review and Feasibility-Level Technical Report (the “Report”) for the Escobal mine in Guatemala. This review is warranted by the ongoing operation of the mine and the classification of reserves since the previous Preliminary Economic Assessment (PEA) in May 2012.

Tahoe is the sole proprietor of the Escobal mine through its subsidiary, Minera San Rafael, S.A. (“MSR”). The mine is located on the Escobal Exploitation concession granted on April 3, 2013.

This Report uses metric measurements, except where noted. The currency used in the Report is U.S. dollars. The local currency of Guatemala is the Quetzal. At the Report effective date, the exchange rate was US$1 equals 7.65 Quetzals.

2.2SOURCES OF INFORMATION

Tahoe previously filed Technical Reports on the Escobal project, including the following:

·Escobal Project Guatemala NI 43-101 Technical Report, dated 30 April 2010. This report was prepared by AMEC Americas Limited of Vancouver, Canada under the guidance of Mr. Greg Kulla, P. Geo., a Qualified Person (QP) as defined by NI 43-101.

·Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment, dated 29 November 2010. This report was prepared by M3 under the guidance of Mr. Conrad Huss, P.E., a QP as defined by NI 43-101. The November 2010 PEA reported an increase in the mineral resources of the Project and provided technical and economic analyses of the potential viability of those mineral resources.

·Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment, dated effective 7 May 2012 and amended on 24 July 2013. This report was prepared by M3 under the guidance of Mr. Conrad Huss, P.E., a QP as defined by NI 43-101.

Additional information was obtained by M3 or provided by Tahoe, and is contained herein.

2.3QUALIFIED PERSONS AND SITE VISITS

The Qualified Person and Principal author for this report is Conrad Huss, P.E., of M3 Engineering & Technology Corporation. All M3 personnel for this project are supervised by Conrad Huss. Mr. Huss visited the Project site for one day on 1 December 2010.

The Qualified Person responsible for the review of the civil and environmental controls for the Escobal project is Daniel Roth, P.E., of M3 Engineering & Technology Corporation. Mr. Roth visited the Escobal project site numerous times from 2010 to 2014.

The Qualified Person responsible for the review of the metallurgical testing and flow sheets for the Escobal project is Thomas L. Drielick, PE, of M3 Engineering & Technology Corporation. Mr. Drielick has not visited the project site.

The Qualified Person responsible for the review of the geology, exploration, drilling, sampling method, sample preparation and analysis, data verification, and resource estimate for the Escobal project is Paul Tietz, CPG, of Mine Development Associates, an independent mining consulting firm. Mr. Tietz visited the Escobal project site in 2010, 2012, and most recently in January 2014.

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The Qualified Person responsible for mineral reserve estimates, mining studies and mine operating and capital cost estimates for the Escobal mine is Matthew Blattman, PE, of Blattman Brothers Consulting LLC. Mr. Blattman has visited the project site numerous times from 2012 to 2014.

The Qualified Person responsible for dry stack tailings design for the Escobal mine is Jack Caldwell, PE, of Robertson Geoconsultants Inc. Mr. Caldwell has visited the project site numerous times from 2011 to 2013.

Table 2-1 shows the QP names, certifications, affiliations, and section responsibilities. These responsibilities are repeated in the QP certificates.

Table 2-1: List of Qualified Persons

Author	Company	Designation	Site Visit	Section Responsibility
Conrad E. Huss	M3	P.E.	1 December 2010	Section 1 and 22.
Thomas L. Drielick	M3	P.E.	N/A	Sections 13, 17, and corresponding items in Sections 1, 25 and 26.
Daniel Roth	M3	P.E.	Multiple Site Visits 2010-2014	Sections 2, 3, 4, 5, 6, 15, 16, 18, 19, 20, 21, 23, 24, 25, and 26, and corresponding items in Section 1.
Paul Tietz	MDA	C.P.G.	7-Sep to 10-Sep 2010 6-Feb to 9-Feb 2012 28-Jan to 31-Jan 2014	Sections 7, 8, 9, 10, 11, 12, 14, and corresponding items in Sections 1, 24, 25 and 26.
Matthew Blattman	Blattman	P.E. SME Reg. Member	Multiple Site Visits 2012-2014	Sections 15, 16, and corresponding items in Sections 1, 21, 22, 24, 25 and 26.
Jack Caldwell	Robertson	P.E.	Multiple Site Visits 2011-2013	Section 18.6.

2.4EFFECTIVE DATES

The effective date of the Escobal Mineral Resource Estimate is January 23, 2014. The effective date of the Escobal Mineral Reserve Estimate is July 1, 2014. The effective date of the Study is November 5, 2014.

2.5UNITS AND ABBREVIATIONS

The report considers US Dollars ($) only. Unless otherwise noted, all units are metric. However, as noted and as standard for projects of this nature, certain statistics are reported as avoirdupois or English units, grades are described in terms of percent (%), grams per metric tonne (g/t) or troy ounces per short ton (oz/t). Salable base metals are described in terms of metric tonnes and English pounds. Salable precious metals are described in terms of troy ounces.

Table 2-2 shows the abbreviations used in this report.

Table 2-2: Terms and Abbreviations

Abbreviation	Unit or Term	Abbreviation	Unit or Term
% (grade)	Percent by weight (grade)	4WD	Four-Wheel Drive
2-D	Two-Dimensional	AA	Atomic Adsorption
3-D	Three-Dimensional	AAS	Atomic Absorption Spectrometry

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Abbreviation	Unit or Term	Abbreviation	Unit or Term
ABA	Acid Base Accounting	k	thousands
AG	Autogenous Grinding	kg	kilograms
Ag	Silver	kg/t	kilograms per metric tonne
AgEq	Silver Equivalent	km	kilometer
AGP	Acid Generation Potential	Km2	square kilometer
ANP	Acid Neutralization Potential	kPa	kilopascal
ARD	Acid Rock Drainage	kV	kilovolt
AT	Assay Ton	kW	kilowatt
Au	Gold	kW-h	Kilowatt-hour
Blattman	Blattman Brothers Consulting LLC	L	Liters
cfm	Cubic feet per minute	lb	pound
Chemex	ALS Chemex	LOM	Life of Mine
CO3	Carbonate	m	meter
COG	Cutoff grade	m3	cubic meter
Company	Tahoe Resources Inc.	m3/h	cubic meter per hour
Cu	Copper	Ma	Million years old
CV	Coefficient of Variation (standard deviation/mean)	MARN	Ministerio de Ambiente y Recursos Naturales (Ministry of Environment and Natural Resources)
DDH	Diamond Drill Hole
dmt/h	dry metric tonnes per hour	masl	meters above sea level
EIA	Environmental Impact Assessment	MDA	Mine Development Associates
EIS	Environmental Impact Study	MEM	Minesterio de Energía y Minas (Ministry of Energy and Mines)
EPCM	Engineering, Procurement and Construction Management
		Mn	Manganese
FA	Fire Assay	MSR	Minera San Rafael, S.A.
Fe	Iron	MSR	Minera San Rafael, S.A., the Guatemalan operating company of Tahoe
g	gram
g Ag/t	grams of silver per metric tonne	Mt	Megatonnes, or one thousand metric tonnes
g AgEq/t	grams of silver equivalent per metric tonne	MTPD	Metric Tonnes per Day
g Au/t	grams of gold per metric tonne	MW	megawatt
g/cm3	grams per cubic centimeter	MWMP	Meteoric Water Mobility Procedure
g/t	grams per metric tonne	MY	Million years old
g/t Ag	grams of silver per metric tonne	NGO	non-governmental organizations
g/t Au	grams of gold per metric tonne	NNP	Net Neutralization Potential
GPS	Global Positioning System	NPV	Net Present Value
ha	hectare	NSR	Net Smelter Return
HC	Humidity Cell	opt	Troy ounces per English ton
HP / hp	Horsepower	oz/t	troy ounce per short ton
ICP	Inductively-Coupled Plasma	PA	Preliminary Assessment or Preliminary Economic Assessment
ICP	induced-coupled polarization
ID3	Inverse Distance Cubed	PAX	Potassium Amyl Xanthate
INAB	Instituto Nacional de Bosques (National Forestry Institute)	Pb	Lead
		PEA	Preliminary Economic Assessment
Inspectorate	Inspectorate, a division of Bureau Veritas; formerly BSI Inspectorate	ppb	part per billion
		ppm	Part per million
IRR	Internal Rate of Return	PSD	Particle Size Distribution
Ja	joint alteration	QA/QC	Quality Assurance/Quality Control
Jn	joint number	RC	Reverse Circulation
Jr	joint roughness	RMR	rock mass rating
Jw	joint water reduction factor	rpm	revolutions per minute

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Abbreviation	Unit or Term	Abbreviation	Unit or Term
RQD	rock quality designation	tpy	Tonnes per year
S	Sulphur	US$/ USD	United States Dollars
Sb	Antimony	UTM	Universal Transverse Mercator coordinate system
SRF	stress reduction factor
t/d	metric tonnes per day	VFD	Variable Frequency Drive
t/h	metric tonnes per hour	WCN	WCM Minerals
Tahoe	Tahoe Resources Inc.	XRD	X-Ray Diffraction
tonne	metric tonne	XRF	X-Ray Fluorescence
tonnes	dry metric tonnes (where one tonne = 1.1023 short tons)	Zn	Zinc
		μm	micrometer or micron
tpa	Tonnes per annum

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3RELIANCE ON OTHER EXPERTS

The QP as author of this Report states that he is a qualified person for the Report as identified in the “Certificate of Qualified Person” attached to the Report. The author has relied upon information derived from the following expert reports pertaining to mineral rights, surface rights, and permitting issues.

In cases where the M3 PEA author, Conrad Huss, P.E., Qualified Person, has relied on contributions of the Qualified Persons listed in Appendix A, the conclusions and recommendations are exclusively the Qualified Persons’ own. The results and opinions outlined in this report that are dependent on information provided by Qualified Persons outside the employ of M3 are assumed to be current, accurate and complete as of the date of this report. See Section 2.3 for a tabulation of QP responsibilities.

Reports received from other experts have been reviewed for factual errors by Tahoe and M3. Any changes made as a result of these reviews did not involve any alteration to the conclusions made. Hence, the statements and opinions expressed in these documents are given in good faith and in the belief that such statements and opinions are not false and misleading at the date of these reports.

Metallurgical testing done by Tahoe’s consultants depends on the samples’ accuracy representing the Escobal deposit.

The base case metal prices utilized herein were provided by M3.

3.1MINERAL TENURE

M3 has not examined mineral tenure, nor independently verified the legal status or ownership of the Project area or underlying property agreements. M3 has fully relied upon independent legal experts for this information through the following documents:

·Arenales & Skinner-Klee, 2010: unpublished legal opinion letter prepared by Arenales & Skinner-Klee for Entre Mares, S.A., 23 February, 2010.

·Arenales & Skinner-Klee, 2010: unpublished legal opinion letter prepared by Arenales & Skinner-Klee for Entre Mares, S.A., 21 May, 2010.

Data and information is derived from work done by previous owners of Escobal and more recent work by Tahoe Resources Inc.

3.2SURFACE RIGHTS, ACCESS, AND PERMITTING

M3 has fully relied on information regarding the status of the current Surface Rights, Road Access and Permits through opinions and data supplied by independent legal experts through the following documents:

·Arenales & Skinner-Klee, 2010: unpublished legal opinion letter prepared by Arenales & Skinner-Klee for Entre Mares, S.A., 23 February, 2010.

·Arenales & Skinner-Klee, 2010: unpublished legal opinion letter prepared by Arenales & Skinner-Klee for Entre Mares, S.A., 21 May, 2010.

Data and information for the Escobal Exploitation License was provided by Tahoe Resources Inc.

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4PROPERTY DESCRIPTION AND LOCATION

4.1LOCATION

The Escobal mine is located in southeast Guatemala, approximately 40 kilometers east-southeast of Guatemala City and two kilometers east of the town of San Rafael Las Flores in the Department of Santa Rosa (Figure 4-1). The mine is centered at UTM coordinates 806,500E 1,601,500N (NAD27, Zone 15).

Figure 4-1: Escobal mine Location Map

The Escobal mine is situated entirely within the Escobal Exploitation Concession, which is 2,000 ha in size. Within the exploitation concession, the Escobal mine property encompasses approximately 280 ha.

4.2MINERAL TENURE AND TITLE

There are three types of mining licenses (concessions) recognized in Guatemalan law — reconnaissance, exploration, and exploitation. According to Guatemala law, reconnaissance licenses are granted for a six-month period. Exploration concessions are granted for an initial period of three years which can be extended for two additional periods for two years each, for a total holding period of seven years. Following the seven year exploration period, no additional extensions are permitted to the exploration license and application must be made an exploitation license or new exploration concession.

In Guatemala all concessions are “coordinate staked” as referenced by UTM coordinates of the concession corners. No physical survey of concession boundaries is required and no monuments are located on the ground. Individual exploration concessions can cover up to 100 km2; exploitation concessions can cover up to 20 km2.

Through its wholly owned subsidiary, Minera San Rafael, S.A. (MSR), the Company holds mineral rights to 150.7 square kilometers in Guatemala through exploration and exploitation concessions. MSR’s concession package

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includes the Escobal exploitation concession (20.0 km2) which covers the Escobal vein and mine operation facilities, and the Andres, Lucero, and Juan Bosco exploration concessions which cover an additional 130.7 km2 in the region. The Escobal exploitation concession and the Andres, Lucero, and Juan Bosco exploration concessions have all been approved by the Ministry of Energy and Mines (Ministerio de Energía y Minas, MEM) and are in good standing with environmental management plans in place and approved by the Ministry of Environment and Natural Resources (Ministerio de Ambiente y Recursos Naturales, MARN). Applications for a number of additional exploration and reconnaissance concessions have been submitted to MEM and are pending approval.

All approved and pending concessions are shown in Figure 4-2; descriptions of each concession are summarized in Table 4-1.

Figure 4-2: Minera San Rafael Concessions — Active and Pending

No exploration or reconnaissance license applications have been approved by MEM during the past year and additional concession approval is unlikely during the remainder of the current government term (2016).

Table 4-1: San Rafael Concessions

Concession			Application	Approval	1st Ext.
License No.	Type	Area (Km2)	date	date	approved	Remarks
OASIS LEXR-040-06	Explor.	40.0	10/25/2006	3/15/2007	4/27/2010	Converted to Escobal Exploit concession 2013
LUCERO LEXR-041-06	Explor.	30.8	10/25/2006	7/20/2007	3/29/2012	2nd Ext applied 6/18/2012
SOLEDAD SR-03-06	Recon.	802.5	12/6/2006	NA
ANDRES LEXR-030-07	Explor.	40.0	5/18/2007	11/6/2007	3/29/2012	2nd Ext applied 12/12/2012
EL OLIVO SEXR-029-07	Explor.	36.0	5/18/2007	NA

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Concession			Application	Approval	1st Ext.
License No.	Type	Area (Km2)	date	date	approved	Remarks
JUAN BOSCO SEXR-089-08	Explor.	59.9	11/12/2008	5/9/2012
PUENTE QUEBRADO SEXR-049-09	Explor.	3.0	10/9/2009	NA
MELISSA SEXR-050-09	Explor.	3.0	10/9/2009	NA
VALENCIA SEXR-050-10	Explor.	7.0	8/23/2010	NA
GRANADA SEXR-054-10	Explor.	5.0	10/6/2010	NA
CRISTINA SEXR-055-10	Explor.	52.5	10/6/2010	NA
EL SILENCIO SR-06-10	Recon	1,098.1	11/4/2010	NA
CIPRESES SEXR-048-09	Explor.	3.0	10/09/2009	NA
PAJAL SEXR-058-11	Explor.	66.0	5/4/2011	NA
ESCOBAL LEXT-015-11	Exploit	20.0	7/8/2011	04/03/2013
EL DURAZNO SEXR-104-11	Explor.	48.8	7/29/2011	NA
PAJARITA SEXR-104-11	Explor.	57.0	7/29/2011	NA
TERESA SEXR-109-11	Explor.	68.5	8/17/2011	NA
OASIS I SEXR-117-11	Explor.	12.8	8/31/2011	NA
OASIS II SEXR-118-11	Explor.	7.0	8/31/2011	NA
OASIS III SEXR-119-11	Explor.	0.2	8/31/2011	NA

Yearly payments are made to MEM for each concession based on concession size and a graduating “concession age” factor. For exploration concessions the current holding cost amounts to a Q3,000 to Q9,000 (~US$380 to US$1,135) concession holding fee per square kilometer. For exploitation concessions a fixed cost of Q12,015 (~US$1,500) is charged per square kilometer. Required payments are current for all of the Company’s currently held concessions.

There are no defined work requirements to keep an exploration concession valid, although exploration activity (sampling, mapping, etc.) must to be conducted and results filed with the MEM on an annual basis. The Company has filed exploration activity reports with MEM for all exploration and exploitation concessions each year as required.

4.3SURFACE RIGHTS

In Guatemala, surface rights are independent of mining rights and must be negotiated separately. There is no allowance for expropriation in Guatemala. The Company has purchased approximately 281 ha of surface rights for the area required for mining, processing plant and ancillary facilities, surface operations, and tailings and waste rock disposal, as shown in Figure 4-3. Annual property taxes to the San Rafael municipality are approximately US$150,000 per year. No additional surface area is required for the mine expansion to 4,500 t/d.

In areas peripheral to the project where surface rights have not been purchased, annual rental fee agreements are in place with a number of land owners to provide for access and site preparation to accommodate exploration activities and drilling. No liabilities currently exist for land usage.

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Figure 4-3: Surface Ownership Map

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4.4PROPERTY AGREEMENTS

The Company established a profit sharing program that provides a 0.5% net smelter return (NSR) to an association of the former land owners of the Escobal mine property. Ten percent of this money is deposited in a special fund, administrated by the association’s board of directors, to be used for improvements in local communities.

4.5PERMITS

The Escobal mine operations are conducted under an Environmental Impact Assessment (EIA) and an Exploitation License.

The EIA was submitted to MARN in August 2011 and approved by the ministry in October 2011 (Resolution 3061-2011). The EIA required documentation of baseline conditions, a project description, and an analysis of potential impacts and their mitigation measures. Public disclosure and involvement was been required and developed throughout each stage of the permitting process. The approval of the EIA allowed the company to begin full construction of the mine, process plant and all other surface facilities. The Company files quarterly environmental monitoring reports with MARN as required by the resolution.

Mineral production is licensed through MEM; as such, receipt of this license is required before production could commence. Application for the Exploitation License was submitted to MEM in November 2010 and approved in April 2013 (License no. LEXT-015-11). The Company files annual reports with MEM as required by the license stipulations.

Environmental requirements for surface exploration activities are specified in Resolution 4590-2008/ELER/CG, which was issued by MARN to Entre Mares in December 2008. These requirements and approvals were transferred from Entre Mares to MSR in September 2010 as specified in MARN Resolution 1918-2010/ECM/GB.

The Company obtained permits from the National Institute of Forests (INAB) for the specific areas where tree cutting was required for surface facility construction. Land use changes in the mine area have also received INAB approval as required.

4.6ENVIRONMENTAL LIABILITIES AND MANAGEMENT

There are no known environmental liabilities to which the Escobal mine property is subject, other than the obligation by the Company to operate and reclaim the Escobal mine in compliance with the applicable MARN and MEM requirements.

The Company believes that the Escobal mine warrants a high level of environmental stewardship and operates with a mandate to meet or exceed standards of sustainability and environmental management consistent with North American practice and regulation. This section summarizes the elements of design and practice relating to environmental management and stewardship that were considered during construction of the mine.

No impacted waters and materials will be directly discharged from the site. Impacted water is routed to lined containment, sampled, and treated (if necessary) prior to being released to the environment. The environmental management program includes:

·Primary Watershed Considerations

·Dry Stack Tailings

·Lined stormwater and waste facilities

·Concurrent Reclamation

·Process water recovery and recycling

·Process/Contact Water Treatment Facility

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·Underground Paste Backfill

·Geochemical Characterization

·Environmental Management Program

4.6.1Primary Watershed

The plant site facilities were designed and located such that the upstream primary natural watershed was not significantly diverted. Only the portion of the drainage near the operational facilities was realigned into a concrete-lined channel. The drainage returns to its original channel prior to exiting the mine site. The avoidance of diverting this major watershed reduced the overall area of disturbance and maintained the historic flow of water through the property.

4.6.2Dry Stack Tailings

Dry stack tailing management provides significant environmental and operational advantages over traditional wet or slurry tailings disposal and impoundment methods. The primary benefit derived from the dry stack tailing system is water balance. As the tails are filtered to ~15% moisture, the remaining water is returned to the process stream providing a direct offset to make-up water normally obtained from ground water pumping. Another benefit of the dry stack tailing system is the reduced footprint compared to a wet system.

4.6.3Lined Stormwater and Waste Facilities

All facilities located on permeable ground that contains or receives impacted waters is lined for containment.

4.6.4Concurrent Reclamation

Concurrent reclamation of the tailings and waste rock storage facility outslopes allows for early reclamation with either native seed mix or a return to agricultural crops. Natural landform grading is incorporated to provide a more stable, sustainable and natural functioning final surface.

4.6.5Process Water Recovery and Recycling

The process design in both the tails and concentrate thickening/filtration circuits maximizes process water recovery and reuse. In addition, contact water from the dry stack tailings facility as well as contact water from haul roads and active mill and plant areas is collected in channels and lined stormwater ponds for reuse in the process stream. Recycling and utilizing this water for operational uses reduces the need for make-up process water from fresh water sources and minimizes the potential for aquifer impacts in the region.

4.6.6Paste Backfill

Approximately 50% of the tailings produced is mixed with cement and water in a batch plant and disposed of underground as paste backfill, providing several environmental advantages.

·Provides stability to the underground workings, increasing safety and reducing the possibility of subsidence expressions reaching the surface.

·Provides an opportunity to encapsulate any potentially acid generating development materials, isolating them from water and oxygen thus preventing any potential metals leaching or acid generation.

·Provides reduction of storage area required on the surface, about 50% of the tails produced disposed of above ground.

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4.6.7Geochemical Characterization

Geochemical Characterization of the waste rock and tailings demonstrate a large net neutralizing capacity, with no deleterious metals in the waste rock or tailings effluent exceeding regulatory limits. Waste rock and tailings samples are collected systematically and regularly with consistently favorable results indicating almost no potential for acid generation. Sampling and characterization of the waste rock and tailings will continue throughout the life of the mine.

4.6.8Environmental Management Program

Potential impacts from mining and processing operations are monitored and managed under a comprehensive Environmental Impact Management Program developed specifically for the Escobal mine. Based on North American standards, the program functions to avoid, minimize, mitigate and remediate, in that order, all potential impacts. The management plan is designed to comply with the requirements of the MARN and MEM resolutions approving the Escobal mine.

4.7RISKS TO ACCESS, TITLE OR OPERATIONS

No significant or specific factors or risks have been identified that directly affect access, title, or the ability of the Company to operate the Escobal mine other than those associated with operating in a developing country.

Some communities and non-governmental organizations (NGO) have been vocal and active with respect to mining and exploration activities in Guatemala. These communities and NGOs have taken such actions as road closures, opposition to infrastructure construction, work stoppages, and law suits for damages.

4.8RECLAMATION

The Company has an obligation to reclaim its properties. The Company recognizes the present value of liabilities for reclamation and closure costs in the period in which they are incurred. A corresponding increase in the carrying amount of the related assets is recorded and amortized over the life of the asset. As at June 30, 2014, the Company has estimated the present value of the future reclamation obligation arising from its activities to be approximately $6 million.

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5ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1ACCESSIBILITY

The Escobal mine is easily accessible via 70 km of paved four- and two-lane highways from Guatemala City to the town of San Rafael Las Flores. The main entrance to the mine property is situated along the highway just outside of San Rafael Las Flores. The center of the mine facilities is approximately 1.5 km on an improved gravel road from the main entrance. The mine is accessible all year. Access to the upper elevations of the mine property (outside of the industrial area) is generally limited to four-wheel drive vehicles on less developed roads and trails.

5.2CLIMATE

The local climate consists of two major seasons; a “rainy” season between May and November and a “dry” season between November and May.

Average annual precipitation amounts to 1,689 mm with June and September the rainiest months with 315 mm and 335 mm, respectively. January and December are the driest months with 0.6 mm and 10.6 mm of rain, respectively.

Average temperatures vary between April, the hottest month with average lows of 19°C and highs of 33.1°C and January, the coldest month, with average lows 14°C and highs of 30°C. Climate measurements are from a combination of sources including the Company’s weather station at the mine site, and weather stations in Los Esclavos, Cuilapa, and Santa Rosa located 30 km southwest of the mine area.

Mining and exploration activities are carried out year-round without interruption due to weather.

5.3LOCAL RESOURCES AND INFRASTRUCTURE

The town of San Rafael Las Flores has a population of approximately 3,500 people. According to Guatemala’s National Institute of Statistics (Census 2002), San Rafael Las Flores’ population is 99.6% “Ladino”, i.e., of Hispanic origin and non-indigenous.

San Rafael Las Flores has basic services such as schools, banks, hotels, restaurants, small shops, and a health center. Mataquescuintla (population of approximately 8,000), located approximately 7 km north from San Rafael and Nueva Santa Rosa (population of approximately 15,000), located approximately 20 km south from San Rafael are more developed with diverse banking, commerce and health services. Housing and food services for many of the Company’s employees are located in the town of San Rafael.

Although there was some historic mining in the area, there was no local workforce experienced in modern mining and the Company instituted appropriate training programs for the local workforce. Several smaller villages surround the mine property and contribute to the mine labor pool.

Most general supplies required for the Escobal mine operations are sourced in Guatemala, but major mining-specific equipment and supplies are not available in-country and have been imported.

5.4EXISTING INFRASTRUCTURE

There is a 13.2 kV medium voltage line to the town of San Rafael Las Flores; however, this line is not capable of handling the load requirements for the project. The mine’s power requirements are met by on-site power generation, capable of 15.5 MW, provided by contractor supplied diesel generators. The mine has 5.5 MW of additional power generation capacity. Several power options are currently being assessed as improvements to the mine economics can be realized with lower-cost alternative power sources.

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Internet service and traditional and cellular telephone service are available at the mine site and in San Rafael Las Flores. In addition, a fiber optic communication line is also available in San Rafael and at the mine facilities.

Process and other industrial use water is supplied by mine dewatering wells and from water pumped from the underground workings. Domestic water is supplied by dedicated water wells located on the mine property.

The Company owns sufficient land to conduct its operations at the Escobal mine, including areas for mining, processing plant, tailings dry stack, and all other facilities. No additional land purchases are required for the mine expansion discussed in this report.

5.5PHYSIOGRAPHY

The Escobal mine lies within mountainous terrain interspersed with rolling hills and valleys. Elevations range from 1,300 masl in the valley to the west of the mine to 1,800 masl to the east. The high mountain range of Montaña Soledad Grande north and east of the mine rises to an elevation of 2,600 masl.

Vegetation is characterized by natural mountain forest species that consist of oak, pine and cypress tree varieties and lower strata scrub-brush species.

Agricultural products in the area include corn and beans for local consumption, and commercial production of onions, tomato and coffee.

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6HISTORY

6.1INTRODUCTION

Guatemala does not have a well-recognized mining history, though an area near Mataquescuintla, approximately 7 km north of the Escobal Project, is the site of copper-silver production from underground mining around the turn of the 20th century.

A small underground operation was developed on an antimony showing at the Loma Pache prospect 600 m north of the Escobal vein in the 1970s. There are no drilling records available from the development of the Loma Pache prospect. Production records for both operations are incomplete; the Mataquescuintla (a.k.a. Colis) mine reportedly produced between 8,000 and 10,000 tons of concentrate of unknown grade. Underground grades are reported to be 217 g/t silver (Ag), 0.2 g/t gold (Au), 1.27% copper (Cu) and 24% sulphur (S).

6.2GOLDCORP / ENTRE MARES 1996-2010

Interest in the Escobal area dates back to 1996 when Entre Mares de Guatemala S.A., the Guatemalan subsidiary of Goldcorp, prospected in the area and identified high-grade gold values associated with surface quartz veins in the western portion of the Escobal vein zone. Size potential of the zone was deemed uneconomic at the time and exploration activities were discontinued later that year. In 2006, Entre Mares reinitiated regional exploration in the area, partially based on verifying geochemical anomalies in the company database. In late 2006, significant silver and gold grades were detected from surface sampling along an extensive alteration zone developed over the Escobal vein. An exploration concession was applied for in October 2006 and was granted in March 2007. Entre Mares commenced exploration drilling May 2007; Tahoe has continued exploration drilling on the property which is ongoing as of the effective date of this Report.

In early 2010 Goldcorp, predecessor to Tahoe in ownership of Escobal, reported a Measured and Indicated mineral resource estimate for Escobal of 6.97 Mt at 0.63 g/t Au and 580.3 g/t Ag and an Inferred mineral resource of 13.15 Mt at 0.53 g/t Au and 443.4 g/t Ag (February 17, 2010 Goldcorp news release). Goldcorp did not release a technical report to support the mineral resource declaration at that time.

6.3TAHOE RESOURCES

In May 2010, Tahoe executed an agreement to acquire the Escobal Project from Goldcorp’s indirectly wholly owned subsidiaries, Goldcorp Holdings Barbados Ltd. and Guatemala Holdings Ltd., which respectively held 9.1% and 91.9% of Entre Mares de Guatemala S.A. On June 8, 2010 upon successful completion of Tahoe’s Initial Public Offering (IPO), Tahoe acquired all of the common shares of Entre Mares including the Escobal project and the exploration concessions discussed in this Report.

In preparation of Tahoe’s acquisition of the Escobal project, AMEC Americas Ltd. authored an independent Technical Report for Tahoe in April 2010 and reported an NI 43-101 compliant resource estimate for the Escobal deposit based on 46,333 m of drilling in 175 surface holes. AMEC reported an Indicated Mineral Resource of approximately 100 million ounces of silver contained in 4,570,000 tonnes at an average silver grade of 684 g/t and an Inferred Mineral Resource of approximately 176 million ounces of silver contained in 12,800,000 tonnes at an average silver grade of 427 g/t (AMEC, 2010).

Tahoe engaged M3 Engineering & Technology Corporation (M3) in 2010 to prepare the Escobal Preliminary Economic Assessment (2010 PEA), dated November 29, 2010 that contained an updated NI 43-101 compliant mineral resource estimate based on data from 61,469 meters in 220 surface diamond drill holes. The Preliminary Assessment reported an Indicated Mineral Resource of 245.2 million ounces of silver contained in 15.3 million tonnes at an average silver grade of 500 g/t and an Inferred Mineral Resource of 71.7 million ounces of silver contained in 8.3 million tonnes at an

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average silver grade of 271 g/t. In addition, both mineral resource categories included significant amounts of gold, lead, and zinc. The PEA indicated that a 3,500 t/d underground mine producing lead and zinc concentrates over a production life of 18 years was economically viable. The metallurgical studies that had been completed to date confirmed that processing through differential flotation would produce marketable lead and zinc concentrates. The PEA showed an after tax net present value at a five percent discount rate of $1.729 billion at the base case metal prices of $18.00/oz silver, $1,100/oz gold, $0.95/lb lead and $.090/lb zinc and an after tax internal rate of return (IRR) of 51% on an initial capital cost of $326.6 million (M3 Engineering and Technology Corporation, 2010).

Indicated and Inferred mineral resources were the basis for the 2010 PEA. By definition, the 2010 PEA was preliminary in nature, in that it included Inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves.

Based on the results of the 2010 PEA, Tahoe’s Board of Directors approved management’s recommendation to proceed to mine development and production. Tahoe contracted M3 for the engineering, procurement and construction management (EPCM) activities to develop the Escobal mine.

In February 2011, the Guatemalan Ministry of Environment and Natural Resources (Ministerio de Ambiente y Recurcos Naturales, MARN) approved the Company’s initial Environmental Impact Assessment (EIA) which allowed for the initiation of underground development, construction of the access road from the main highway and establishment of temporary surface facilities to support the underground development. In October 2011, MARN approved the Company’s EIA for the design, construction, operation and closure of the Escobal mine at which time M3 began mobilizing contractors to commence earthwork and construction of the process plant and associated support facilities in support of a 3,500 t/d mining and processing operation at Escobal. Construction of the surface facilities continued through 2013.

Tahoe’s continuing exploration successes at Escobal prompted the Company to contract M3 to conduct a second Preliminary Economic Assessment (2012 PEA) analyzing increased mine and plant throughput associated with the extraction of additional resources. The 2012 PEA included an updated NI 43-101 compliant Mineral Resource estimate and reported Indicated mineral resources of 27.1 million tonnes with average grades of 422 g/t silver, 0.43 g/t gold, 0.71% lead and 1.28% zinc. Indicated silver ounces totaled 367.5 million. Inferred mineral resources for the Escobal deposit were reported at 4.6 million tonnes with average grades of 254 g/t silver, 0.59 g/t gold, 0.34% lead and 0.66% zinc. Inferred silver ounces totaled 36.7 million. The 2012 PEA indicated that throughput increases from 3,500 t/d to 4,500 and/or 5,500 t/d would improve the economics of the project. The 2012 PEA showed an after tax net present value for the 4,500 t/d case at a five percent discount rate of $2.94 billion at the base case metal prices of $25.00/oz silver, $1,300/oz gold, $0.95/lb lead and $.090/lb zinc and an after tax IRR of 68.3% on an initial capital cost of $372.8 million. After tax net present value for the 5,500 t/d case at a five percent discount rate was $2.99 billion at the same base case metal prices with an after tax IRR of 68.5% on initial capital costs of $405.4 million (M3 Engineering and Technology Corporation, 2012).

As with the 2010 PEA, Indicated and Inferred mineral resources were the basis for the 2012 PEA. By definition, the 2012 PEA was preliminary in nature, in that it included Inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves.

Based on the results of the 2012 PEA, the crushing and grinding circuits, flotation circuit, and concentrate thickening and filtering processes were designed to accommodate an increase throughput without major equipment replacement required. Likewise, the flotation, concentrate and tailings filtration buildings were constructed to accommodate additional equipment necessary for increased mill throughput.

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The 2012 PEA was subsequently revised and reissued in July 2013 with no substantive changes to the results and conclusions of the 2012 PEA.

The Escobal processing plant was commissioned in September 2013, producing its first concentrates on September 29, 2013. The plant reached the 3,500 t/d design throughput rate in June 2014. Through June 30, 2014, the Escobal mine had produced 12 million ounces of silver, 7,600 ounces of gold, 6,600 tonnes of lead and 7,300 tonnes of zinc contained in lead and zinc concentrates from mill feed averaging 581 g/t silver, 0.47 g/t gold, 1.0% lead and 1.4% zinc. Underground development through June 30, 2014 totaled approximately 17,100 m with an additional 229 m of vertical ventilation raise development completed.

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7GEOLOGICAL SETTING AND MINERALIZATION

The Escobal deposit is an intermediate-sulfidation fault-related vein system formed within Tertiary sedimentary and volcanic rocks within the Caribbean tectonic plate. The Escobal vein system hosts silver, gold, lead and zinc, with an associated epithermal suite of elements, within quartz and quartz-carbonate veins. Quartz veins and stockwork up to 50 m wide, with up to 10% sulfides, form at the core of the Escobal deposit and grade outward through silicification, quartz-sericite, argillic and propylitic alteration zones.

Drilling to date has identified continuous precious and base metal mineralization at Escobal over 2,400 m laterally and 1,200 m vertically in four zones; the East and East Extension zones (collectively, the East Zone), the Central Zone, and the West/Margarito zones (collectively, the West Zone). The vein system is oriented generally east-west, with variable dips. The East and East Extension zones dip from 60° to 75° to the south with recent drilling showing a change to vertical dip at depth. The majority of the mineralized structure(s) in the Central and West/Margarito zones dip from 60° to 75° to the north, steepening to near-vertical at depth. The upper eastern portion of the Central Zone dips 60° to 70° to the south, as in the East Zone.

7.1REGIONAL GEOLOGY

Guatemala comprises two geologic terrains formed as the result of the convergence of a major tectonic plate boundary. The North American plate comprises the northern half of Guatemala, and the Caribbean plate comprises the southern half, with three major east-west trending, left-lateral transform faults forming the plate collision boundary. From north to south this boundary is defined by the Polochic, Motagua and Jocotan fault systems (Figure 7-1). The Escobal deposit lies within the Caribbean plate, south of the Jocotan fault. The northern side of the Jocotan fault system contains Paleozoic metasediments, schist and gneiss, while the south side contains a series of Tertiary mafic volcanic eruptive events composed mostly of dacitic to andesitic tuff, lahar and andesitic to basaltic flows. These eruptive units are separated by thin beds of water-lain sediments consisting mostly of fine to medium grained clastic and tuffaceous sediments. Tertiary volcanics are commonly covered by Quaternary dacitic volcanic eruptive ash units. The Escobal deposit is within the Tertiary mafic eruptive units that trend subparallel to the Jocotan fault system.

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Figure 7-1: Regional Geology

7.2LOCAL AND PROPERTY GEOLOGY

Project area surface geology is shown in Figure 7-2. The area is underlain by the Eocene Subinal Formation, a series of interbedded volcaniclastic sediments that include siltstone, fine-and coarse-grained sandstone, tuff, and limestone-clast conglomerate. This formation is unconformably overlain by a package of medium-grained, massive porphyritic andesite and lithic tuff composed of fine- to coarse-grained lapilli. Magnetic andesitic dikes, the youngest Tertiary lithological units in the project area, cross-cut all older rock units. A thin unit of Quaternary pyroclastic ash irregularly overlies all lithological units over large portions (approximately 60%) of the project area.

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Figure 7-2: Local Geology

7.3LITHOLOGY

Specific lithological units from oldest to youngest in age and their corresponding map designations are described below:

Volcaniclastic-Clastic Sequence (Subinal Fm) (cr):

·A volcaniclastic sedimentary sequence related to regional redbeds forms the local basement in the Escobal area. These rocks are believed to correlate with the Subinal Formation, a continental clastic sequence that is distributed throughout central and southeast Guatemala. The volcaniclastic sequence at Escobal contains subunits of lapilli, andesite and crystal tuff intercalated with siltstone, sandstone and conglomerate. Individual beds range in thickness from five to 200 m. The unit is exposed as irregularly-distributed windows in drainages and has a minimum thickness of 500 m.

·Sedimentary and volcanic subunits prove to be difficult to use as marker beds, due to their irregular distribution and repetitive occurrence. Recent drilling in the West Zone has identified a specific narrow sub-horizontal andesite unit in the extreme west drill sections. Distribution of this andesite bed shows a marked displacement; 150 to 250 m down to the west, suggesting normal basin margin faulting around the 805,900E section.

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Porphyritic Andesite (ap):

·A sub-horizontal shaped body of porphyritic andesite unconformably overlies the basement volcaniclastic Subinal Formation sediments throughout the Escobal area. The unit is massive to medium-grained and porphyritic, with feldspar, biotite and quartz phenocrysts in a fine-grained matrix.

·This unit is thought to be hypabyssal or intrusive in origin as it is texturally very consistent and shows no mineralogical zonation. The unit forms rare outcrops in the Escobal area and has been defined in drilling over a thickness of 500 m. Based on regional geological relationships, the porphyry is believed to be Upper Miocene in age.

·Recent drilling in the extreme west portion of the project area encountered a thick unit of andesite breccia that is interpreted as a flow within or proximal to a volcanic vent. The monolithic matrix-supported breccia is composed of sub-angular to sub-rounded porphyritic andesite clasts within a similar composition matrix. Due to limited drilling in the area, the size and distribution of this unit is not currently known.

Lithic Tuff (tl):

·A unit of young post-mineral lithic tuff overlies the andesite porphyry in the northeast and far-west portions of the Escobal area. The unit consists of white, non-welded ashflow tuff with angular to sub-rounded, lapilli to pebble-sized lithic fragments of basalt to rhyolite composition. This unit has an observed thickness of 50 to 150 m, thickening to the east where less erosion is evident, and masks the eastern extension of the East Zone of the Escobal vein.

Andesite Dikes (ad):

·Post-mineral andesite dikes cut the three major lithological units and are believed to be of late-Tertiary (post- Miocene) age. Dikes occur in the eastern portion of the Escobal vein where they can be followed for three kilometers along a N40W regional trend. The dikes are near-vertical tabular bodies that range in width from 20 cm to 10 m and primarily occupy the footwall of the East Escobal vein. The dikes are composed of euhedral feldspar crystals immersed in a very fine grained matrix and are generally fresh to weakly altered, containing rare minor quartz veinlets. The dikes are variably magnetic, relative to the degree of alteration/weathering.

Quaternary ash-airfall tuff (Qc/Qph):

·Non-lithified ash and pumice-rich tuff is widespread and covers most ridges and topographic highs in the project area. Thickness is variable, though is commonly several meters thick on hilltops and slopes. Ash is typically eroded from drainages and valleys, though where reworked and transported, can form up to 20-m- thick deposits.

·The ash unit comprises two layers; a basal very coarse-grained, unconsolidated, heterolithic layer; and an upper layer of medium- to fine-grained unconsolidated ash.

7.4STRUCTURE

The dominant structural trend in the region parallels the regional Montagua and Jocotan fault systems along an east-west to N60E trend. At Escobal, this structural trend is represented by a series of east-west trending normal faults that generally exhibit down-to-the south movement, typical of an extensional structural regime. These faults display lithologic displacement, shear and gouge zones, and are host to the high-angle south dipping veins that define the East Zone of the Escobal vein and upper and lower limbs of the variably-dipping Central Zone.

Dilational jogs, or tensional shears, are commonly observed in extensional fault terrains (pull-apart basins or grabens) where the area between individual normal faults exhibit wide zones of disruption as a response to the structural event.

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These zones are commonly wider and more gently dipping than the primary steeply dipping structures. The wider moderate north-dipping Central mineralized zone is interpreted as occupying a dilational jog between the normal faults represented by the East Escobal vein zone and the upper and lower limbs of the Central vein zone (Figure 7-3). These dilational jogs are also interpreted where changes in dip occur in deeper portions of the West/Margarito and East Escobal vein zones.

A N40W trending structure dissects the Escobal vein between the East and Central zones. This feature is evidenced by the occurrence of steep dipping (70°SW) andesite dikes and an apparent (approximately 100 m) left-lateral shift of mineralization. Based on relative elevations of mineralization and lithologic markers, this is a normal fault with vertical movement on the order of 50 m (down to the southwest). Similar northwest to northeast trending faults may interrupt or offset the Escobal vein along its eastern extension.

In the extreme west margin of the Escobal vein, drilling delineated an andesite marker horizon within the volcaniclastic sequence that suggests a fault at the margin of the San Rafael valley. Relative location of the sub-horizontal andesite suggests normal fault displacement on the order of 150 to 250 m (down to the west). This offset is believed to occur along a series of en echelon faults that progressively offsets stratigraphy and veins west into the San Rafael valley.

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Figure 7-3: Interpretation of Central Escobal Vein (North-South Section — Looking East)

7.5MINERALIZATION

Economic mineralization at Escobal comprises silver, gold, lead, and zinc hosted within quartz veins, stockwork zones and hydrothermal breccias. The deposit predominantly comprises sulfide mineralization. Silver, lead, and zinc sulfide mineralization predominates in the Central and West zones though elevated gold values also occur at depth in the Central Zone. In the East Zone, gold-rich mineralization is associated with the upper mixed sulfide-oxide horizon. Silver mineralization in all zones shows a close association with galena and low-iron sphalerite.

A petrographic study of vein samples indicated a fairly simple and consistent paragenesis. Stage I veining consists of banded to massive chalcedony intercalated with quartz and carbonate. This is the volumetrically-dominant vein event and contains the bulk of sulfide minerals. Volumetrically lesser Stage II consists of sulfide-bearing granular chalcedony. Various episodes of post-sulfide quartz, and late barren calcite veining locally cut and/or overprint the main banded vein.

Based on analysis of petrographic characteristics at least five events of quartz veining are interpreted (Figure 7-4). These include, from oldest to youngest:

1)Dominant banded quartz-chalcedony vein.

2)Silica flooding event (quartz-chalcedony)

3)Narrow chalcedony/quartz veinlets.

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4)Narrow euhedral quartz veinlets

5)Late hematite-goethite-chlorite-sericite and calcite replacement veinlets

Gold-silver and base-metal mineralization occurs exclusively in the first two events. Narrow later-stage quartz and chalcedony veinlets are not considered precious or base metal depositing events.

Figure 7-4: Vein Episodes and Generalized Relationships

Silver minerals are dominantly proustite (+/- pyrargyrite), silver sulfide (acanthite), lesser amounts of other silver sulfosalts (stephanite, polybasite and silver-bearing tetrahedrite) and minor native silver. Gold minerals include electrum and native gold. These minerals and other sulfides occur as aggregates of abundant finely disseminated grains most commonly in chalcedony, and as grains interstitial to quartz in select bands, and as more isolated grains, especially with visible gold, throughout the vein in chalcedony/quartz. Aggregates of grains commonly consist of pyrite, acanthite, proustite, visible gold, ± galena, ± sphalerite, and ± chalcopyrite. Acanthite, proustite, and visible gold commonly are found together as disseminated aggregates exhibiting no, or rare, mutual contacts. In places, gold exhibits mutual contacts with acanthite and proustite.

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Figure 7-5: Escobal Central Zone — Breccia with Red Proustite Bands (Drill Hole E08-110)

There is no definitive boundary to the overall vein width or to the down dip extensions of the vein. Silicification and stockwork concentrations increase inwards towards massive banded and/or brecciated quartz ± carbonate material within the vein center. Mineralization may start abruptly or may gradually increase through the stockwork. The vein appears to be better constrained in the volcanic host rocks and more diffuse and unconstrained in the sediment host rocks.

Drilling in the far west end of the Escobal vein encountered a wide zone of massive gypsum veins associated with hornfels and/or contact metamorphic minerals. This is the only gypsum occurrence recognized in the project area. Because it is associated with a similarly rare wide zone of brecciated andesite, it is believed to represent a low-temperature hot springs environment related to the margin of an andesitic volcanic vent.

7.6ALTERATION

Alteration mineralogy is typical of intermediate sulfidation epithermal systems. Quartz veins and stockwork up to 50 m wide, with up to 10% sulfides, form at the core of this alteration pattern and grade outward through silicification, quartz-sericite, argillic and propylitic zones. Recent petrographic studies from samples in the deep West extension of the Escobal vein have identified hornfels and other contact metamorphic effects possibly related to hydrothermal or intrusive events in this area.

The following descriptions provide additional detail on the alteration types:

Silicification

·Pervasive silicification is intimately related to zones of mineralization and forms as halos on both sides of the principal veins. Silica replacement is common in the matrix and occasionally replaces minor accessory minerals. Where strongly silicified, the rock is totally replaced leaving only casts of replaced minerals. This

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thoroughly pervasive texture is common where hydrothermal breccia occurs. Disseminated pyrite is commonly associated with silica replacement. Silicification halos surround mineralized veins for thicknesses up to 50 m.

Quartz-Sericite alteration

·Quartz-sericite alteration surrounds veins and silicified zones and generally indicates the zones proximal to mineralization. The alteration is characterized by homogenous zones of mixed quartz and sericite that can be up to 100 m thick.

Argillization

·Argillic alteration commonly forms in select fault and shear zones. Commonly clay (kaolinite), sericite, and jarosite form within narrow (centimeter to meter wide) zones as the alteration products of feldspar, biotite and rock matrix.

Propylitization

·Propylitic alteration forms as weakly pervasive and stronger fault-controlled zones of chlorite-calcite-pyrite. Propylitic alteration forms furthest from mineralization and commonly borders fresh-unaltered rock.

7.7ESCOBAL VEIN ZONES

The Escobal vein is divided into three primary zones, as described below and illustrated in Figure 7-6.

East Zone

·In the East Zone, the Escobal vein follows an east-west to N80E normal fault that dips variably (60-75°) to the south and can be followed for at least 600 m on strike (807,200E to 807,800E).

·The East zone demonstrates a dilational jog similar in character and at roughly the same elevation as found in the Central Zone. The upper reaches of the East Zone vein follows south dipping fault for roughly 500 to 600 m to approximately the 1250 m elevation where it changes dip towards to vertical or steeply towards the north. This deeper extension remains partially open and untested down-dip to the north and laterally to the east and west.

·Widely-spaced step-out drilling in the deep eastern margin of the East Zone has identified mineralization that can be followed for 350 m along an east-west strike and to a 400 m depth. The “East Extension” comprises multiple zones of low-grade, moderate- width (two to 20 m) veins and stockwork zones with near vertical dips. Mineralization extends from 1400 to 1000 meter elevations and remains open to depth. The zone is capped by a 300-m-deep zone of un-mineralized narrow veining that was recognized in earlier drill campaigns.

·Geochemistry in the main East Zone is characterized by a gold-rich sector in the near surface oxide/mixed sulfide-oxide zone that abruptly changes at depth to silver-rich mineralization across the sulfide interface. Lead and zinc concentrations show a strong correlation to silver mineralization in the lower portion of the sulfide zone and increase with depth relative to silver. A gradational zoning pattern is observed with silver giving way to lead and then zinc with depth. The East Extension is characterized by high silver and relatively low grades of lead, zinc and gold.

Central Zone

·The Central portion of the Escobal vein is the thickest part of the vein system. The zone extends 700 m on strike and covers a nearly 800 m vertical range, from outcrop at 1500 m elevation to the deepest drill intercept at 700 m elevation. The zone strikes east-west with the main portion of the vein dipping moderately (60-70°)

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to the north. Flexures in mineralization in the upper and lower reaches of the Central Zone are controlled by high-angle faults. The wide, moderately north-dipping main portion of the vein represents mineralization along the dilational jog, or tensional shears between two major normal faults.

·High-grade silver occurs throughout the Central Zone, the bulk of which forms a wide roughly horizontal zone related to the north-dipping dilational structure. The zone narrows towards the east and exhibits greatest vertical extent on its western margin where it abruptly terminates along a barren “gap zone” bounding the western Margarito area. Lead and zinc concentrations correlate extremely well with silver grades in the Central Zone, though silver grades are maintained at depth, contrary to the gradational Ag-Pb-Zn vertical zoning that is evident in the East Zone.

West Zone

·The West mineralized zone is characterized by surficial gold occurrences that give way to a wide zone of silver, gold, and base metal rich vein stockwork at depth. The deeper “Margarito” mineralization is a discrete shoot as it is separated from the Central Zone and the upper gold zone by a 50 to 100 m wide barren “gap” in mineralization. The zone as currently modeled extends over a 350 meter strike length and spans +550 m vertically, raking down to the east. The top of mineralization is entirely preserved with significant grades commencing 250 m below the surface. The zone is partially open to depth, while the western margin is believed to be down-dropped further west along a normal basin-bounding fault, interpreted through marker-bed offset.

·The West Zone follows a semi-arcuate trace with moderate north dips in the upper reaches of the zone giving way to steep, near vertical inclination at depth. The upper portion of the zone is characterized by high gold values in the mixed-oxide zone. The deeper Margarito shoot exhibits very wide (30-50 m) zones of stockwork-veining with moderate silver grades and moderate-high gold grades throughout. Base metal values show a marked increase in the lower portion of the zone.

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Figure 7-6: Escobal Long Section (East-West Section — Looking North)

Oxidation

·The bulk (~95%) of the deposit is unoxidized. Wall rock oxidation was modeled on 50-m-spaced north-south drill sections with oxidation to depths up to 200 m while the vein itself, due to its relative permeability, is partially oxidized from the surface to locally 250 m depth. Secondary manganese oxide is concentrated near the base of the oxide zone.

·The lower boundary of mixed oxide-sulfide is defined by the last observation of limonite. Oxidation within vein intervals above the sulfide boundary vary from moderately to completely oxidized in drill holes. Primary sulfide versus oxide concentrations increase with depth, from none near surface, to 100% unoxidized at the sulfide boundary. There are intensely oxidized vein intersections with high gold grades in the upper levels of the East Zone which may be amenable to leach processes not considered at this time.

7.8VEIN MODEL

Vein attributes have been compiled to support metallurgical sample collection and to aid on-going exploration. Physical attributes include estimated true vein width and vein volume percent across the defined zones. Mineralogic attributes include observed mineral abundance estimates of iron oxide/sulfate, manganese oxide, proustite, and total sulfide. Geochemical attributes include average Ag, Au, Pb, Zn, Cu, As, and Sb contents for each vein intercept, as well as calculated Ag/Au, Ag/Pb, and Ag/Cu ratios. All vein attributes were contoured on a long section in the plane of the vein

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(vein intercepts were projected horizontally at 90 degrees to a common east-west plane). Key observations include the following:

·Four high-grade (+500 g Ag/t) Ag-(Au-Pb-Zn) “ore shoots” are defined by drilling. The East, Central and West/Margarito zones are well-defined by drilling while the East Extension zone remains partially open. The East, Central and West/Margarito zones contain bonanza-grade (plus 1,000 g Ag/t) intervals.

·The East Zone ore shoot begins approximately 100 m beneath the surface, is at least 600 m in strike length, and spans a 500-600 m vertical range.

·The Central Zone ore shoot begins approximately 50 m beneath the surface, is at least 700 m in strike length and spans approximately 800 m vertically.

·The East Extension ore shoot begins approximately 300 m beneath the surface. The zone is partially defined by drilling and is open to the west, east and down dip. As currently defined, the zone covers a 500 m strike length and 400 m vertical range.

·The West/Margarito Zone begins approximately 250 m below the surface and as currently defined extends over a 350 m strike length and 550 m vertical range. The zone remains open to depth, and also to the west where it is believed to be offset by faulting.

·Distribution of gold, silver, lead and zinc show a general trend of mineralization with a gentle (~20°) rake, down to the west; in effect the East Zone is about 200 m higher in elevation than the Central Zone, which is in turn is about 200 m higher in elevation than the Margarito Zone.

·Individual ore shoots in the East and Margarito zones show rakes 20-50°± to the east in the plane of the vein. The Central Zone is roughly horizontal with a more extensive vertical plume along its western margin. The East Extension, as currently defined, trends sub-horizontal with an apparent moderate (~ 50°) rake to the east where open to depth.

·Gold and arsenic are erratically anomalous above and peripheral to the higher grade Ag-Pb-Zn-(Au) in partially oxidized vein in the West and East zones. Deep drilling in sulfide-rich portions of the western Central and Margarito zones show irregular zones of elevated gold values.

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8                      DEPOSIT TYPES

The Escobal deposit formed in an intermediate-sulfidation epithermal quartz vein system of probable Upper Miocene to Lower Pliocene age. These deposits are commonly included in the low-sulfidation epithermal class of deposits. Distinguishing characteristics of an “intermediate sulfidation” environment include mineral assemblages indicating a sulfidation state between those of high and low sulfidation types, relatively high total sulfide content of 5 to 10 percent, low-iron “blond” sphalerite, presence of silver sulfosalts, and association with andesitic to dacitic volcanics. Magmatic-associated fluids are implied.

Epithermal deposits form as high-temperature mineralizing fluids rise along structural pathways and deposit quartz and precious- and base-metal minerals in open spaces in response to boiling, which is usually coincident to a release of pressure within the hydrothermal system. This quartz and metal deposition, followed by resealing of the system, is repeated over the life of the hydrothermal system resulting in crosscutting and overprinted breccia and vein textures. Typically, the largest and highest grade deposits are associated with long hydrothermal systems marked by complex overlapping veins.

These deposits are strongly structurally controlled. Mineralizing fluids are directed along structural pathways with high-grade “ore shoots” typically concentrated in open dilatant zones. These dilatant zones commonly form where inflections occur vertically and laterally along the vein.

Metal deposition and zoning in epithermal deposits are related to the level of boiling. Typically, precious metals deposit at or near the boiling level while base metals precipitate below. Boiling may occur at different levels as the hydrothermal system evolves producing an overprint of various episodes.

The Escobal deposit occurs in a similar geologic setting with host rocks, vein characteristics and mineralogy typical of other intermediate-sulfidation systems. Specific definitive features include banded, cockscomb, and drusy vein textures; massive, stockwork and breccia veins; intermediate argillic and quartz-sericite alteration; appreciable base-metal and silver-sulfosalt mineralogy, and associated arsenic and antimony.

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Figure 8-1: Spatial Relationship of Intermediate Sulfidation Deposits (after Corbett, 2002)

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9                      EXPLORATION

The exploration strategy at Escobal utilizes straightforward exploration techniques that include surface prospecting for vein, altered outcrops and float; subsequent detailed geologic mapping and surface geochemicial sampling; and drill testing. The most effective exploration tool at Escobal has been core drilling, the results of which are discussed in Section 10, Drilling.

Evolution of the mineralization model at Escobal has been instrumental in identifying the project’s resource potential. The recognition of deeper silver and base metal zones below the remnant near-surface gold-bearing cap was paramount in the discovery success. Of equal importance was the understanding of the structural controls of the variably-dipping Central Zone vein and, more recently, recognition of the change in dip of the deep East Zone vein. Discoveries of the West/Margarito and East Extension zones continue to add resources to the project as deeper portions of the vein zones continue to be defined.

Step-out drilling has been successful in identifying mineralization both laterally and below the originally defined Escobal resource areas. These discoveries resulted from systematic drilling along the projections of geochemical trends modeled from prior drill data. The new information gained contributes to the understanding of the distribution, zoning and strength of the mineralized system. Based on the results to date, it is believed that there remains significant potential for discovery of still unrecognized mineralization along the Escobal structure.

Several other vein targets have been identified in the district using the geologic model developed from the Escobal vein. At the same time, geologic mapping and prospecting continues to help identify other styles of mineralization in the district that may include mineralized intrusive and breccia bodies.

Supplementary studies have been undertaken to aid interpretation and discovery of buried veins throughout the region. An alteration zonation model to enhance interpretation and help to identify other district exploration targets was conducted using spectroscopic (Terraspec®) analysis on the Escobal vein and other veins proximal to Escobal. Results to date suggest illite and kaolin concentrations in wallrock surround the Escobal vein and grade outwards to smectite-rich alteration both laterally and to depth.

9.1                  GEOPHYSICS

Both ground magnetic and Induced Polarization (IP)/resistivity orientation surveys have been conducted by Tahoe over portions of the Escobal and neighboring secondary veins. Ground magnetics is not viewed as a valid exploration method as the low magnetite content associated with mineralized veins and the interference from magnetite-rich capping Quaternary sediments (ash and alluvium) provides chaotic magnetic results. IP and resistivity results show better correlation with mineralized veins only where veins with higher sulfide content extend to a shallow level (~150m depth) below minimal Quaternary cover.

9.2                  GEOCHEMISTRY

Silver and gold mineralization in the Escobal vein is typical of intermediate-sulfidation deposits with an associated epithermal suite of elements including arsenic, antimony, lead and zinc. Generally, high arsenic, lead and zinc grades correlate with anomalous silver mineralization. Moderate correlations are also observed between silver and antimony, and between gold, arsenic and lead. Manganese is anomalous throughout the deposit, both as pyrolusite in the oxide portion and possibly as a product of sphalerite in the non-oxide portion of the deposit.

Soil sampling was completed by Entre Mares, S.A. prior to Tahoe’s acquisition of the Oasis concession in 2010. Entre Mares sampled at 100 m by 25 m spacing over the entire Escobal vein and adjoining areas. Soil and rockchip anomalies confirm trends identified through geological mapping and drilling (Figure 9-1 and Figure 9-2).

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Figure 9-1: Soil and Rockchip Chemistry — Silver

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Figure 9-2: Soil and Rockchip Chemistry — Zinc

The style of mineralization and geochemical zoning patterns vary laterally across the deposit as determined from geochemical data collected from surface sampling programs and geochemical analyses of drill core.

East Zone: East Zone geochemistry is characterized by a gold-rich signature in the near surface mixed/oxide zone (supergene enrichment) that abruptly changes with depth to silver-rich mineralization at the oxide/sulfide interface. A small zone of gold mineralization occurs deep at the east edge of the East Zone, at a similar elevation and possibly related to elevated gold mineralization in the deep Central Zone.

Anomalous lead and zinc concentrations are related to silver mineralization in the sulfide zone. A gradational zoning pattern is observed with silver giving way to lead and then zinc with depth. Absolute lead and zinc values increase with depth relative to silver, suggesting that a pure base metal zone may occur below the extent of the current drilling.

Arsenic is coincident with gold mineralization in the East Zone. Generally, arsenic is vertically constrained throughout the deposit, occupying a horizon between 1200 to 1600 m. In the East Zone, anomalous arsenic correlates with the two east-plunging gold “ore shoots” in the mixed/oxide zone, and is irregularly dispersed below gold mineralization in the sulfide zone. Anomalous antimony correlates well with silver and shows a slightly wider dispersion pattern than arsenic.

Central Zone: Geochemistry in the Central Zone is distinctive as no significant near-surface gold is observed. Silver mineralization occurs at a slightly lower elevation than the East Zone and remains open to the east. Anomalous gold grades occur at depth in the central core of the Central Zone with anomalous gold grades at 1350 m elevation (approximately 100-150 m below surface) extending to a deeper intercept at 1025 m elevation. This deep gold

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mineralized core of the Central Zone is believed to represent a distinct zone unrelated to the near-surface gold zones observed in the East and West zones.

Lead and zinc concentrations correlate extremely well with silver in the Central Zone, though gradational silver-lead-zinc vertical zoning is not evident as in the East Zone. Arsenic forms as a large dispersion halo west and above Central Zone mineralization. Antimony correlates well with and shows a minor dispersion around gold-silver mineralization.

West Zone: The West Zone is the most geochemically inconsistent portion of the deposit characterized by surficial gold occurrences, lack of distinct zone of silver mineralization and irregular lead, antimony and arsenic anomalies below the surface gold zone. The West Zone surface gold anomaly occupies a similar elevation range as the upper mixed/oxide East gold zone and is interpreted as the erosional remnant of the same zone. As exploration drilling in the west zone has been largely geared towards definition of the near-surface gold targets, deeper drilling is required to explore zones of deeper silver and gold mineralization and evaluate geochemical signatures in this area.

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10                   DRILLING

Drill targeting of the Escobal and ancillary veins has been conducted by Entre Mares and Tahoe from 2007 to the present, with 946 exploration and infill/definition drill holes totaling 249,392 m completed within the boundaries of the Escobal exploitation concession through July 1, 2014. Surface drilling was done using both contractor- and company-owned drill rigs; beginning in 2012, infill and definition drilling from underground drill stations has been conducted using drills owned and operated by the Tahoe.

Data acquired through January 23, 2014 have been used for the Escobal resource model and estimate reported herein; the dataset used for resource estimation is comprised of 842 drill holes totaling 231,326 m. The dataset includes 21 metallurgical and three piezometer drill holes.

Summaries of the drilling completed to date and drill holes used for the resource estimate are presented in Table 10-1 and Table 10-2, respectively. Figure 10-1 is a plan map illustrating the surface drill hole locations; Figure 10-2 and Figure 10-3 are drill hole cross sections through the Central and East zones of the Escobal deposit, respectively.

Table 10-1: Escobal Exploitation Concession Drilling (Project-to-Date through July 1, 2014)

Company	Target	Purpose	No. Drill Holes	Total Length (m)
Entre Mares	Escobal	Exploration	213	58,156
	Areneras	Exploration	3	601
	Granadillo	Exploration	2	425
		Entre Mares Total	218	59,182

Tahoe Resources*	Escobal	Exploration/Infill (from surface)	222	128,811
	Escobal	Infill/Definition (from underground)	498	56,460
	Escobal	Metallurgy	21	4,943
	Escobal	Piezometers	3	900
	Areneras	Exploration	3	1,465
	Beto	Exploration	5	2,631
		Tahoe Resources Total	752	195,210
	Escobal Exploitation Concession Total		970	254,392

*excluding ancillary drill holes not sampled for assay, i.e., water wells, monitor wells, geotechnical holes, etc.

Table 10-2: Escobal Drilling Included in Resource Estimate (Drill Data Through January 23, 2014)

Area	No. Drill Holes	Total Length (m)
East Zone / East Zone Ext.	181	74,671
Central Zone	560	107,527
West Zone / Margarito	89	44,674
Other	12	4,454
Total	842	231,326

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Figure 10-1: Escobal Drill Hole Location Map

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Figure 10-2: Escobal Central Zone Cross Section 806800E

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Figure 10-3: Escobal East Zone Cross Section 807500E

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10.1SURFACE DRILLING

Nearly all surface drilling at Escobal has been by diamond drill (core) methods, using 1.52-m and 3.04-m (5-ft and 10-ft) core barrels. The majority of mineralized intercepts were drilled using NTW-size or larger drill core, with lesser amounts of NQ2-, BTW-, and BQTK-size drill core. Six diamond drill holes were precollared through unmineralized rock using reverse circulation (RC) drilling; four drill holes precollared by RC have yet to be continued with diamond drilling. In addition, 37 small diameter (AQ-size) Winkie core holes were drilled at Escobal in 2010 and 2011; these drill holes were used as a ‘first pass’ prospecting tool or to gather near-surface geologic data for future plant site construction planning. No RC or Winkie drill samples are included in the drill hole database used for the resource estimate.

10.1.1Drill Campaigns

10.1.1.1Entre Mares

Entre Mares conducted drill campaigns in the Escobal project area from 2007 to June 2010, during which time they completed 218 diamond drill holes totaling 59,182 m. Entre Mares’ drilling was conducted by Kluane Guatemala S.A. (a division of Kluane International), using KD600 and KD1000 drill rigs, and by Entre Mares personnel, using company-owned Hydracore drills.

10.1.1.2Tahoe Resources

Upon acquisition of the Escobal property in June 2010, Tahoe continued the exploration drilling program begun by Entre Mares. From June 2010 through June 2014, 222 drill holes totaling 129,676 m were completed on the Escobal vein and eight drill holes totaling 4,064 m completed on the ancillary Areneras and Beto veins. Surface drilling has been done by Kluane Guatemala S.A. using the same KD600 and KD1000 drill rigs previously used during the Entre Mares drill campaigns. In addition, Tahoe purchased a Hydracore 2000 man-portable drill and a Longyear LM-75 drill in early 2011; both of which were utilized throughout the 2011-2013 exploration drill programs. Tahoe has since continued surface drilling using the Hydrocore drill.

Beginning in 2011, larger drills were contracted to achieve greater depth to targets in the Escobal vein. Canchi Drilling drilled several holes using a JS-1500 drill in the West Extension zone in 2011 but failed to achieve desired target depths. Island Drilling was contracted in 2012 and 2013 and completed a number of deep (+1500 m) holes in West Extension zone utilizing AR-250 and LF-230 truck-mounted drills. In early 2013, Kluane Drilling also provided a larger man-portable drill (GM-282) capable of drilling +1200 m depths to test deep targets throughout the project area.

From mid-August 2010 to early 2011, Tahoe conducted a diamond drill program to acquire core samples specifically for metallurgical and physical property testing. Rodio-Swissboring Guatemala S.A. was contracted to drill five large diameter (PQ-size) core holes to acquire samples for comminution testing. No assay data was obtained from this drilling. An additional 16 core holes were drilled by Kluane Guatemala S.A. and Tahoe (HQ- and NTW-size drill core) to obtain samples for metallurgical variability tests. Assays from the variability test samples are included in the resource estimate.

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Figure 10-4: Surface Exploration Drilling

10.2UNDERGROUND DRILLING

In 2012, Tahoe initiated resource infill and stope definition drilling from underground drill stations located along the primary and secondary development headings in preparation of production mining in the Escobal Central. All underground drilling has been done by Tahoe personnel using two Longyear LM-50 drills and one Longyear LM-75 drill. Through July 1, 2014, the company had completed 498 underground drill holes totaling 56,460 m. Drilling from underground stations has been by diamond drill methods using 1.52-m (5-ft) core barrels with NQ- and BQ-size tools. To date, underground drilling has been limited to the Central Zone. Infill and stope definition drilling will continue throughout the life of the mine.

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Figure 10-5: Underground Infill Drilling

10.3DATA COLLECTION

Data collection procedures for surface drilling remained generally consistent between Entre Mares and Tahoe; hence, the following descriptions are applicable to both companies’ drilling programs, except as noted. Likewise, data collection procedures for underground drilling are consistent with the surface drilling protocols, except as noted.

10.3.1Drill Core Handling

As the core barrel is retrieved from the drill hole, the core is removed and placed in wooden core boxes along with markers labeled with the downhole distance. The core boxes are labeled and transported by to the company’s core logging facility at the project site, where company geologists and technicians wash and photograph the core, record the geologic and geotechnical characteristics, and mark drill core intervals for sampling. Core from the surface drill holes is sampled by sawing the core in half longitudinally. Core from the underground drill holes is sampled in its entirety, with periodic sample intervals sawed in half for archive. After logging and sampling are complete, the core boxes are transported either to a secured storage facility in San Rafael Las Flores or to storage facilities at the project site, where it is stored on racks inside covered buildings.

In September 2012, roughly 63,000 m of core from 106 drill holes (E11-287 through E12-389) were lost in a fire at the mine core storage facility and were unavailable to MDA for examination. All of the core from these drill holes had been logged, photographed, sampled and assayed prior to the fire.

10.3.2Drill Collar Surveys

At the completion of each surface drill hole, the collar locations are marked in the field with a four-inch plastic (PVC) pipe cemented into the top of the drill hole. The drill hole identification number is indicated by permanent marker on the PVC pipe and etched in the cement at the collar. All drill collar locations were surveyed by Sergio Diaz (2007-

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2010) or Geotecnología S.A. (2009-2014), both independent professional surveyors based in Guatemala City. Collar locations were determined using non-differential global positioning system (GPS) instruments and post-processed. In some cases where topography or heavy vegetation prevented collection of accurate measurements by GPS, the surveyor employed a Total Station instrument using established drill collar locations or project control points with known coordinates as bench marks. All drill hole collar coordinates are reported in UTMm coordinates, NAD 27, Zone 15 and converted to the Escobal site Cartesian coordinate system. The original reports received from the surveyor are archived at the Escobal project office and the reported collar coordinates stored in the Escobal project digital database.

Underground drill hole collars are surveyed by the Escobal engineering department in the site Cartesian coordinate system using a Total Station instrument. The survey data is downloaded and stored in the Escobal digital project database.

10.3.3Downhole Surveys

Downhole survey measurements of surface drill holes are taken by the drill contractor or company personnel at approximate 50 m intervals and at the final depth of each drill hole. Entre Mares used a Tropari down-hole survey instrument through to mid-2009, after which they used a Reflex EZ-shot digital downhole survey tool. All surface holes drilled by Tahoe were surveyed using the Reflex EZ-shot tool.

Underground drill holes are surveyed downhole at approximate 50 m intervals and at the final hole depth by company personnel using a Reflex EZ-shot digital tool.

A 3° west magnetic declination correction is routinely applied to raw azimuth readings for all drill holes. The survey readings are entered into the Escobal project digital database with the original survey datasheets archived at the Escobal exploration or mine offices.

10.3.4Geological Logging

Geologic data from drill core is originally recorded on paper logging forms and then entered into the digital project database. Data documented from the drill core includes lithology; primary and secondary rock textures, vein lithology; mineralization and alteration; estimated sulfide content; structural features, including the angle of structure to the core axis; and degree of iron oxidation.

10.3.5Geotechnical Logging

The majority of drill core has been logged for geotechnical data. Geotechnical data collected from the drill core includes core recovery, hardness, rock quality designation (RQD), joint number (Jn), joint roughness (Jr), joint alteration (Ja), joint water reduction factor (Jw), and the stress reduction factor (SRF); all of which is entered into the project database. From this data, geomechanical classifications — tunneling quality index (Q rating) and rock mass rating (RMR) — are calculated to identify the ground control measures appropriate for the rock quality anticipated during underground excavation.

10.4DRILLING SUMMARY AND RESULTS

Both the Entre Mares and Tahoe surface drilling programs at Escobal targeted the vein system with diamond drill core holes oriented perpendicular to the general east-west strike direction of the deposit, at varying inclinations to explore the deposit along dip. Underground drill holes completed by Tahoe were collared on the south side of the vein system and drilled northerly at varying azimiths and inclinations to infill between vein intercepts from surface drilling and to define ore boundaries for stope design.

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To date, drilling has defined mineralization over approximately 2,400 m of strike length and 1,200 m of total vertical extent; elevations range from 800 to 1720 m above sea level (masl) in the East Zone, from 700 to 1540 masl in the Central Zone and from 550 to 1400 masl in the West Zone. In general, the Escobal deposit has been drill-delineated from the surface in the east-west direction (i.e., along strike) on nominal 50-m spaced intervals, with numerous holes drilled between the 50-m intervals, particularly in the East Zone. Additional drilling from underground has supplemented the surface drill spacing, locally decreasing the drill spacing in the current and near-future production areas to as much as 20 m along strike and 10 m vertically. The drill sample spacing is sufficient for the geologic modeling and resource estimation of the Escobal vein system.

The East Zone and East Zone Extension strike approximately azimuth 80° and dip to the south from 60° to 80°, with an average dip of approximately 70° south. As such, the majority of holes drilled from the surface are oriented to the north, with a few holes oriented south to explore for north-dipping secondary veins. Drill hole lengths range from 9.9 meters to 1021.1 m, with an average drill hole length of 417.3 m.

The Central Zone strikes approximately east-west with variable dips, with the western portion (West/Margarito Zone) trending slightly to the north of west. The mineralized structure generally dips from 60° to 70° to the north from the surface down to around 1200 m elevation and steepens to near-vertical at depth. The upper portion of the deposit in the eastern half of the Central Zone dips 60° to 70° to the south (East Zone orientation). Accordingly, drill holes were oriented northerly to explore the south-dipping portion of the mineralization and southerly to explore the north-dipping portion of the mineralization. Surface drill hole lengths in the Central Zone range from 8.2 m to 1016.2 m, averaging 368.0 m. Underground drill hole lengths range from 28.0 m to 302.1 m, averaging 113.4 m.

10.5CORE RECOVERY — METAL GRADE ANALYSES

MDA evaluated the relationship between the surface drilling core recovery and metal grades in 2010 and the underground core recovery in 2014.

Overall, the surface drill core recovery data indicate good to excellent core recovery within the mineralized horizons. The underground drilling has a greater proportion of moderate to poor core recovery drill intervals, as compared to the surface drilling, but there is no evidence of a grade bias being imparted into the drill data due to the core loss. The core recovery data and metal grade analyses support the estimation of the Escobal resource to Measured and Indicated levels.

10.5.12010 Analyses — Surface Core Recovery versus Metal Grades

Tahoe provided MDA with the core recovery data for all 218 core holes used in the 2010 resource estimate. MDA checked the recovery data calculations and spot-checked the measurements against the core photos.

The average core recovery for all surface exploration drill intervals is approximately 96 percent. The average core recovery for the mineralized intervals used in the resource estimate is approximately 95 percent. Approximately 65 percent of the core recovery measurements have values of 100 percent recovery. The prevalence of exact 100 percent core recovery values is indicative of the massive, weakly fractured nature of the country rock but also suggests possibly less rigorous measurement techniques.

MDA analyzed the relationship between metal grades and core recovery. All four metals (silver, gold, lead, and zinc) were reviewed independently. Figure 10-6 shows the relationship between silver grades and core recovery. The silver grade and number (“Count”) of core recovery intervals are presented in the left-hand and right-hand y-axis, respectively. These values are sorted into core recovery “bins” of regular 10 percent intervals as noted along the x-axis. (Each bin represents all intervals within each 10 percent interval; for example, recovery column “80” shows the average silver value and number of sample intervals for all intervals with core recovery values between 80 and 89 percent.) The data

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shown in Figure 10-6 were filtered for only those sample intervals with silver grades greater than 10 g Ag/t to better represent the core recovery effect on significant silver grades.

The data in Figure 10-6 show a noticeable decrease in average silver grade when core recovery drops below 80 percent. Only a small fraction (<10 percent) of all sample intervals have recovery values below 80 percent, so the observed change in silver grade is not believed to result in a material error in the current resource estimate. Since the assay values have been potentially down-graded due to the core recovery loss, the current resource estimate is on the conservative side, indicating a small upside to the resource estimate.

MDA analyzed the gold grade versus core recovery data, and a similar pattern as seen in the silver data was observed.

Figure 10-6: Core Recovery — Silver Grade Comparison

The lead versus core recovery data indicate that there is an approximate 25 percent increase in lead grade when core recovery decreases from 100 percent into the 80 percent recovery category. Below 80 percent, the lead data have a similar, though somewhat more erratic, decrease in values as was seen above in the silver data. The latter observed decrease in lead grade is not considered significant due to the small fraction (<10 percent) of all samples with recoveries less than 80 percent. An analysis of the zinc data shows the same relationship with core recovery as seen in the lead data.

The increase in lead and zinc grades associated with the 80 percent and 90 percent core recovery intervals could be directly related to a selective increase in grade from core loss. It also could be a natural function of the geology in which the higher-grade, base-metal mineralization occurs preferentially within highly fractured structural intervals. The observed lower core recoveries within these intervals would then have a purely spatial correlation, not genetic, with the increased grades. Further analyses of the data would be needed for a more precise determination of the relationship between core recovery and base-metal grades.

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10.5.22014 Analyses — Underground Core Recovery versus Silver Grades

Tahoe provided MDA with the core recovery data for the underground core holes used in the 2014 resource estimate. MDA checked the recovery data calculations and spot-checked the measurements against the core photos. Only minor errors were found in the calculations which were corrected before the core recovery/silver grade analyses.

The average core recovery for all underground drill intervals is approximately 89 percent while the average core recovery for the mineralized intervals assaying greater than 50 g Ag/t is approximately 86 percent. These average core recovery values are lower than the surface exploration core recovery values.

MDA analyzed the relationship between silver grades and underground core recovery (Figure 10-7) using only those intervals assaying greater than 50 g Ag/t to better represent the sample intervals contributing to the current resource estimate. As in Figure 10-6, the silver grade and number (“count”) of core recovery intervals are presented in the left-hand and right-hand y-axis, respectively. The histogram bars represent the average silver grades for each core recovery “bin” while the “count” at each bin is indicated by the light blue line connecting the red data points.

Figure 10-7: Underground Core Recovery — Silver Grade Comparison

Figure 10-7 shows that there is no apparent relationship between core loss and silver grades within the underground drilling. Average silver grades remain unchanged with decreasing core recovery though there is some increased variability in the lower recoveries, as would be expected due to the smaller population of samples within these core recovery ranges. Core loss has not introduced a grade bias into the underground drill data and subsequent resource estimate.

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11SAMPLE PREPARATION, ANALYSES AND SECURITY

Sample preparation, analyses, and security procedures of surface drill hole samples have generally remained consistent between Entre Mares and Tahoe. The Escobal exploration manager and many of the on-site geologic personnel responsible for the drill core sampling remained in place following the transfer of the property from Entre Mares to Tahoe Resources in 2010; as such, the following descriptions are applicable to the practices of both Entre Mares and Tahoe, except as noted. Likewise, sample procedures used by mine geology personnel for underground drill samples are generally consistent with those used for surface drill samples, except where noted.

With few exceptions, Entre Mares and Tahoe used BSI Inspectorate (now named Inspectorate, a division of Bureau Veritas) as their primary analytical laboratory. Inspectorate operates a sample preparation laboratory in Guatemala City where sample pulps are prepared and shipped to Inspectorate’s laboratory facility in Reno, Nevada. There is no affiliation between Inspectorate and Tahoe except that of an independent laboratory/client relationship.

Inspectorate holds current ISO 17025 and ISO 9001:2000 certifications. ISO 17025 specifies the general requirements for the competence to carry out tests and or calibrations, including sampling. ISO 9001-2000 specifies requirements for a quality management system where an organization needs to demonstrate its ability to consistently provide a product that meets customer and applicable regulatory requirements and aims to enhance customer satisfaction through the effective application of the system, including processes for continual improvement of the system and the assurance of conformity to customer and applicable regulatory requirements.

11.1SAMPLE METHOD AND APPROACH

Company geologists determine drill core sample intervals once the drill core has been logged for geologic and geotechnical properties (as described in Section 10.3, Data Collection). Surface drill sample intervals generally vary from less than one meter to one-and-one-half meters in zones of discreet mineralization, and from three meters to locally six meters in weakly mineralized or altered areas. These sample lengths are appropriate for the differing styles and distribution of mineralization at Escobal, though it is recommended that sample intervals do not extend across obvious mineralogic contacts. Intervals of ‘fresh’ unaltered rock are normally excluded from sampling. Underground drill sample intervals are usually 1 m or 1.5 m in length, though irregular sample lengths are common as the sample breaks are based on geologic and mineralogic contacts. Once the sample intervals are determined, the core is marked, sample tags are stapled to the core box dividers, and the core boxes are photographed.

Core samples selected for analysis from surface drill holes are cut lengthwise using mechanized diamond saws. One-half of the core is placed in a plastic sample bag with a sample tag. The remaining half core is replaced in the core box for future reference. The mineralized zones at Escobal are often quite wide (up to 50 m) and complex (multiple cross-cutting vein events). The practice of submitting one-half of the core provides a reasonable representation of the mineralization for analysis. Stope definition and infill drill core obtained from underground drilling is normally not split, with the whole core sent for analyses.

11.2SAMPLE SECURITY

After the drill core is logged, photographed, and sampled at the project site, the samples are taken to San Rafael Las Flores, where they are stored in the company’s secured office/warehouse facility or held at the mine site until delivered to Inspectorate’s sample preparation laboratory in Guatemala City. From 2007 to 2008, all samples were picked up by Inspectorate at the San Rafael office. Since 2008, the samples have been delivered to the Inspectorate prep lab using Tahoe’s drivers and vehicles. BSI holds duplicate sample pulps in secured storage in Guatemala City and returns them on a routine basis to the Tahoe exploration department in San Rafael or to the mine.

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11.3LABORATORY SAMPLE PREPARATION

Since the initiation of drilling at Escobal in 2007, drill core samples have been prepared by BSI Inspectorate at their preparation facility in Guatemala City. After drying, core samples were crushed to >80 percent passing 2 mm (10 mesh) using a jaw crusher and roll mill. The crushed samples were then passed through a Jones riffle splitter to obtain a nominal 300 g sample for pulverization. The 300 g subsample was pulverized to >90 percent passing 150 mesh and split into two sample pulps for primary and check analyses. Barren sand is used to clean the pulverizer after every sample; one sample of the barren sand is inserted into the sample stream per batch where it is reported as an internal laboratory blank. Inspectorate packaged and air-freighted one set of pulps to their laboratory in Reno, Nevada for analysis and delivered the second set of pulps to Entre Mares and Tahoe at site.

11.4LABORATORY ANALYSES

Inspectorate in Reno, Nevada is the primary laboratory for nearly all of the drill core analytical work at Escobal, with the exception of 79 metallurgical samples from Entre Mares’ 2008 drilling campaign that were assayed ALS Chemex (Chemex) for Au and Ag, 315 samples from Tahoe’s 2010 metallurgical program that were assayed at Cardwell Analytical (wet assays) and Chemex (ICP), and 1,284 samples from ten underground infill drill holes that were assayed at the Escobal on-site laboratory.

Inspectorate determined silver grades using aqua regia digestion followed by atomic absorption spectrometry (AAS) and, to a lesser extent, induced-coupled polarization (ICP). The use of ICP for silver grade determination was discontinued in late 2007. For initial silver results exceeding 200 g/t, Entre Mares instructed Inspectorate to automatically re-analyze the sample using fire assay with gravimetric finish. Tahoe continues this practice, but uses a lower grade threshold of 100 g/t.

Gold analyses were done by fire assay (one assay-ton) followed by AAS. Samples returning more than 3 g/t were re-assayed with a gravimetric finish. Assayed sample mass varied depending upon level of sulfide in the sample to limit losses caused by boil-over in the assaying process.

Sample pulps were also often analyzed for a multi-element geochemical suite using aqua regia digestion followed by ICP. Lead, zinc, and copper values exceeding 1% were re-analyzed by aqua regia/AAS, which has a higher grade determination threshold than ICP. Base metal samples exceeding the threshold of AAS were assayed using titration methods.

11.5QUALITY ASSURANCE / QUALITY CONTROL PROCEDURES

Quality assurance/quality control (QA/QC) procedures for drill core sample analyses include the use of standard reference materials, sample blanks, check assays, and duplicate samples. Results and analysis of the QA/QC data are presented in 12.0, Data Verification.

11.5.1Standard Reference Materials

Tahoe incorporates blind standard reference material (assay standards) of varying metal grades into the sample stream prior to submission of the samples to Inspectorate at a rate of one standard per 20 drill samples (5%). Tahoe used four different standards for the 2010 and 2011 surface drill samples and eight standards for the 2012 and 2013 surface drill samples; all of which were commercial standards prepared by CDN Resource Laboratories of Langley, B.C., Canada. Six commercial standards obtained from WCM Minerals (“WCM”) of Burnaby, British Columbia were used for submission with underground drill samples. Entre Mares did not use assay standards in their QA/QC program at Escobal.

Inspectorate also included reference materials (both in-house and certified reference materials) in its QA/QC program.

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11.5.2Blanks

Entre Mares and Tahoe inserted sample blanks into the sample stream at irregular intervals to check for contamination during the laboratory sample preparation stage. Samples assaying less than five parts per billion gold and 0.1 parts per million silver were collected from local outcrops at Escobal for use as sample blanks. These samples are valid as gold and silver blanks, but do contain trace amounts of lead and zinc. Inspectorate also monitored pulverizer contamination by collecting and analyzing barren sand used to clean the pulverizer between samples.

11.5.3Check Assays

Check assay programs have been continually in effect at Escobal since the initiation of the exploration drilling campaigns by Entre Mares, with sample pulps re-assayed by laboratories other than Inspectorate, including samples submitted to more than one outside laboratory for redundant check assaying.

Entre Mares submitted the second pulp split prepared by Inspectorate to a second laboratory for check assaying for nearly all mineralized drill intercepts. Entre Mares used the SGS-managed laboratory at the Marlin Mine in Guatemala for silver and gold check assays from May 2007 through May 2008, and again from July 2009 through the end of their involvement with the property in June 2010. From June 2008 through July 2009, Entre Mares used CAS Honduras for the silver and gold check assaying. A small percentage of samples were also re-assayed by ALS Chemex in Vancouver.

SGS and CAS Honduras both analyzed for silver and gold using fire assay with AAS finish, with high grade results re-assayed by fire assay with gravimetric finish. Chemex analyzed for silver and gold using fire assay with gravimetric finish and analyzed for lead and zinc using four-acid digestion with AAS. High grade ‘overlimit’ lead and zinc results were re-analyzed by volumetric methods (titration).

For Tahoe’s check assay program, Inspectorate shipped 5% of the assay pulp splits to Chemex in Vancouver, BC or Reno, Nevada for re-analysis (in 2011, Inspectorate shipped 25% of mineralized sample interval pulps to Chemex in Vancouver for check analyses). Chemex analyzed for silver and gold by fire assay and gravimetric finish and for lead and zinc using four-acid digestion with AAS. High grade ‘overlimit’ lead and zinc results are re-analyzed by titration.

11.5.4Duplicates

Duplicate samples are collected after the first stage of crushing (coarse rejects) as opposed to check-assay samples, which are sample pulps.

From May 2007 through July 2009, Entre Mares submitted coarse reject duplicates generally at the rate of one in 15 samples to CAS Honduras for analysis of gold and silver by fire assay/AAS. From July 2009 through May 2010, Entre Mares sent coarse reject splits to Chemex in Vancouver, though on a much more irregular schedule. Chemex completed the sample preparation process and analyzed the new sample pulps for silver and gold using fire assay with gravimetric finish and for lead and zinc using four-acid digestion with AAS. High grade ‘overlimit’ lead and zinc results were re-analyzed by titration. Tahoe has discontinued the use of coarse reject duplicate analyses from surface drill samples.

Duplicate samples from the underground drilling are from coning and quartering crushed whole core samples, with each half-sample analyzed by Inspectorate. The company has discontinued this practice based on MDA’s recommendation.

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11.6CONCLUSIONS

MDA believes that the core sampling procedures, sample analyses, QA/QC procedures, and sample security have provided samples that are of sufficient quality for use in the resource estimation discussed in Section 14.

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12DATA VERIFICATION

Mine Development Associates (MDA) verified the Escobal database on three occasions; in 2010 and 2012 for the 2010 and 2012 mineral resource estimates, and in 2014 for this current resource estimate. The project drill data included in the 2010, 2012, and 2014 data audits are surface exploration drill holes up through E13-443 and underground drill holes UG-12-0010 through UG-14-0389.

The verification work consisted of 1) completing an audit of the full assay database; 2) checking a significant percentage of the drill location and survey data; 3) conducting three site visits, which included verification sampling and a review of sample handling and logging procedures; and 4) reviewing the QA/QC data. The results of this verification program support the estimation of the Escobal resource and the assignment of Measured and Indicated classifications to much of the stated resource.

12.1DATABASE AUDIT

12.1.1Drill Collar Database

2010 Audit: The drill-hole collar locations for the Entre Mares drill holes were audited against the original spreadsheet data from the third-party surveyor, and no errors were found. After updating the database with the survey data for the 2010 Tahoe drilling, MDA checked all locations by plotting the drill-hole collars on cross-sections and comparing the locations with the digital topography. A number of drill holes in areas of steep topography had a ±3 m elevation difference with the topography and/or adjoining drill holes. After discussion with Tahoe, MDA adjusted the elevations on 21 drill holes to better match the existing data. The uncertainty in some of the collar elevations is not considered significant for the resource estimation or classification.

2012 and 2014 Audits: In 2012, Tahoe converted the project database to the Escobal-site Cartesian coordinate system. Using the X, Y conversion formulas provided by Tahoe, MDA in 2012 checked for consistency all of the Entre Mares and 2010 Tahoe Resources drill collar coordinates against the database. MDA then audited the post-2010 surface exploration drill holes against the original spreadsheet data from the third-party surveyor. The surveys are recorded in NAD27 UTM “ground” coordinates, typical of most survey coordinates, and then converted to the Cartesian grid system used in the mine. Only minor errors were noted and the database was revised to include all final data. All drill-hole locations were also checked by plotting the hole collars on cross-sections and comparing the locations with the digital topography.

The underground collar locations were surveyed by Escobal mine survey department personnel. MDA checked the underground drill collar database against a spreadsheet file “Pozos_Mina” provided by Tahoe. The underground collar audit resulted in the switching of collar coordinates for holes UG-12-0146 and UG-12-149. There are no surveys for 13 holes, though these holes are either well outside the modeled resource or are very recent holes in which the database does not yet have assays. In either case, the approximate locations in the database will not affect the resource estimate.

12.1.2Down-Hole Survey Database

Down-hole survey measurements for both the underground and exploration drilling are collected on approximate 50 m drill-depth intervals. A final reading is taken at the bottom of the drill hole.

Over the course of the three separate audits, MDA checked approximately 15% of the survey data against the original survey reports. For 17 of the Entre Mares holes audited, MDA compared the database values against a visual inspection of the original Sperry Sun camera discs produced by the Tropari survey instrument; only occasional minor discrepancies (<2 degrees) were noted between MDA’s reading of the discs and the current database. For the remainder of the drilling, MDA compared the database survey data against the original survey coupons created on-site

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by the survey crew. Five errors were noted, only two were significant, and the database was corrected to reflect the new data.

12.1.3Assay Database

The Escobal assay database includes results of the primary analyses for silver, gold, lead, and zinc, plus a 32-element geochem suite, completed by Inspectorate. The project database also includes the secondary check and duplicate analyses completed by ALS Chemex (Chemex). The assay database does not include any of the pre-2010 check assay data from the Marlin Mine lab or the CAS Honduras lab due to the lack of back-up data and the inability to verify any of these analyses.

For the drill assay data used in the 2010 resource estimate, each sample interval’s final “accepted” metal value was an average of multiple analyses if duplicate and/or check assay values were present. A number of different analytical techniques was employed by the labs which resulted in some uncertainty as to what values should be used in determining the accepted database metal value. After discussions with Tahoe, a hierarchy of assay techniques was established for each metal, and it was decided that the “accepted value” would be calculated using only the value(s) for the highest-ranked technique. The 2012 and 2014 databases were standardized to use only the primary assay values from Inspectorate and no further reconstruction of the assay data was required.

The sample data sent to the laboratory uses a sample ID code that is blind to the laboratory as to the drill hole number and from-to location. For each of the three resource estimates, MDA audited the assay database by downloading the assay data directly from the laboratories and compared with the database values. Minor clerical errors were noted in the 2012 database while both random and systematic errors were noted in the more recent drilling evaluated in 2014. Systematic errors included: variable treatment of “less-than-detection” values, not replacing original fire assay or ppm value with gravimetric or ore-grade re-analysis, and entering incorrect values from the original laboratory spreadsheet data. All assay database errors were corrected and the less than detection values were standardized to have a “0” value for use in the resource estimate.

Along with the assay data, MDA checked the sample ID/drill hole “from-to” correlation by comparing the hard-copy sample selection data against the database. Approximately 15 percent of the database was checked and only three minor errors were noted and corrected. The drill sample down-hole locations were also checked while on site by comparing the database from-to values directly against the sample intervals marked within the core boxes for 20 drill holes. No errors were noted.

The assay database is considered very clean and ready for use in the resource estimate.

12.2SITE VISITS

Paul Tietz of MDA visited the project site on September 7th through the 10th, 2010, again on February 6th through the 9th, 2012, and for the third time January 28th through January 31st, 2014. The purpose of the visits was to review the Escobal deposit drilling and sampling procedures, results, and geology in preparation for the resource modeling, and to complete the remaining data audit tasks required for the 2010, 2012, and 2014 resource estimates. Specific data verification items included database construction and recordation, drill-hole location and down-hole survey validation, and QA/QC methods. A limited amount of core was evaluated, and verification samples were collected in 2010 from four core holes. During each site visit, time was spent in the field verifying and discussing drilling and sampling procedures and geologic concepts with project personnel.

12.2.1Drilling Operations

Core rigs were operating on the property during all of MDA’s site visits. A detailed description of the core drilling campaigns and procedures is provided in Section 10, Drilling. The drilling procedures observed while on-site were

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consistent with industry standards, and no drilling issues were observed or discussed with project personnel which would negatively impact the resource estimate.

12.2.2Sampling and Logging Procedures

The current logging and sampling procedures meet industry standards/practices and are sufficient to allow for confidence in the resource estimate. The one improvement suggested by MDA is that the full length of the underground drill holes should be sampled due to the close proximity to the ore body. Unsampled intervals create modeling and estimation concerns especially when the sampled intervals end in mineralization.

12.2.3Drill Collar Verification

MDA used a hand-held GPS to check the locations of sixteen surface drill collars. The hand-held GPS cannot achieve survey-level accuracy, but it serves to verify that in general terms drill holes are where the database indicates they should be. MDA did not identify any discrepancies in the locations of drill holes. Due to the constructed mine facilities, many of the surface drill holes within the Central and West areas are no longer visible.

12.2.4Verification Sampling

MDA collected eight quarter-core verification samples from typical moderate- to high-grade sample intervals within four core holes in 2010. The MDA samples were delivered to the Inspectorate prep lab in Guatemala City and were analyzed using the same analytical techniques as employed for the Escobal drilling.

The verification sample results show similar mean values for silver, lead, and zinc, though individual intervals have differences of up to 80 percent. The MDA gold values are predominantly lower than the original samples, which could be a result of sampling bias or inherently more erratic gold mineralization. The MDA sample results are not considered to be statistically meaningful due to the limited sampling but serve primarily as a general verification of the Escobal metal grades.

12.3QUALITY ASSURANCE AND QUALITY CONTROL

MDA evaluated the quality control and quality assurance (QA/QC) data for the Escobal project in 2010, 2012, and 2014. In 2010, MDA evaluated the QA/QC data for holes drilled in 2007 and into 2010, up to and including number E10-225. In 2012, MDA evaluated the QA/QC data for holes drilled in late 2010 and 2011 (E10-226, up to and including E11-348 and PZ11-02). For the current 2014 resource estimate, MDA evaluated the QA/QC data for exploration holes drilled in 2012 and 2013 (E12-349, up to and including E13-444) and the 2012-early 2014 underground holes (UG-12-010, up to and including UG-14-434).

The QA/QC data and procedures are generally similar over time though there are differences in the exploration drilling and underground drilling campaigns. In all drill campaigns, Inspectorate was the primary lab used by Tahoe and the QA/QC data consists of blanks, standards reference material, and duplicate samples. A small number of underground drill samples were analyzed at the on-site lab at Escobal. The QA/QC from the batches analyzed onsite returned disappointing results, and use of that lab for drill samples has been discontinued until improvements are made.

Tahoe provided the following description of protocols used from 2010 to the present:

Assay Standards:To be inserted into sample stream prior to submission of the samples to Inspectorate at a rate of one assay standard per 20 drill samples (5%). Assay standards are selected based on which standard’s grade more closely matches the geologist’s estimate of grade for the proximal drill samples.

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Assay Blanks:To be inserted into the sample stream at irregular intervals; particularly internal or immediately following high-grade sample intervals. Inspectorate also monitored pulverizer contamination by collecting and analyzing the barren sands used to clean the pulverizer after each sample.

Check Assays:Inspectorate shipped 5% of the assay pulp splits to ALS Chemex (Vancouver, BC or Reno, Nevada) for reanalysis. Due to a mis-communication, in 2011 Inspectorate shipped approximately 25% of mineralized sample interval pulp splits to ALS Chemex for check analyses. Tahoe used ALS’s Vancouver facility exclusively in 2011.

12.3.1Assay Standards

Analyses of standards are available for silver, gold, lead and zinc data. For the 2010 and earlier drilling, there was a limited data set and no analyses has been conducted on these few standards. Four different standards were used in the 2011 surface drilling while eight standards were used in the 2012 and 2013 surface exploration drilling. All surface exploration drilling standards are commercial standards prepared by CDN Resource Laboratories of Langley, B.C., Canada (“CDN”). The underground drilling uses 6 commercial standards obtained from WCM Minerals (“WCM”) of Burnaby, British Columbia.

MDA evaluated the results obtained for the standards using a variation of the common Shewhart-type control chart. An example for silver standard CDN-ME-7 used in the 2011 exploration drill program is shown in Figure 12-1. Similar charts were prepared for each of the standards analyzed eight or more times, for each of silver, gold, lead and zinc.

Figure 12-1: Control Chart for Silver in CDN-ME-7

Notes:The red upper and lower warning lines are the Best Value ± 2 * Std. Dev., using statistics calculated by CDN.

The red upper and lower control lines, UCL and LCL, are the Best Value ± 3 * Std. Dev., using statistics calculated by CDN.

The three low-side failures noted were excluded from the calculation of the mean and standard deviation for this data set, shown by blue lines

Where Tahoe identify analytical failures, their policy is to re-run the affected sample batch for the element concerned. The final exploration data made available to MDA incorporates any such re-runs, so the failures identified by Tahoe in

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the batches that were re-run are not “visible” to MDA and are not counted as failures. Tahoe has advised MDA that they were most rigorous about identifying failures and having batches re-run for silver, somewhat less rigorous in the case of gold, and least rigorous in the cases of lead and zinc. Silver carries in excess of 85% of the value of the project, and silver also has the fewest standard analyses failures.

Any set of analyses by a single lab will in all probability have some biases relative to the accepted values of standards. Four standards had biases whose magnitudes exceed 5%, though three of the four were biased are on the low side; in other words the lab produced results that are biased low relative to the accepted values of the standards. The biases seen in the Escobal exploration data are typical of those that MDA encounters in assay data sets and none are considered significant enough to have a material effect on the resource.

The failure rates in the 2012-2013 exploration standards are higher than MDA would normally expect. In order to gain an overview of the performance of the lab, MDA constructed a “Z Score” chart for all metals in all the exploration drilling standards. The Z Score is calculated as (observed value — expected value)/(expected standard deviation). It has the effect of normalizing the results from different standards to a common base and scale, permitting them to be plotted together on a single chart. The Z score chart for silver in the 2012 and 2013 standards is shown in Figure 12-2.

Figure 12-2: Control Chart, Silver Z Score All Standards

Figure 12-2 shows that from the beginning of 2012 through to about the beginning of August, 2013, analyses of silver in the standards for the most part fell within the acceptable range, with no high failures and few low failures. Subsequently, from the beginning of August 2013 through the remainder of that year, the results became much more erratic, with some extreme failures, both high and low. Something changed in August 2013. Changes might have consisted of any or all of deterioration in field procedures, deterioration in the qualities of the standards, or some systematic change at the laboratory producing a reduction of precision in silver analyses. Z-score charts for gold, lead, and zinc all show the same increase in erratic results beginning in about August 2013.

A rather large number of failures were identified in the underground drilling (156 out of 1656 total analyses) though only eight of the failures were flagged by Tahoe as requiring re-assays. This is because only standards from mineralized batches were flagged for re-assay.

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12.3.2Sample Blanks

For MDA’s 2010 and 2012 QA/QC analyses, Tahoe describes the blank material as “collected from local outcrops of unaltered andesite, which is the one of two primary host rocks for the deposit. The samples were combined and twelve splits were sent to the lab for Au and Ag analysis only; all samples came back <5 ppb Au and <0.1 ppm Ag. However, there were no base metal analyses performed.”

In email correspondence during May of 2014, Tahoe described the material used not as “unaltered andesite”, but as “pumice tuff” of unspecified lithologic affinity.

The blank material described in 2010 and 2012 remained in use during the 2012 — 2013 period. MDA has reviewed the data from the blanks for gold, silver, lead and zinc, but concludes that the material is not suitable as a blank for lead and zinc.

MDA evaluated the results of the analyses of blank material making use of charts like the example in Figure 12-3.

Figure 12-3: Blanks in Inspectorate Gold Analyses

Figure 12-3 shows a plot of the results of the analyses of blanks, with a superimposed plot of the analyses of the samples numerically-preceding each blank. The purpose of superimposing the analyses of the preceding samples is to gain a visual impression as to whether a high grade in a preceding sample tends to produce a higher grade in the immediately-following blank. The red “warning” line on Figure 12-3 is arbitrarily set at five times the lab’s lower detection limit. Three to five times the detection limits are typical industry rules of thumb, when no more rigorously-determined rule is available. Two failures are evident in Figure 12-3. The grades in the two failures are so high that MDA suspects they are due to record-keeping errors rather than analytical failures.

There are 2,796 analyses reported to be of blank material used in the 2007 thru 2010 drilling. Of this total, 50 count as failures using five times the lower detection limit as a rule of thumb. This is a rate of 1.8%. MDA cannot determine which of the 50 are record-keeping errors and which are analytical failures. The post-E10-225 exploration data set includes 1,253 analyses of material identified as blanks. Continuing to use five times the detection limit as the failure criterion for gold and silver, the failure rate for gold in the blanks is negligible (<0.5%) while the failure rate for silver is 1.5%. Just two of the silver failures have grades high enough, 77.5 g Ag/t and 78 g Ag/t, that were they real samples, they could have a material local effect on estimated economic silver grades.

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The underground data contains 388 analyses of blank material which MDA matched to samples in the assay database used in the resource estimate. The blank material is from the same source as the material used for the exploration drilling. As noted above, the material does not usefully serve as a blank for lead and zinc. For silver, lead and zinc there is a statistically meaningful correlation between the grades in the blanks and those in the preceding samples, whereas for gold such a correlation is not evident. Batches having significant mineralization, in which an analysis of a blank was identified as a failure, were sent for re-analysis. The results of these re-analyses are discussed below in Re-Analyses of QA/QC Failures.

12.3.3Duplicate Samples

For the 2007 through 2010 drill holes up to and including E10-225, the duplicate samples consist of thirty-two combinations of pulp “check assay” duplicate samples and coarse reject “preparation” duplicate samples, in different pairings of analyses done at Inspectorate, Chemex, CAS in Honduras and at the Marlin Mine lab. A total of 12,970 individual duplicate pairs are available for gold, silver, lead, and zinc.

For exploration drilling after E10-225, the duplicate sample data set consists of 539 pulps prepared and analyzed by Inspectorate which were subsequently sent to Chemex for check analyses. A small number of same-lab check analyses were also re-run at Inspectorate in 2011.

The underground drilling duplicate samples are field “split-core” duplicates with each half-sample analyzed by Inspectorate. MDA noted that the coning and quartering process used to create the field duplicate is probably the single largest source of error in the sample preparation process, and, in MDA’s opinion, the field duplicates provide little to no information about the precision of the un-duplicated samples.

An example of MDA’s analyses of the duplicate pairs data is shown in Table 12-1.

Table 12-1: Summary Comparison of 2011 Original and Check Analyses

			Mean Values of Parameters
		Erratics	Original	Chemex	Pct Diff of	Relative Pct	Abs Rel
Metal	Count	Rejected	ppm	ppm	Means	Diff	Pct Diff
Gold	194	4	0.382	0.391	2.4	5.7	14.5
Silver	204	7	308.7	311.4	0.9	1.6	8.6
Lead	207	none	0.85	0.876	3.1	14.1	17.3
Zinc	230	none	1.063	1.115	4.9	11.1	15.5

Notes:Relative Percent Difference is calculated as 100 x This is a “worst case” calculation that gives a more extreme value than the more statistically rigorous 100 x

MDA also evaluated the silver, gold, lead and zinc check assays using scatterplots and relative percent difference plots. Examples for silver data from the 2012 and 2013 drill period are shown in Figure 12-4, Figure 12-5, and Figure 12-6.

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Figure 12-4: Scatterplot, 2012 and 2013 Silver Check Assays vs. Originals

Figure 12-5: Silver Relative Percent Difference (2012 and 2013 Check Assay vs. Original)

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Figure 12-6: Silver Absolute Relative Percent Difference (2012 and 2013 Check Assay vs. Original)

The results of MDA’s analyses indicate that for the 2007 through 2010 data, in almost all comparisons involving the Marlin Mine lab, the Mine analyses come out low. Inspectorate appears to be biased high relative to Chemex, though not significantly, and the Marlin Mine lab but may be either high or low relative to CAS, depending on the metal and method being evaluated.

The 2011 pulp check assay data (Table 12-1) contained higher percentages of erratically-large differences between data sets, about 2.1% for gold and about 3.4% for silver. On average, Chemex’ analyses are marginally higher than those of the primary lab, Inspectorate, though, at a 95% confidence level, the set of check analyses is not statistically distinguishable from the set of original analyses.

The results of the check assay analyses on the 2012-2013 data (as shown in Figure 12-4, Figure 12-5, and Figure 12-6) indicate that the Chemex silver values are lower grade than the original Inspectorate assays, especially at silver grades above 60 g Ag/t. MDA evaluated the apparent 5% low bias in the Chemex data by reviewing the two labs’ respective biases with respect to standards. For the standards at approximately 100 g Ag/t and approximately 200 g Ag/t, Inspectorate’s average analysis is quite close to the expected grades whereas Chemex’s average analysis is biased low. This may help to explain in part Chemex’s low bias relative to Inspectorate in the 60 g Ag/t to 500 g Ag/t range but there is insufficient data to compare the two labs’ performance on higher-grade silver standards.

12.3.4Re-Analyses of QA/QC Failures

In early 2014, Tahoe sent 309 underground drill samples from batches containing failed blanks or standards, along with appropriate QA/QC control samples, to Inspectorate for re-analysis. Tahoe provided the results of the re-analyses to MDA in May of 2014.

The database that MDA used for the resource estimation described in this report contains the original analyses, not the re-analyses. To assess the consequences of continuing to use the original analyses, MDA compared the silver re-analyses to the silver originals in the MDA database, to determine whether the use of the older or the newer analyses would have made any material difference in the outcome of the resource estimate where it is influenced by affected samples.

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For the evaluation, MDA used charts similar to the examples provided in Figure 12-4 through Figure 12-6. MDA concludes that in grade ranges material to the resource estimate (exceeding 30 g Ag/t), the differences between the original assays and the re-assays are not material.

12.4CONCLUSIONS AND RECOMMENDATIONS

The number of failures and biases seen in the Escobal exploration standards are typical of those that MDA encounters in assay data sets and are not considered significant. The exception is the high number of failures observed in the period from August 2013 through the remainder of that year. It is not known the reason for the high failure rate. The standards used in the underground drilling also had a high number of failures (9% of the total standards). Tahoe did re-assay the sample batches for those “failure” standards associated with mineralized drill intervals.

The analyses of the exploration blanks yielded a failure rate for silver of about 1.5% while the failure rate for gold is <1%. Only two silver failures are significant and the overall failure rate for silver is not so high as to disqualify the data set for use in the resource estimate described in this report. There is apparent evidence for some contamination in blanks in the underground data, though the degree is limited and MDA does not believe that this has a significant effect on the resource estimate.

MDA has the following comments respecting the exploration QA/QC data set:

1.It is not clear to MDA what actions, if any, were taken in response to failures in the standards or blanks. MDA suggests that these be clearly documented.

2.MDA is not aware of any field, preparation or pulp duplicate samples in the exploration data set. These types of duplicates would be useful tools for evaluating the precision of the primary lab’s analyses and MDA recommends their use.

3.The combination of standards, blanks and external check assays is adequate to support the use of the assay data set in the resource estimate.

The following comments pertain to the underground drilling QA/QC data set:

4.Failures in the standards or blanks in batches of samples containing significant mineralization, were followed up with re-analyses,

5.Field duplicate samples were collected and analyzed, but the method used to split the samples in the field to create the duplicates significantly reduces their usefulness. MDA suggests that field duplicates made by cutting the core lengthwise, as is common in the industry, would be more useful. MDA further suggests that preparation and pulp duplicates be used.

6.No check assays were obtained from a second, external lab. MDA recommends the use of such checks.

As a general comment, there appears to be an unnecessary disconnect between the QA/QC procedures used with the exploration samples and those used with the underground samples.

7.The exploration group makes use of external check samples; the underground group does not,

8.The underground group collects and analyzes field duplicates; the exploration group does not,

9.The two groups use different sets of standards from different suppliers. This is not necessarily a problem, and the use of numerous different standards in a sense provides some additional assurance. Nevertheless, MDA suggests that the use of at least some common standards would add some continuity to the QA/QC data.

While there are some deficiencies in the QA/QC data for Escobal, in general Tahoe’s drill procedures, sampling programs, data collection and management, and QA/QC programs are consistent with industry standards, have

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become more thorough over time, and have yielded results that support the use of the project’s assay database for the resource estimate described in this report.

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13MINERAL PROCESSING AND METALLURGICAL TESTING

Independent and well-respected international testing facilities that have performed metallurgical test work on the Project include:

·FLSmith Dawson Metallurgical Laboratories; Salt Lake City, Utah, USA.

·McClelland Laboratories (McClelland); Nevada, USA

·SGS Lakefield; Canada

·Economic Geology Consulting (EGC); Nevada, USA

·Phillips Enterprises, LLC (PE); Colorado, USA.

·Silver Valley Laboratories (SVL); Idaho, USA.

·Kappes Cassiday Associates, Reno, Nevada, USA.

·Hazen Research, Golden, Colorado, USA

The laboratories do not hold ISO certification for metallurgical testing activities; this is typical for metallurgical test work facilities. Test work was performed on behalf of Entre Mares during 2008—2009.

Previous metallurgical test work conducted by McClelland Laboratories (McClelland), Utah, USA and Kappes Cassiday Associates, Reno, Nevada, USA (KCA) concluded that differential lead/zinc flotation producing a high value lead concentrate containing most of the silver and gold in the mill feed and a saleable lower value zinc concentrate was the optimum processing route.

The following conclusions were drawn from the test work conducted by McClelland Laboratories in May 2009:

·The Escobal sulfide and mixed oxide/sulfide composites did not respond particularly well to gravity concentration treatment, at an 80%-106mm feed size.

·The Escobal sulfide composites responded well to conventional bulk sulfide flotation treatment for recovery of gold and silver, at an 80%-75mm feed size

·The Escobal sulfide composites showed good potential for selective flotation of contained lead and zinc.

·The Escobal mixed oxide/sulfide composite did not respond as well to conventional bulk sulfide flotation treatment.

·The Escobal composites were moderately amenable to whole ore milling/cyanidation treatment, at an 80%-75mm feed size.

·The EC08-127 composite may have displayed a moderate preg-robbing tendency during whole ore cyanidation.

·Adding activated carbon during whole ore cyanidation (CIL) leaching generally was effective in significantly improving gold and silver recoveries.

·Cyanidation of flotation products, including regrind/intensive cyanidation of flotation rougher concentrates, was not particularly effective in increasing overall leach recoveries, when compared to whole ore CIL/cyanidation leaching.

FLSmidth Dawson Metallurgical Laboratories was selected in June 2010 to conduct a comprehensive metallurgical test program on a composite sample representative of the Escobal deposit. The objective of the Dawson Metallurgical testwork was to advance the design of the differential flotation circuit to process the Escobal ores. The sequence of flotation testwork conducted by Dawson Metallurgical included the following:

·Grind time determination

·Reagent screening for lead rougher flotation

·Lead rougher flotation

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·Reagent screening for zinc rougher flotation

·Zinc rougher flotation

·Preliminary grind size optimization

·Locked cycle testing

·Tailing and concentrate physical property characterization

Hazen Research, Inc. located in Golden Colorado was selected in June 2010 to conduct a comprehensive comminution test program using PQ core samples drilled specifically for these tests. The tests include: JK drop weight tests, Bond rod mill tests, Bond ball mill tests, Bond crushing tests and abrasion testing.

13.1SAMPLING

A total of 46 buckets of drill core samples were received for sample preparation and assay at Phillips Enterprises and stage crushed to minus 10-mesh prior to flotation testing.

Head assays for the master composite were conducted to determine gold, silver, lead, zinc, copper, iron, and antimony. Also carbon (total/organic), non-sulfide lead and non-sulfide zinc were determined. The results of the head assay analysis for the master composite are presented in Table 13-1.

Table 13-1: Master Composite Head Assay Results

Au, ppm	Ag, ppm	%Pb	%Zn	%Cu	%Fe	%Sb	%Ctot	%Corg	Pbns	Znns
0.412	569	0.984	1.56	0.041	2.48	0.044	1.35	0.05	0.10	0.097

13.2GRINDING TESTS

Two kilogram samples were ground in a mill at 20, 30, 40, 50, and 60 minute intervals to establish the relationship between the grind sizes (P80) and grind times. The time required to achieve various grinds were obtained and tests were run at different grind sizes to ascertain the relationship between P80 versus metal recovery.

The test results shown in Table 13-2 indicate that grinding beyond 105 microns did not result in any significant increase in metal recoveries. It was therefore decided that 105 microns was the optimum grind size for rougher flotation. This grind was used for further testwork as well as the design of the process plant.

Table 13-2: P80 Versus Metal Recovery to the Lead Rougher Concentrate

Test No.	Grind P80 microns	Metal Recovery to Pb Rougher Concentrate with SIPX as Collector
		%Au	%Ag	%Pb	%Zn	%Cu	%Fe	%Sb
9	231	55.3%	69.5%	83.6%	22.2%	57.7%	16.6%	28.0%
10	144	61.6%	75.5%	88.2%	20.8%	64.3%	18.3%	30.3%
11	105	67.0%	78.7%	88.7%	19.1%	64.0%	18.9%	30.8%
4	74	64.9%	82.9%	89.4%	17.9%	64.8%	22.8%	30.8%
12	46	67.9%	79.6%	82.6%	14.5%	64.6%	13.1%	30.0%
13	37	68.5%	76.5%	67.5%	11.4%	56.4%	10.3%	28.0%

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Test No.	Grind Time (P80,microns)	Metal Recovery to Pb Rougher Concentrate With PAX as Collector
		%Au	%Ag	%Pb	%Zn	%Cu	%Fe	%Sb
22	35(150)	67.6%	81.1%	88.9%	31.9%	71.7%	34.7%	31.4%
23	45(105)	71.8%	81.3%	89.2%	25.7%	70.6%	37.2%	30.8%
24	60(75)	73.9%	81.2%	89.4%	24.1%	71.8%	36.4%	32.3%
25	85(53)	73.9%	80.0%	80.1%	15.1%	64.9%	21.2%	37.9%

13.3GRINDABILITY TESTS

Phillips Enterprises, LLC conducted three ball mill grindability tests on Escobal Project samples as part of the McClelland Laboratories test work. The resulting ball mill work indices (Wi) from ball mill grindability tests conducted at closing screen of 100 mesh (150 micron) are shown in Table 13-3.

Table 13-3: Resulting Ball Mill Work Indices from Ball Mill Grindability Tests

Sample	Wi (kW-hr/st)	Wi (kW-hr/mt)
EC08 - 122	14.07	15.55
EC08 - 125	17.22	18.99
EC08 - 127	16.44	18.13
Average	15.91	17.56

A circuit consisting of one 5 m by 8.5 m ball mill in closed circuit with a hydrocyclone classifier was selected as a circuit that would likely meet the design tonnage. This circuit was based on the results from the comminution tests to produce a primary grind size of 80% passing 105 µm.

13.4REAGENT SCREENING TESTS

Initial lead rougher flotation tests were conducted with Sodium Isopropyl Xanthate and Sodium Ethyl Xanthate and 3418A as the main collectors. It was observed that each collector tried worked well with the flotation being very fast and essentially being completed after 4 minutes. Microscopic examination of the concentrates indicated that concentrate contained galena as the main product with pyrite and gangue as the main contaminants with lesser but significant amounts of sphalerite as the third most common contaminant. Examination of the tails showed that the predominant sulfide minerals were pyrite and sphalerite with pyrite being in the majority. The minerals were very liberated with the only locking seen being small blebs of pyrite attached to gangue. It was also found that Sodium Isopropyl Xanthate (SIPX) performed better than Sodium Ethyl Xanthate. More reagent screening tests were conducted with the stronger xanthate, Potassium Amyl Xanthate (PAX), which improved the recovery of precious metals when compared with Sodium Isopropyl Xanthate as shown in the Table 13-4.

Table 13-4: Typical Metal Recovery to Lead Rougher Concentrate

Collector Type	%Au	%Ag	%Pb	%Zn	%Fe
SPIX	74.5	79.7	90.9	26.0	30.1
PAX	77.3	83.4	91.8	30.4	37.5
Difference	2.73	3.68	0.91	4.36	7.40

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With the objective to maximize the precious metals recovery to the lead concentrate, Potassium Amyl Xanthate (PAX) was used in subsequent tests as the main lead rougher collector.

The tests showed that galena floated well with xanthates as the main collectors. Pyrite and sphalerite floated with galena but not in unusual quantities considering the use of very strong collectors to maximize precious metals recoveries and the fact that pyrite is the most abundant sulfide in the material tested. The high degree of liberation indicated that a grind coarser than 75-microns is possible and that the rougher concentrate will clean well.

Initial zinc reagent screening tests were conducted with the tailings from the lead rougher flotation. Lime (Ca(OH)2), copper sulfate (CuSO4.5H2O) and Potassium Amyl Xanthate (PAX) were added to the lead rougher tails slurry, conditioned for 5 minutes and a zinc rougher flotation step was completed after adding X-133 frother. The pH of the lead rougher tails slurry was raised to 9.5 with lime to depress pyrite and the copper sulfate was used to activate sphalerite.

Sphalerite floated well with PAX as the main collector at pH 9.5 and with about 30 g/t copper sulfate dosage for sphalerite activation. The results shown in the Table 13-5 indicate good recoveries of both lead and zinc are achievable in the rougher concentrates.

Table 13-5: Typical Metal Distribution in Rougher Flotation Products

Product Identification	%Au	%Ag	%Pb	%Zn	%Fe
Lead Ro Con	76.72	82.29	91.10	31.7	29.83
Zinc Ro Con	5.56	4.91	2.94	61.89	13.39
Pb +Zn Con	82.28	87.20	94.03	93.64	43.22
Zinc Ro. Tail	17.72	12.80	5.97	6.36	56.78

More tests were conducted in September 2010 to address the following:

·Recycle water. The Escobal mine was designed to recycle as much water as possible minimizing treatment and discharge. Calcium in the lime used to raise the pH in zinc flotation depresses galena and precious metals and the copper sulfate used to activate sphalerite will increase the amount of sphalerite and pyrite that float to the lead rougher flotation concentrate if process water from the zinc flotation is recycled.

·Flotation Objectives. The best economic benefit for the project was to maximize the recovery of precious metals to the lead rougher circuit where the highest value can be achieved. Flow sheet design therefore focused on recovery of precious metals to lead concentrate and achieving a marketable lead concentrate grade at the expense of some zinc recovery.

The flotation test program addressed the above objectives, including:

·The use of co-collectors to improve the recovery of precious metals to the lead rougher concentrate

·The use of rougher concentrate regrind to improve both the lead and zinc concentrate cleaner flotation response

·The reduction or elimination of lime and or copper sulfate in the zinc rougher flotation circuit

·The investigation of process water treatment options to remove lime and copper sulfate from recycle water.

Thirteen tests were run with co-collectors, 3418A, AF31, Aero 3477, AF 208, Flomin C-4920, Flomin C-4132, Flomin C-4150, Flomin C-7436, Flomin C-4930, Flomin C-7931. The co-collectors were either used with PAX in the lead rougher flotation or used in place of PAX in the zinc rougher flotation. The -10 mesh samples used for the tests were

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all ground to a P80 of 105 microns and X-133 was used as the frother for all the tests. Lime was not added to the zinc flotation circuit.

Table 13-6: Typical Metal Distribution in Rougher Flotation Products

Test #	Product Identification	Co-Collectors Used	%Au	%Ag	%Pb	%Zn	%Fe
42/43	Lead Ro Con	AF31/C-4920	74.39	89.11	94.10	38.21	39.36
42/43	Zinc Ro Con	3418A	15.39	2.85	1.79	56.89	1.88
44/45	Lead Ro Con	3477	78.28	88.00	93.24	34.82	37.20
44/45	Zinc Ro Con	C-7436	10.42	3.90	2.33	60.49	6.86
47	Lead Ro Con	A-208	81.80	88.76	93.51	36.55	38.27
47	Zinc Ro Con	SEX/ C-4132	4.74	2.75	1.65	56.96	4.84
48/49	Lead Ro Con	C-4150	81.44	87.88	93.27	38.10	39.48
48/49	Zinc Ro Con	SIPX/C-4132	6.69	2.83	1.80	56.40	3.59
50/51	Lead Ro Con	C-7436	79.99	86.86	92.68	33.80	38.99
50/51	Zinc Ro Con	SEX/C4150	5.71	3.86	2.35	60.28	4.82
52	Lead Ro Con	C4930	80.98	87.17	93.20	37.04	39.96
52	Zinc Ro Con	PAX/H2SO4	4.34	4.73	3.01	60.35	5.97
53/54	Lead Ro Con	C-7931	84.35	87.17	93.89	37.14	36.64
53/54	Zinc Ro Con	C-7931	2.42	3.26	1.94	57.22	5.30

The results of the tests showed that additional silver (and gold) could be recovered to the lead rougher concentrate by using co-collectors. The best results for silver recovery were obtained with a combination of AF31/Flomin C-4920 and Aerofloat 208 as co-collectors. The tests gave silver recoveries of 89.11% and 88.76% to the lead rougher concentrates. A closer examination shows that Aerofloat 208 may be a better co-collector since it floated more gold (81.80% vs 74.39%) and less zinc (36.55% vs 38.21%) and iron (38.27% vs 39.36). than the AF31/Flomin C-4920 combination. Unfortunately the co-collectors also improved the flotation of the main contaminants sphalerite and pyrite to the lead rougher concentrate. More tests and mineralogical studies were needed to ascertain whether there is some association of silver with sphalerite and pyrite. Testing analyzed the alternative of regrinding the rougher concentrate to improve mineral liberation, the use of sphalerite, and pyrite depressants to improve both the recovery of precious metals to and the grade of the final lead concentrate.

The results of the tests using very selective zinc collectors and co-collectors produced good results with less than 6% zinc left in the zinc (final) rougher tails in all the tests. The best result was achieved with SEX/Flomin C-4150 co-collector combination where 60.28% of the zinc in the ore reported to the zinc rougher concentrate with only 2.35% lead and 4.82% iron reporting in the zinc rougher concentrate. The 3418A co-collector had the lowest amount of contaminants of 1.79% lead and 1.88% iron but had only 56.89% of zinc reporting to the zinc rougher concentrate.

The following conclusions were drawn from the test work conducted by Dawson Metallurgical:

·The Escobal sulfide ore is amenable to selective flotation producing a lead concentrate with most of the silver and gold in the lead concentrate and a clean zinc concentrate with some precious metals content.

·Grinding the ore to 80 percent passing 105 microns produced mineral liberation suitable for the flotation process.

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·Ore floated well with normal flotation reagents such as; potassium amyl xanthate (PAX), sodium isopropyl xanthate (SIPX), copper sulfate (CuSO4), zinc sulfate, Aerofloat 208 and Aerofroth X-133.

·Very selective collectors and co-collectors can be used in the zinc circuit at lower pH to eliminate the use of lime.

13.5DETERMINATION OF RECOVERIES AND REAGENT CONSUMPTIONS

The following recoveries and reagent consumptions were used in the process design based on test work and determinations described in the preceding sections:

Table 13-7: Escobal Concentrator Operational Parameters

Parameter	Value	Units
Silver Recovery	86.7	Percent
Gold Recovery	75.1	Percent
Lead Recovery	82.5	Percent
Zinc Recovery	82.6	percent
Bond Work Index	17.56	kW-hr/Mt
Primary Grind Size (P80)	105	Microns

Table 13-8: Reagent Consumptions

Parameter	Value	Units
Collectors	60	g/t
Activators	30	g/t
Depressants	80	g/t
Frothers	30	g/t
Flocculant	60	g/t

13.6ESTIMATED METALLURGICAL RECOVERIES FOR PROCESS DESIGN

Process development to determine concentrator unit operations and to set the design criteria for the unit operations were done by McClelland Laboratories Inc. of Reno, Nevada and FLSmidth Dawson Metallurgical Laboratories of Salt lake City, Utah. M3 reviewed the data supplied by McClelland Laboratories Inc. and Dawson Metallurgical Laboratories and relied on it to develop the process design criteria used for the design of the process facilities. The metallurgical testing program followed industry accepted practices and was believed to be technically sound and representative for the deposit. M3 recognized that the preliminary design criteria could change as more computer simulation, laboratory, or pilot plant performance testing becomes available.

McClelland’s and Dawson Metallurgical’s froth flotation test data from samples of the Escobal sulfide resources showed 75.1% gold recovery, 86.7% silver recovery, 82.5% lead recovery and 82.6% zinc recovery from feed grades averaging 415 g/t silver, 0.47 g/t gold, 0.72% lead and 1.23% zinc.

McClelland’s tests also showed that flotation recoveries of 66% and 84% for gold and silver were achievable from a mixed oxide/sulfide sample that had only about half the amount of sulfide sulfur contained in the sulfide samples. Based on experience at comparable deposits it was reasonable to assume that similar recoveries can be achieved for oxide material blended with sulfide material. Escobal metallurgical and process plant personnel are currently conducting testwork on mixed oxide/sulfide material.

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13.7POST-PLANT DESIGN METALLURGICAL TESTS

13.7.1Pilot Plant

In 2013, Tahoe commissioned a metallurgical pilot plant to validate the flotation process selected for the Escobal project, evaluate reagents type and usage, and to produce samples of lead and zinc concentrates for marketing to smelters.

Approximately ten tonnes of ore was collected from the Escobal vein in the underground workings for the pilot plant test work. Overall feed grade to the pilot plant was 463 silver grams per tonne, 0.16 gold grams per tonne, 0.59 percent lead, 0.91 percent zinc and 2.24 percent iron. Pilot plant throughput of 100-kg per hour produced about 75 kg of lead concentrate and 170 kg of zinc concentrate; a portion of which underwent additional cleaning to create the final concentrate products. Pilot plant work and concentrate cleaning were conducted at Hazen Research, Inc. of Golden, Colorado under the direction of Mr. Gregory Sharp, independent consulting metallurgist.

Results of the pilot plant are summarized in Tables 13-9 and 13-10.

Table 13-9: Overall Pilot Plant Results

	Concentrate Grade					Metal Recovery
Product	Au g/t	Ag g/t	%Pb	%Zn	%Fe	%Au	%Ag	%Pb	%Zn	%Fe
Lead Concentrate	5.27	26,080	31.3	18.3	11.7	45.7	76.0	73.5	28.5	7.2
Zinc Concentrate	1.23	25.65	3.6	31.2	15.7	15.8	11.9	13.3	62.3	12.2
Total						61.6	87.8	86.7	90.8	19.4

Table 13-10: Final Cleaned Lead and Zinc Concentrate Grades

Product	Concentrate Grade
	Au g/t	Ag g/t	%Pb	%Zn	%Fe
Lead Concentrate	10.9	69,400	64.4	2.9	2.5
Zinc Concentrate	0.85	2,290	4.8	45.1	9.0

Total metal recovery for silver and lead from the pilot plant was slightly higher than the recoveries predicted by prior metallurgical test work. The silver grade of the final cleaned lead concentrate was nearly double of that observed in the 2010 flotation tests, suggesting higher silver grades in the lead concentrates produced at the mine may be achievable, though likely not of the magnitude seen from the pilot plant results. Lead and zinc grades of the respective pilot plant concentrates also exceeded predicted grades by several percent.

Total metal recovery for gold and zinc was less than that indicated by the previous test work, though the recoveries were better than expected considering the low gold feed grade and, to a lesser extent, the low zinc feed grade to the pilot plant relative to previous flotation test head grades.

Overall, the pilot plant validated the results of the prior flotation tests and confirmed the process design selected for the Escobal mine would produce highly marketable concentrates.

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13.7.2Flotation Tests

Additional flotation tests were conducted in 2014 on drill samples collected from new areas in the East Extension Zone, deep Central zone, and West/Margarito Zone. Although mineralization in each of these areas appeared to be same mineralogically as samples previously tested from the Escobal deposit, rougher flotation tests were completed as a check of metallurgical compatibility with the Escobal process plant design. Sample composites across the entire Escobal vein were collected from seven drill holes and submitted for testing. FLSmidth conducted the flotation tests. Flotation test results are summarized in Table 13-11.

Table 13-11: Metal Distribution in Rougher Flotation Products (2014)

Deep Central (3 Composites)

Product	%Au	%Ag	%Pb	%Zn	%Fe
Lead Ro Con	59.70	71.34	84.73	34.97	17.34
Zinc Ro Con	14.13	8.76	4.74	53.12	10.58
Pb +Zn Con	73.83	80.09	89.47	88.09	27.92
Zinc Ro Tail	26.17	19.90	10.54	11.91	72.08

East Extension (2 Composites)

Product	%Au	%Ag	%Pb	%Zn	%Fe
Lead Ro Con	26.93	72.89	68.38	22.87	11.89
Zinc Ro Con	12.11	11.92	13.88	54.92	15.50
Pb +Zn Con	39.03	84.81	82.26	77.79	27.39
Zinc Ro Tail	60.97	15.19	17.75	22.22	72.61

West/Margarito (2 Composites)

Product	%Au	%Ag	%Pb	%Zn	%Fe
Lead Ro Con	60.86	70.27	93.83	25.22	33.69
Zinc Ro Con	22.37	15.81	2.40	68.65	13.69
Pb +Zn Con	83.23	86.08	93.23	93.87	47.38
Zinc Ro Tail	16.78	13.93	3.77	6.14	52.63

Results from the rougher flotation tests demonstrate that metal recovery from ore in each of the three of new areas is similar to previous flotation test results and will not require alterations to the process design.

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14                   MINERAL RESOURCE ESTIMATES

14.1                INTRODUCTION

The mineral resource estimate described in this technical report is an update of a previous mineral resource estimate completed by MDA and publically reported on May 7, 2012. Mineral resource estimation for the Escobal project follows the guidelines of Canadian National Instrument 43-101 (“NI 43-101”). The modeling and estimation of silver, gold, lead, and zinc resources were done under the supervision of Paul G. Tietz, a qualified person under NI 43-101 with respect to mineral resource estimation. Mr. Tietz is independent of Tahoe Resources by the definitions and criteria set forth in NI 43-101; there is no affiliation between Mr. Tietz and Tahoe Resources except that of an independent consultant/client relationship.

The mineral resource described in this report has an effective date for data input of January 23, 2014 and includes the data and analyses resulting from Tahoe’s 2012, 2013, and early 2014 work program up to and including exploration drill hole E13-443 and underground definition drill hole UG-14-0389. For the current resource estimate, MDA audited the data derived from drilling through January 20, 2014 , analyzed QA/QC data, conducted a site visit, and collected samples of drill core for verification purposes. All of these subjects are discussed in Sections 11 and 12 of the technical report.

The Escobal deposit was modeled and estimated by evaluating the drill data statistically, interpreting mineral domains on cross sections and then level plans, analyzing the modeled mineralization statistically to establish estimation parameters, and estimating silver, lead, gold, and zinc grades into a three-dimensional block model. All modeling of the Escobal resources was performed using GEOVIA SurpacTM software.

All of the procedures and methods used to model and estimate the Escobal deposit are similar to those used by MDA in the previous 2010 and 2012 resource estimate. Specific data and results have been updated to reflect the current work.

Although MDA is not an expert with respect to any of the following factors, MDA is not aware of any unusual environmental, permitting, legal, title, taxation, socio-economic, marketing, or political factors that may materially affect the Escobal mineral resources as of the date of this report.

14.2                DATA

The Escobal deposit mineral resource reported in this technical report is based on project drill database consisting of 842 drill holes totaling 231,326 m. The large majority of the drilling has been by diamond core drilling methods with the database containing 445 surface diamond core holes for 184,361 m and 387 underground diamond core holes for 43,869 m. The remaining drilling was by reverse circulation (RC) methods (four drill holes for 674 m) or a combination of RC and diamond core (six drill holes for 2,422 m). The Escobal drill-hole assay database contains 58,610 silver assays, 58,609 gold assays, 58,515 lead assays, and 58,516 zinc assays. The geology database includes drill-hole lithology, alteration, vein type and percentages, sulfide content, and oxidation state data. Digital topography at 2-m contours was supplied by Tahoe Resources.

14.3                DEPOSIT GEOLOGY PERTINENT TO RESOURCE MODELING

Mineralization within the Escobal deposit is characterized by multi-phase brecciation and quartz±sulfide veining emplaced within a generally east-west trending structural zone. The structural zone is divided into Central and East zones which represent structurally off-set portions of the same mineral system. Similar vein orientations and relationships are observed in each zone though the south-dipping high-level vein seen in the East zone has not been as well preserved in the topographically lower Central zone.

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Mineralization within the Central zone (the deposit west of ~807200E) extends 1,500 m along a general east-west strike and has total vertical extent of 1,000 m. (The western portion of the Central zone is often referred to as the West zone, or Margarito zone, by project personnel. For the purposes of the resource model, the Central zone refers to the combined West and Central zones.) The upper portion of the Central zone, interpreted to be a dilational jog between south-dipping normal faults, primarily dips steeply to the north at approximately 60-70 degrees and has a vertical extent of approximately 300 m. This portion of the deposit is the focus of the current underground development and contains the greatest thickness of relatively continuous mineralization. At depth, the structural zone becomes more vertical and mineralization generally weakens though the deposit appears to have a shallow westerly rake and high-grade drill intercepts are encountered at depth at the western end of the Central zone (the Margarito zone). Within its eastern half, the upper portion of the Central structure is truncated against a weakly to moderately mineralized, east-dipping structure whose orientation is sub-parallel to the main East zone structure. Structural interpretations suggest that this Central zone east-dipping structure is the structurally-offset western extension of the East zone structure.

A late, generally weakly mineralized, quartz-calcite vein event occurs within the middle and eastern portions of the Central zone mineralized structure. The late veining occurs as distinct, predominantly post-mineralization veining, often less than 1 m thick, that cuts through intervals of higher-grade mineralization. The late veining is prevalent in almost all intercepts to some degree, with the percentage of late vein directly affecting the assay grade of the individual sample intervals. Where the late quartz-calcite vein attains an appreciable thickness (up to 10 m), a “metal void” is created in the otherwise generally continuous high-grade mineralization. Isolated instances of higher-grade mineralization within the vein often are associated with the presence of sulfide-rich, clasts or remnant slivers of the mineralized wallrock.

East zone mineralization extends for 850 m along a general N80E strike. The upper portion of the East zone (the East vein) dips to the south at approximately 60-70 degrees and has a vertical extent of 450 m. The south-dipping vein transitions to a steeply north-dipping structural zone similar in style to the productive portion of the Central zone system but the vertical extent (200 m) is not as great and mineral continuity and thickness are not as well-developed as in the Central zone. Below the north-dipping zone are near-vertical to steep south-dipping structures which mimic the deeper portions of the Central zone. Extensions of these near-vertical structures up into the East vein hanging wall are marked by narrow, though often high-grade, quartz-sulfide veins. Total vertical extent of the East zone is 950 m though the deeper portions are not well defined due to the limited drilling.

The tuff lithology which overlies the andesite in the East zone is a post-mineralization unit, and the East zone mineralization truncates against the base of the tuff.

The mineralized structure(s) can be up to 50 m wide though widths of 10-30 m are more common. The structure/wallrock boundaries are often not distinct sharp contacts, but consist of a gradual decrease in brecciation/veining over a 5-10 m distance. This gradation is especially common within the hanging wall andesite wallrock in the Central and East zones. Conversely, the footwall boundary is often more clearly defined within the sedimentary rocks in the deeper sections of the Central zone. Peripheral to the main mineralized structures, mineralization occurs within thin (<0.5 m) sulfide-bearing quartz veins and breccias.

The Central and East zones contain predominantly sulfide mineralization, with minor oxide and mixed oxide/sulfide material within the upper portions of both zones. Silver, lead, and zinc occur throughout the Central and East zones, with better grades generally occurring within the sulfide mineralization. Gold also occurs throughout the deposit, though the richest gold mineralization is within the oxide and mixed material within the upper levels of the East zone.

14.4                GEOLOGIC MODEL

A cross-sectional geologic model of the Escobal deposit was created by Tahoe and MDA. The cross-sections looked due east and are numbered using the project’s UTM Easting coordinates with the westernmost section being section

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805250E and the eastern section at 808400E. The sections are evenly spaced on 50 m intervals for those portions of the deposit to the east and west of the underground development (from 805250E to 806200E and then again from 807200E to 808400E). Within the 1,000 m length of current underground development (from 806200E to 807200E), the sections are spaced at 20-m intervals that generally match the underground drill spacing.

The geologic model constructed for Escobal data included 1) the wallrock lithologies, with all apparent structural offsets, 2) the oxidation boundaries showing oxide, mixed oxide and sulfide, and sulfide material, and 3) the mineralized structures within the Central and East zones. The latter structural zone model was used as a template to guide the mineral-domain modeling (discussed below).

Tahoe’s exploration group provided MDA the geologic model on the 50-m-spaced sectional intervals while the mine geology group provided the 20-m-spaced geologic model. MDA combined the two models making sure to correlate comparable geologic features and, if needed in localized areas, bringing consistency to differing interpretations and/or spatial locations. To aid in the modeling review, MDA used the core photos of almost all of the drill holes to evaluate structural information, especially the angle to core axis orientation of the mineralized veins and breccias. The core photo review also allowed for greater definition on the vein types and structural zone contacts.

Included in the Central zone structural/vein geology was the modeling of the distinct through-going, late quartz-calcite vein(s). The location and thickness of the late vein was determined by those drill intercepts which are dominantly composed (~ >50 percent) of massive, late quartz-calcite veining. As discussed in the previous section, the late veining can occur throughout the full width of the mineralized structure zone as less than 1-m-thick veins. Multiple late veins that are highly variable in thickness and continuity are modeled within the underground development but away from the closely-spaced drilling, the late vein is modeled as a fairly regular single vein. As such, the current model outside of the underground development is simplistic in its representation of this weakly mineralized veining.

The inability to accurately estimate the weakly mineralized, late-stage veining, and its effect on the metal-grade distribution, is an uncertainty in the current resource model.

The lithology and oxidation models were converted into 3-dimensional solids which were used to code the block model. The lithology and oxidation codes were used to assign density values to the block model (see Section 14.6 for details on the block model density), while the oxidation coding was also used for resource classification.

14.5                MINERAL-DOMAIN GRADE MODELS

Cross-sectional mineral-domain models for each of the four metals were created for the Central and East zones. Distribution plots of silver, gold, lead, and zinc grades were made to help define the natural populations of metal grades to be modeled on the cross sections. The natural populations from the distribution plots were checked against the drill data and geologic model to determine if the populations represented realistic, continuous mineral types. Low-grade, moderate-grade, and high-grade mineral domains (domain codes 100, 200, and 300, respectively, in the block model) were determined for all four metals.

The resulting grade populations used to create the mineral domains are shown in Table 14-1.

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Table 14-1: Mineral Domain Grade Populations

Metal	Zone	Low-Grade Domain (domain 100)	Mid-Grade Domain (domain 200)	High-Grade Domain (domain 300)
Silver	Central	~ 5 – 150g Ag/t	~ 150 – 1500g Ag/t	~ >1500g Ag/t
	East	~ 5 – 150g Ag/t	~ 150 – 800g Ag/t	~ >800g Ag/t
Gold	Central	~ 0.06 – 0.5g Au/t	~ 0.5 – 3.0g Au/t	~ >3.0g Au/t
	East	~ 0.06 – 0.37g Au/t	~ 0.37 – 1.0g Au/t	~ >1.0g Au/t
Lead	Central	~ 0.007 – 0.2% Pb	~ 0.2 – 1.0% Pb	~ >1.0% Pb
	East	~ 0.007 – 0.08% Pb	~ 0.08 – 0.5% Pb	~ >0.5% Pb
Zinc	Central	~ 0.017 – 0.2% Zn	~ 0.2 – 1.5% Zn	~ >1.5% Zn
	East	~ 0.017 – 0.3% Zn	~ 0.3 – 0.85% Zn	~ >0.85% Zn

The mineral domains as modeled and drawn on the cross sections are not strict “grade shells” but are created using geologic information for defining orientation, geometry, continuity, and contacts in conjunction with the grades. Each of these domains represents a distinct style of mineralization. While all metals are generally spatially related, they are not always exactly coincident, thereby requiring separate domain models for each metal.

The unique metal-grade and geologic characteristics of the late, quartz-calcite vein required the creation of a unique mineral domain for it within each of the four metals (domain code 110). The late vein mineral domain included all late vein assay values and restricted estimation to within the late vein.

At the start of the mineral-domain modeling, it was realized that the low-grade gold domain was smaller in cross-sectional area than the low-grade silver domain. To assure that some level of gold mineralization would be estimated into all blocks containing silver, a dilutional domain (domain code 10) was added to the gold mineral-domain model.

Each metal was modeled independently, though the completed silver sectional model was used to help guide the general trends of the gold, lead, and zinc domain models. For the current resource estimate, a spatial and statistical analysis indicated a close relationship between the lead and zinc low- and mid-grade sectional domains. Accordingly, the low- and mid-grade zinc sectional domains were used as a proxy for the corresponding lead sectional domains. A unique high-grade lead sectional domain was created due to the increased variation from the zinc high-grade domain.

The mineral domain cross sections were three-dimensionally rectified to the drill data and then sliced at 5 m intervals that coincide with the center of the block-model’s vertical block size. The sectional slices were used to create 5 m level plans that were used to code domain percentages into the block model.

Typical cross sections through the Central zone silver domain model are shown in Figure 14-1 and Figure 14-2, while the East zone silver domain model is shown in Figure 14-3. Also included on the cross-section figures are estimation areas used to define orientations for estimation search ellipsoids. See Sections 14.7 and 14.8 for further details on the estimation areas.

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Figure 14-1: Section 806300 - Escobal Central Zone Silver Geologic Model

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Figure 14-2: Section 806800 - Escobal Central Zone Silver Geologic Model

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Figure 14-3: Section 807500 - Escobal East Zone Silver Geologic Model

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14.6                DENSITY

The density values used in the current resource estimate are based on 3,397 density measurements collected in 2007 through 2011 by Tahoe Resources from diamond drill core in the Escobal resource area. The samples were grouped according to lithology, oxidation state, and lead mineral domains, as discussed in Sections 14.4 and 14.5. The oxidation state (oxidized, mixed oxidized and sulfidic, and sulfidic) and lead domains (100, 110, 200, and 300) were used as a spatial control on model density due to the correlation with sulfide content, which is the dominant factor in density variation within the mineral model. After completing a statistical review of the density data, MDA eliminated two samples as being outliers or improbable and capped two measurements. Due to potential sample collection bias (the use of whole solid core versus fractured, possibly less-dense core), MDA lowered the values of each group by about 1% for use in the current resource estimate. The lithology, oxidation state, and mineral domains were used to code the block model with the assigned density values.

The density values used in the estimate are shown in the “Model” columns in Table 14-2 and Table 14-3. Table 14-2 lists the density values used for the wallrock lithologies. There are no density data for the overlying Tertiary tuff lithology, and a density value of 2.54 g/cm3 was assigned to this lithology. The lack of density data for the tuff is not considered significant to the current resource estimation because no mineralization has been modeled into this post-mineralization rock unit.

The density values used for the mineralized material, which have the most impact on the resource estimate, are represented in Table 14-3. Within all mineral domains, there was a natural grouping of density values by oxidation state; lower density values for the oxidized and mixed material as compared to the higher density values within the sulfide material. The mineral domain densities vary from 2.52 g/cm3 within the oxidized and mixed low-grade (100 domain) material, to 2.80 g/cm3 within the sulfidic high-grade (300 domain) material within the Central zone. The increased presence of massive sulfide in the Central zone as compared to the East zone is demonstrated in the density values; as a result, unique values for the 300 lead domain mineralization are assigned to each zone. This difference in density between East and West zones for the lower grade domains is not observed, so one density value is used for each of the 100 and 200 mineral domains in both zones. The limited sampling in the oxidized/mixed 300 domain resulted in MDA assigning the same values as those for the 200 domain for this rock type. Additional density measurements for the oxidized/mixed high-grade mineralization are recommended.

The block’s density value is the volume-weighted average of the unmineralized lithologic density, combined with the volumes of lead mineralization domains.

Table 14-2: Lithology Density Values Used in Model

Lithology	#	Mean	Median	Min.	Max.	Std.Dev.	Model
andesite	291	2.63	2.65	2.17	2.87	0.09	2.61
sediment	246	2.65	2.67	2.14	2.99	0.12	2.64
tuff	No data						2.54

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Table 14-3: Mineral Domain Density Values Used in Model

Pb Domain	Ox. State	#	Mean	Median	Min.	Max.	Std.Dev.	Model
100	Ox-Mix	186	2.55	2.55	2.16	2.79	0.09	2.52
200	Ox-Mix	114	2.57	2.57	2.26	2.84	0.09	2.54
300	Ox-Mix	6	2.51	2.52	2.30	2.67	0.15	2.54
110	all	141	2.64	2.63	2.44	3.64	0.10	2.60
100	Sulfide	487	2.65	2.66	2.17	3.10	0.09	2.63
200	Sulfide	1648	2.68	2.68	1.81	3.39	0.10	2.65
300 (East)	Sulfide	123	2.77	2.73	2.45	3.31	0.15	2.72
300 (Central)	Sulfide	605	2.85	2.77	2.15	4.30	0.27	2.80

14.7                SAMPLE CODING AND COMPOSITING

Drill-hole assays were coded by the sectional mineral-domain polygons. After reviewing the statistically similar Central and East assay data, the assays for the two areas were combined into one deposit-wide assay data file per metal for further statistical evaluation and compositing.

MDA analyzed the assay data and capped a total of 164 individual metal analyses which were statistically and spatially deemed beyond a given domain’s natural population of samples. This number of samples capped represents approximately 0.1% of the total assay values within the database. The capped analyses occur within all grade ranges and all estimation areas. Descriptive statistics of the uncapped and capped sample grades by domain are presented in Table 14-4.

Compositing was made at 3 m down-hole lengths, honoring all mineral domain boundaries. Length-weighted composites were used in the block-model grade estimation and the volume inside each mineral domain was estimated using only composites from inside that domain. Composite descriptive statistics for the metal domains are presented in Table 14-5.

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Table 14-4: Escobal mineral Domain Assay Statistics

			Mean	Median			Min.	Max.
Ag Domain	Assays	Count	(g Ag/t)	(g Ag/t)	Std. Dev.	CV	(g Ag/t)	(g Ag/t)
110	Ag	1406	149.1	49.2	338.6	2.27	0.0	7985.4
	Ag Cap	1406	138.5	49.2	236.5	1.71	0.0	1500.0
100	Ag	20071	33.4	15.6	62.7	1.88	0.0	5200.1
	Ag Cap	20071	33.1	15.6	52.8	1.59	0.0	1000.0
200	Ag	9492	373.6	258.4	376.4	1.01	0.6	6696.3
	Ag Cap	9492	372.6	258.4	366.7	0.98	0.6	4500.0
300	Ag	1248	2576.9	1908.1	2183.7	0.85	82.2	26886.6
	Ag Cap	1248	2545.0	1908.1	2040.3	0.80	82.2	15500.0
All	Ag	32217	203.3	33.8	627.3	3.09	0.0	26886.6
	Ag Cap	32217	201.5	33.8	604.4	3.00	0.0	15500.0

			Mean	Median			Min.	Max.
Au Domain	Assays	Count	(g Au/t)	(g Au/t)	Std. Dev.	CV	(g Au/t)	(g Au/t)
110	Au	1406	0.384	0.132	1.410	3.67	0.000	42.628
	Au Cap	1406	0.338	0.132	0.670	1.98	0.000	6.000
10	Au	12964	0.024	0.016	0.037	1.53	0.000	1.808
	Au Cap	12964	0.024	0.016	0.031	1.29	0.000	0.450
100	Au	14407	0.154	0.110	0.157	1.02	0.000	4.860
	Au Cap	14407	0.153	0.110	0.149	0.97	0.000	2.000
200	Au	3135	0.885	0.718	0.645	0.73	0.000	8.208
	Au Cap	3135	0.883	0.718	0.623	0.71	0.000	5.000
300	Au	402	5.308	3.130	10.114	1.91	0.072	167.654
	Au Cap	402	4.698	3.130	5.207	1.11	0.072	40.000
All	Au	32314	0.217	0.053	1.248	5.76	0.000	167.654
	Au Cap	32314	0.208	0.053	0.789	3.80	0.000	40.000

			Mean	Median			Min.	Max.
Pb Domain	Assays	Count	(% Pb)	(% Pb)	Std. Dev.	CV	(% Pb)	(% Pb)
110	Pb	1406	1730	416	4950	2.86	0	70800
	Pb Cap	1406	1653	416	4120	2.49	0	40000
100	Pb	18889	244	122	465	1.91	0	17200
	Pb Cap	18889	241	122	384	1.60	0	6000
200	Pb	12772	2496	1710	2609	1.05	0	92900
	Pb Cap	12772	2489	1710	2466	0.99	0	30000
300	Pb	3519	25788	16000	31213	1.21	41	435950
	Pb Cap	3519	25743	16000	30706	1.19	41	315000
All	Pb	36586	2968	353	11121	3.75	0	435950
	Pb Cap	36586	2958	353	10982	3.71	0	315000

			Mean	Median			Min.	Max.
Zn Domain	Assays	Count	(% Zn)	(% Zn)	Std. Dev.	CV	(% Zn)	(% Zn)
110	Zn	1406	3147	905	8326	2.65	0	146800
	Zn Cap	1406	3046	905	7095	2.33	0	70000
100	Zn	18858	637	368	972	1.53	0	37400
	Zn Cap	18858	626	368	798	1.27	0	9000
200	Zn	12100	4565	3460	4051	0.89	3	80600
	Zn Cap	12100	4563	3460	4009	0.88	3	53000
300	Zn	4217	38683	23700	42705	1.10	147	497200
	Zn Cap	4217	38641	23700	42325	1.10	147	362000
All	Zn	36581	5450	935	17256	3.17	0	497200
	Zn Cap	36581	5436	935	17138	3.15	0	362000

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Table 14-5: Escobal mineral Domain Composite Statistics

		Mean	Median			Min.	Max.
Ag Domain	Count	(g Ag/t)	(g Ag/t)	Std. Dev.	CV	(g Ag/t)	(g Ag/t)
110	692	138.5	60.1	192.3	1.39	0.0	1500.0
100	10877	33.1	19.4	40.7	1.23	0.0	799.7
200	4535	372.6	289.3	285.7	0.77	2.4	3717.5
300	655	2545.0	2150.1	1566.9	0.62	151.5	11209.4
All	16758	201.5	39.4	545.2	2.71	0.0	11209.4

		Mean	Median			Min.	Max.
Au Domain	Count	(g Au/t)	(g Au/t)	Std. Dev.	CV	(g Au/t)	(g Au/t)
110	692	0.338	0.167	0.522	1.54	0.000	6.000
10	7642	0.024	0.019	0.026	1.07	0.000	0.450
100	6847	0.153	0.124	0.112	0.73	0.000	2.000
200	1625	0.883	0.762	0.488	0.55	0.000	4.518
300	248	4.698	3.435	4.294	0.91	0.072	40.000
All	17054	0.208	0.060	0.712	3.43	0.000	40.000

		Mean	Median			Min.	Max.
Pb Domain	Count	(% Pb)	(% Pb)	Std. Dev.	CV	(% Pb)	(% Pb)
110	692	1653	592	3080	1.86	0	40000
100	10982	241	147	307	1.28	0	6000
200	6202	2489	1902	1968	0.79	0	23000
300	1726	25743	18000	26085	1.01	75	299450
All	19602	2958	395	9963	3.37	0	299450

		Mean	Median			Min.	Max.
Zn Domain	Count	(% Zn)	(% Zn)	Std. Dev.	CV	(% Zn)	(% Zn)
110	692	3046	1252	5729	1.88	0	70000
100	10966	626	430	643	1.03	0	9000
200	5941	4563	3809	3128	0.69	3	42200
300	2040	38641	26320	36753	0.95	323	333550
All	19639	5436	1025	15799	2.91	0	333550

14.8                RESOURCE MODEL AND ESTIMATION

The Escobal resource block model replicates the relatively complex metal distributions and geometries observed in the geologic and mineral-domain cross-sectional models. Because of the varied geometries, five separate estimation areas were created at Escobal; one occurring at depth within the Central and East zones (area 1), one unique to the East zone (area 2), and three unique to the Central zone (areas 3, 4, and 5). The locations of these areas relative to the mineral domains are shown in Figure 14-1, Figure 14-2, and Figure 14-3. The estimation areas were modeled with solids, which were used to code the block model.

The portion of each 5 m by 5 m by 2.5 m block inside each mineral domain was estimated using only composites from inside its respective domain. Grade interpolation utilized Inverse Distance Cubed (ID3), with nearest neighbor and ordinary kriging estimates also being made for checking estimation results and sensitivities. Variography and geostatistical evaluations were made to determine distances for search and classification criteria.

Estimation in areas 1, 2, 4, and 5 used two search passes, while estimation in area 3, which encompasses the current underground development and as such has a much greater drilling density than the other four areas, used three search passes. In all estimation areas, successive passes did not overwrite previous estimation passes. The final pass filled the modeled domains. Strict (15 m) search restrictions (pullbacks) were employed for the higher-grade values for all

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four metals within the late-stage quartz calcite vein (domain 110). Search restrictions were also employed within the silver, lead and zinc low-grade domains (domain 100) to control isolated high-grade composites, and in the gold very low-grade and high-grade domains (domains 10 and 300, respectively). The Escobal estimation parameters are given in Table 14-6.

Table 14-6: Escobal Estimation Parameters for Mineral Resources

Description	Parameter
SEARCH PARAMETERS: Estimation Areas 1, 2, 4, and 5
Samples: min./max./max. per hole (1st pass; domain 100)	2 / 18 / 4
Samples: min./max./max. per hole (2nd and 3rd pass; domain 100)	1 / 24 / 4
Samples: min./max./max. per hole (1st pass; domains 200+300)	2 / 12 / 4
Samples: min./max./max. per hole (2nd and 3rd pass; domain 200 +300)	1 / 18 / 4
First Pass Search (m): major/semimajor/minor	150 / 150 / 75
Second Pass Search (m): major/semimajor/minor	500 / 500 / 250
SEARCH PARAMETERS: Estimation Areas 3
Samples: min./max./max. per hole (1st pass; domain 100)	2 / 12 / 3
Samples: min./max./max. per hole (2nd and pass; domain 100)	1 / 18 / 3
Samples: min./max./max. per hole (1st pass; domains 200+300)	2 / 9 / 3
Samples: min./max./max. per hole (2nd and 3rd pass; domains 200+300)	1 / 12 / 3
First Pass Search (m): major/semimajor/minor	75 / 75 / 25
Second Pass Search (m): major/semimajor/minor	150 / 150 / 50
Third Pass Search (m): major/semimajor/minor	500 / 500/ 250
SEARCH ELLIPSOID ORIENTATIONS
Search Bearing/Plunge/Tilt : Estimation area 1	270° / 0° / 90°
Search Bearing/Plunge/Tilt : Estimation area 2	260° / 0° / 65°
Search Bearing/Plunge/Tilt : Estimation area 3	275° / 0° / -60°
Search Bearing/Plunge/Tilt : Estimation area 4	275° / 0° / -55°
Search Bearing/Plunge/Tilt : Estimation area 5	290° / 0° / 90°

SEARCH RESTRICTIONS

Domain	Areas	Grade Threshold	Search Restriction (m)	Estimation Pass
Ag 110	all	>500 g/t	15	all
Pb 110	all	>1.2 %	15	all
Zn 110	all	>3.0 %	15	all
Au 110	all	>2.5 g/t	15	all
Ag 100	all	>150 g/t	40	all
Ag 100	all	>450 g/t	15	all
Pb 100	all	>0.1 %	40	all
Pb 100	all	>0.35 %	15	all
Zn 100	all	>0.25 %	40	all
Zn 100	all	>0.6 %	15	all
Au 10	all	>0.05 g/t	50	all
Au 300	all	>20 g/t	50	all

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14.9                RESOURCE CLASSIFICATION

MDA classified the Escobal resources in order of increasing geological and quantitative confidence into Inferred and Indicated categories defined by the “CIM Definition Standards - For Mineral Resources and Mineral Reserves” in 2005, in compliance with Canadian National Instrument 43-101. CIM mineral resource definitions are given below:

Mineral Resource

Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories. An Inferred Mineral Resource has a lower level of confidence than that applied to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.

A Mineral Resource is a concentration or occurrence of diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals in or on the Earth’s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.

The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which Mineral Reserves may subsequently be defined by the consideration and application of technical, economic, legal, environmental, socio-economic and governmental factors. The phrase ‘reasonable prospects for economic extraction’ implies a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. A Mineral Resource is an inventory of mineralization that under realistically assumed and justifiable technical and economic conditions might become economically extractable. These assumptions must be presented explicitly in both public and technical reports.

Inferred Mineral Resource

An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.

Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be assumed that all or any part of an Inferred Mineral Resource will be upgraded to an Indicated or Measured Mineral Resource as a result of continued exploration. Confidence in the estimate is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure. Inferred Mineral Resources must be excluded from estimates forming the basis of feasibility or other economic studies.

Indicated Mineral Resource

An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics, can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework and to reasonably assume the continuity of mineralization. The Qualified Person must recognize the importance of the Indicated Mineral Resource category to the advancement of the feasibility of the project. An Indicated Mineral Resource estimate

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is of sufficient quality to support a Preliminary Feasibility Study which can serve as the basis for major development decisions.

Measured Mineral Resource

A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough to confirm both geological and grade continuity.

Mineralization or other natural material of economic interest may be classified as a Measured Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that the tonnage and grade of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.

14.10              MINERAL RESOURCES

MDA classified the Escobal resources by a combination of distance to the nearest sample and the number of samples, while at the same time taking into account reliability of underlying data and understanding and use of the geology. There are Measured, Indicated, and Inferred resources within the Escobal deposit at this time. Measured classification is limited to mixed and sulfide material types associated with the closely-spaced underground drilling and development, while Indicated and Inferred resources occur throughout the Escobal deposit. To be classified as Measured, the mixed and sulfide blocks must be within an average distance of 15 m to two silver composites, coming from two different underground drill holes, within a 30 m isotropic search. There are no Measured resources within the oxide portions of the deposit due to some uncertainty in the density data within the oxide material and the limited amount of metallurgical testing of this material type.

To be classified as Indicated, blocks must be within an average distance of 60 m to two silver composites, coming from two different drill holes, within a 120 m isotropic search. All drill holes and oxidation material types are used in the Indicated determination. The gold-dominant upper portions of the East zone are removed from the Indicated classification and considered Inferred only due to limited metallurgy and density analyses, and spatial uncertainty of the high-grade gold. All other estimated mineralization not classified as Measured or Indicated was assigned to be at least Inferred.

None of these issues detract from the overall confidence in the global project resource estimate, but they do detract from confidence in some of the accuracy which MDA believes is required for Measured and Indicated in these specific areas.

Because of the requirement that the resource exists “in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction,” MDA is reporting the resources at an approximate economic cutoff grade that is reasonable for deposits of this nature that will likely be mined by underground methods. As such, some economic considerations were used to determine cutoff grades at which the resource is presented. MDA considered reasonable metal prices and extraction costs and recoveries, albeit in a general sense.

The Escobal reported resource is summarized in Table 14-7, while the Escobal estimation results are tabulated by classification and oxidation state in Table 14-8; the latter table provides the resource numbers at various AgEq cutoff grades to better assess grade-tonnage fluctuation. The stated resource comes from the block-diluted grade within the entire 5 m by 5 m by 2.5 m blocks and is tabulated on a silver-equivalent (“AgEq”) cutoff grade of 130 g AgEq/t. All material, regardless of which metal is present and which is absent, is tabulated. Because multiple metals exist, but on

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a local scale do not necessarily co-exist, the AgEq grade is used for tabulation. Using the individual metal grades of each block, the AgEq grade is calculated using the following formula:

g AgEq/t = g Ag/t + (60 * g Au/t) + (0.0031 * Pb ppm) + (0.003 * Zn ppm)

This formula is based on prices of US$22.00 per ounce silver, US$1,325.00 per ounce gold, US$1.00 per pound lead, and US$0.95 per pound zinc. No metal recoveries are applied, as this is the in situ resource, though expected recoveries are similar across all metals resulting in no change to the stated equivalency formula. Typical cross sections through the Escobal block model showing AgEq block grades for the Central and East zones are given in Figure 14-4, Figure 14-5, and Figure 14-6. These are the same cross-section locations used to depict the mineral-domain models in Section 14.5.

Silver is the dominant metal throughout much of the Escobal deposit accounting for, on average, approximately 80 to 85 percent of the in situ block values. The one exception to the silver-dominant mineralization style is within the upper levels of the East zone where the mineralization is primarily gold with decreased silver, lead, and zinc. The highest estimated gold grades (up to 40 g Au/t) at Escobal occur within this area, and much of the AgEq stated resource is driven by the gold content in that area.

Table 14-7: Escobal Deposit Reported Resource

Escobal Reported Resource (130g AgEq/t cut-off grade)

	Mineral		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
Class	Type	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
Measured	Mixed	260,000	283.9	0.355	0.247	0.570	2,380,000	3,000	1,420,000	3,280,000
Measured	Sulfide	6,270,000	520.9	0.407	0.936	1.628	105,020,000	82,000	129,360,000	225,010,000
Measured	total	6,530,000	511.4	0.405	0.908	1.585	107,400,000	85,000	130,780,000	228,290,000
Indicated	Oxide	590,000	316.4	0.362	0.248	0.513	6,010,000	7,000	3,230,000	6,680,000
Indicated	Mixed	790,000	261.8	0.420	0.163	0.280	6,650,000	11,000	2,840,000	4,890,000
Indicated	Sulfide	31,070,000	314.2	0.316	0.704	1.153	313,860,000	315,000	482,190,000	790,120,000
Indicated	total	32,450,000	312.9	0.319	0.682	1.120	326,520,000	333,000	488,260,000	801,690,000
Meas. + Ind.	Total	38,980,000	346.2	0.334	0.720	1.198	433,920,000	418,000	619,040,000	1,029,980,000
Inferred	Oxide	120,000	112.9	2.830	0.050	0.098	420,000	11,000	130,000	250,000
Inferred	Mixed	330,000	191.4	2.933	0.119	0.217	2,010,000	31,000	860,000	1,560,000
Inferred	Sulfide	970,000	221.2	0.271	0.301	0.568	6,890,000	8,000	6,420,000	12,130,000
Inferred	Total	1,420,000	224.4	1.243	0.250	0.467	9,320,000	50,000	7,410,000	13,940,000

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Table 14-8: Escobal Deposit AgEq Resource Tabulation

Measured material in Mixed:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	390,000	215.0	0.304	0.186	0.424	2,720,000	4,000	1,610,000	3,690,000
100	320,000	248.6	0.334	0.215	0.494	2,560,000	3,000	1,520,000	3,480,000
120	280,000	272.9	0.351	0.237	0.546	2,430,000	3,000	1,450,000	3,330,000
130	260,000	283.9	0.355	0.247	0.570	2,380,000	3,000	1,420,000	3,280,000
140	250,000	294.9	0.367	0.258	0.596	2,320,000	3,000	1,390,000	3,210,000
150	230,000	304.7	0.373	0.268	0.620	2,270,000	3,000	1,370,000	3,160,000
160	220,000	312.8	0.376	0.276	0.639	2,230,000	3,000	1,350,000	3,120,000
170	210,000	321.9	0.385	0.285	0.663	2,170,000	3,000	1,320,000	3,070,000
180	200,000	332.2	0.390	0.296	0.688	2,120,000	2,000	1,290,000	3,010,000
190	190,000	340.1	0.394	0.304	0.709	2,080,000	2,000	1,270,000	2,970,000
200	180,000	347.6	0.399	0.312	0.727	2,030,000	2,000	1,250,000	2,920,000
210	170,000	358.0	0.410	0.323	0.755	1,970,000	2,000	1,220,000	2,850,000
220	160,000	366.0	0.417	0.331	0.774	1,930,000	2,000	1,200,000	2,800,000
230	150,000	377.1	0.423	0.343	0.807	1,870,000	2,000	1,170,000	2,740,000
240	150,000	387.4	0.434	0.354	0.834	1,810,000	2,000	1,130,000	2,670,000
250	140,000	399.4	0.446	0.369	0.872	1,740,000	2,000	1,100,000	2,610,000
300	100,000	463.2	0.484	0.451	1.078	1,450,000	2,000	970,000	2,310,000
350	70,000	548.5	0.501	0.568	1.345	1,180,000	1,000	840,000	1,980,000
400	50,000	640.8	0.499	0.695	1.634	990,000	1,000	740,000	1,730,000
500	30,000	817.0	0.574	0.919	2.128	750,000	1,000	580,000	1,340,000

Measured material in Sulfide:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	7,730,000	437.4	0.353	0.784	1.372	108,720,000	88,000	133,690,000	233,810,000
100	6,900,000	481.5	0.382	0.865	1.508	106,870,000	85,000	131,630,000	229,540,000
120	6,460,000	508.9	0.400	0.914	1.592	105,610,000	83,000	130,110,000	226,480,000
130	6,270,000	520.9	0.407	0.936	1.628	105,020,000	82,000	129,360,000	225,010,000
140	6,100,000	532.7	0.414	0.957	1.663	104,410,000	81,000	128,610,000	223,540,000
150	5,940,000	543.4	0.421	0.976	1.696	103,830,000	80,000	127,940,000	222,210,000
160	5,800,000	553.6	0.427	0.995	1.728	103,250,000	80,000	127,290,000	220,920,000
170	5,660,000	564.2	0.434	1.015	1.761	102,630,000	79,000	126,630,000	219,610,000
180	5,520,000	574.6	0.440	1.035	1.793	102,000,000	78,000	125,940,000	218,260,000
190	5,390,000	585.2	0.447	1.055	1.826	101,330,000	77,000	125,240,000	216,860,000
200	5,250,000	596.0	0.454	1.075	1.860	100,640,000	77,000	124,470,000	215,370,000
210	5,120,000	607.0	0.461	1.096	1.894	99,920,000	76,000	123,670,000	213,770,000
220	4,990,000	618.6	0.468	1.118	1.931	99,140,000	75,000	122,880,000	212,190,000
230	4,860,000	630.0	0.476	1.141	1.968	98,350,000	74,000	122,120,000	210,710,000
240	4,730,000	641.3	0.483	1.163	2.005	97,570,000	73,000	121,330,000	209,160,000
250	4,610,000	653.1	0.491	1.186	2.044	96,740,000	73,000	120,490,000	207,560,000
300	4,020,000	714.4	0.530	1.310	2.247	92,380,000	69,000	116,180,000	199,210,000
350	3,500,000	780.8	0.574	1.444	2.465	87,770,000	65,000	111,340,000	190,020,000
400	3,060,000	848.3	0.619	1.578	2.682	83,320,000	61,000	106,290,000	180,650,000
500	2,370,000	986.4	0.708	1.849	3.126	75,040,000	54,000	96,460,000	163,040,000

Note: prices used for silver equivalent are US$0.95/lb Zn, US$1.00/lb Pb, $22/oz Ag, and US$1,325/oz Au

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Table 14-8: Escobal Deposit AgEq Resource Tabulation (continued)

Indicated material in Oxide:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	800,000	252.8	0.329	0.199	0.415	6,540,000	9,000	3,530,000	7,360,000
100	680,000	286.8	0.349	0.226	0.467	6,270,000	8,000	3,380,000	7,010,000
120	620,000	306.2	0.358	0.240	0.497	6,100,000	7,000	3,280,000	6,790,000
130	590,000	316.4	0.362	0.248	0.513	6,010,000	7,000	3,230,000	6,680,000
140	570,000	324.3	0.365	0.254	0.526	5,930,000	7,000	3,190,000	6,590,000
150	550,000	332.1	0.369	0.261	0.539	5,850,000	6,000	3,150,000	6,510,000
160	530,000	339.4	0.372	0.266	0.550	5,770,000	6,000	3,100,000	6,420,000
170	510,000	347.6	0.377	0.273	0.564	5,680,000	6,000	3,050,000	6,320,000
180	490,000	355.5	0.381	0.279	0.578	5,590,000	6,000	3,010,000	6,230,000
190	470,000	363.6	0.387	0.286	0.592	5,490,000	6,000	2,960,000	6,130,000
200	450,000	370.6	0.391	0.292	0.605	5,400,000	6,000	2,910,000	6,040,000
210	440,000	378.7	0.394	0.298	0.619	5,300,000	6,000	2,860,000	5,940,000
220	420,000	387.0	0.400	0.306	0.636	5,190,000	5,000	2,810,000	5,850,000
230	400,000	396.1	0.405	0.314	0.654	5,070,000	5,000	2,760,000	5,740,000
240	380,000	403.4	0.410	0.321	0.669	4,970,000	5,000	2,720,000	5,650,000
250	370,000	412.6	0.414	0.330	0.686	4,850,000	5,000	2,660,000	5,530,000
300	290,000	460.2	0.440	0.377	0.788	4,230,000	4,000	2,380,000	4,970,000
350	230,000	508.1	0.458	0.428	0.892	3,670,000	3,000	2,120,000	4,420,000
400	180,000	548.7	0.462	0.482	1.008	3,230,000	3,000	1,940,000	4,070,000
500	110,000	639.2	0.496	0.641	1.356	2,280,000	2,000	1,570,000	3,310,000

Indicated material in Mixed:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	1,120,000	207.3	0.368	0.134	0.237	7,470,000	13,000	3,300,000	5,860,000
100	930,000	236.6	0.398	0.150	0.261	7,050,000	12,000	3,060,000	5,330,000
120	830,000	253.8	0.412	0.159	0.274	6,790,000	11,000	2,920,000	5,030,000
130	790,000	261.8	0.420	0.163	0.280	6,650,000	11,000	2,840,000	4,890,000
140	760,000	268.9	0.428	0.167	0.286	6,520,000	10,000	2,780,000	4,750,000
150	720,000	275.8	0.436	0.171	0.291	6,390,000	10,000	2,720,000	4,630,000
160	690,000	283.1	0.444	0.175	0.297	6,250,000	10,000	2,640,000	4,490,000
170	660,000	289.5	0.451	0.178	0.301	6,120,000	10,000	2,580,000	4,360,000
180	630,000	296.6	0.459	0.182	0.306	5,970,000	9,000	2,510,000	4,220,000
190	600,000	303.6	0.471	0.186	0.312	5,810,000	9,000	2,440,000	4,100,000
200	570,000	309.9	0.480	0.190	0.318	5,660,000	9,000	2,380,000	3,980,000
210	530,000	318.2	0.492	0.194	0.325	5,470,000	8,000	2,290,000	3,830,000
220	510,000	325.4	0.502	0.198	0.330	5,300,000	8,000	2,210,000	3,690,000
230	480,000	332.5	0.514	0.201	0.336	5,140,000	8,000	2,130,000	3,560,000
240	460,000	339.2	0.523	0.206	0.343	4,980,000	8,000	2,070,000	3,450,000
250	430,000	346.5	0.534	0.210	0.350	4,810,000	7,000	2,000,000	3,320,000
300	310,000	389.6	0.601	0.235	0.382	3,850,000	6,000	1,590,000	2,590,000
350	180,000	464.6	0.725	0.282	0.432	2,690,000	4,000	1,120,000	1,720,000
400	110,000	560.3	0.852	0.321	0.477	1,940,000	3,000	760,000	1,130,000
500	5 0,000	762.9	0.930	0.407	0.612	1,250,000	2,000	460,000	690,000

Indicated material in Sulfide:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	44,220,000	242.5	0.256	0.556	0.926	344,780,000	364,000	541,870,000	902,640,000
100	36,720,000	279.1	0.286	0.631	1.043	329,530,000	338,000	511,010,000	844,370,000
120	32,700,000	303.3	0.306	0.681	1.119	318,840,000	322,000	491,210,000	807,020,000
130	31,070,000	314.2	0.316	0.704	1.153	313,860,000	315,000	482,190,000	790,120,000
140	29,590,000	324.7	0.325	0.726	1.187	308,900,000	309,000	473,610,000	774,190,000
150	28,210,000	335.0	0.334	0.748	1.220	303,870,000	303,000	465,280,000	758,740,000
160	26,920,000	345.3	0.343	0.770	1.252	298,880,000	297,000	456,830,000	743,180,000
170	25,740,000	355.2	0.352	0.791	1.284	293,920,000	292,000	449,040,000	728,670,000
180	24,630,000	365.0	0.361	0.813	1.316	288,980,000	286,000	441,340,000	714,410,000
190	23,560,000	374.9	0.371	0.834	1.348	283,960,000	281,000	433,360,000	699,910,000
200	22,510,000	385.1	0.380	0.857	1.382	278,700,000	275,000	425,300,000	685,510,000
210	21,530,000	395.1	0.389	0.880	1.416	273,510,000	270,000	417,780,000	671,960,000
220	20,570,000	405.5	0.399	0.905	1.452	268,150,000	264,000	410,220,000	658,270,000
230	19,640,000	416.1	0.410	0.930	1.489	262,680,000	259,000	402,380,000	644,490,000
240	18,750,000	426.6	0.420	0.956	1.528	257,240,000	253,000	395,090,000	631,560,000
250	17,920,000	437.3	0.430	0.981	1.566	251,870,000	248,000	387,620,000	618,390,000
300	14,120,000	495.4	0.486	1.120	1.767	224,810,000	221,000	348,380,000	549,870,000
350	11,210,000	556.9	0.540	1.255	1.962	200,700,000	194,000	310,080,000	484,800,000
400	8,960,000	622.2	0.596	1.391	2.156	179,180,000	172,000	274,770,000	425,770,000
450	7,210,000	692.8	0.651	1.514	2.335	160,500,000	151,000	240,480,000	370,890,000
500	5,810,000	770.9	0.706	1.632	2.501	143,980,000	132,000	208,980,000	320,310,000

Note: prices used for silver equivalent are US$0.95/lb Zn, US$1.00/lb Pb, $22/oz Ag, and US$1,325/oz Au

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Table 14-8: Escobal Deposit AgEq Resource Tabulation (continued)

Inferred material in Oxide:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	160,000	92.6	2.315	0.044	0.093	480,000	12,000	160,000	330,000
100	140,000	101.5	2.560	0.046	0.095	450,000	11,000	140,000	290,000
120	120,000	109.5	2.743	0.049	0.097	430,000	11,000	130,000	260,000
130	120,000	112.9	2.830	0.050	0.098	420,000	11,000	130,000	250,000
140	110,000	116.7	2.918	0.051	0.098	410,000	10,000	120,000	240,000
150	110,000	120.0	2.980	0.052	0.098	410,000	10,000	120,000	230,000
160	100,000	124.6	3.056	0.054	0.099	400,000	10,000	120,000	220,000
170	90,000	129.1	3.128	0.055	0.100	390,000	9,000	110,000	210,000
180	90,000	133.3	3.209	0.057	0.101	380,000	9,000	110,000	200,000
190	80,000	137.3	3.270	0.058	0.102	370,000	9,000	110,000	190,000
200	80,000	140.8	3.330	0.058	0.101	360,000	9,000	100,000	180,000
210	80,000	145.6	3.387	0.060	0.102	350,000	8,000	100,000	170,000
220	70,000	148.5	3.446	0.061	0.103	340,000	8,000	100,000	160,000
230	70,000	153.3	3.501	0.062	0.104	330,000	8,000	90,000	150,000
240	60,000	157.9	3.552	0.064	0.103	320,000	7,000	90,000	150,000
250	60,000	161.1	3.604	0.065	0.104	320,000	7,000	90,000	140,000
300	40,000	185.3	3.871	0.071	0.109	260,000	6,000	70,000	110,000
350	30,000	212.7	4.190	0.079	0.107	210,000	4,000	50,000	70,000
400	20,000	264.1	4.181	0.101	0.114	170,000	3,000	40,000	50,000
500	10,000	379.1	5.082	0.201	0.165	70,000	1,000	30,000	20,000

Inferred material in Mixed:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	420,000	159.8	2.473	0.104	0.197	2,150,000	33,000	960,000	1,820,000
100	370,000	174.4	2.691	0.111	0.208	2,090,000	32,000	910,000	1,700,000
120	340,000	186.5	2.856	0.117	0.215	2,040,000	31,000	880,000	1,610,000
130	330,000	191.4	2.933	0.119	0.217	2,010,000	31,000	860,000	1,560,000
140	310,000	198.2	3.019	0.122	0.220	1,990,000	30,000	840,000	1,510,000
150	300,000	205.9	3.123	0.125	0.225	1,950,000	30,000	810,000	1,460,000
160	280,000	212.5	3.205	0.129	0.231	1,920,000	29,000	800,000	1,430,000
170	270,000	218.7	3.268	0.132	0.236	1,900,000	28,000	790,000	1,410,000
180	260,000	224.5	3.336	0.135	0.241	1,870,000	28,000	780,000	1,380,000
190	250,000	232.3	3.412	0.139	0.248	1,850,000	27,000	760,000	1,350,000
200	240,000	239.2	3.477	0.141	0.252	1,820,000	26,000	740,000	1,320,000
210	230,000	244.2	3.536	0.144	0.256	1,800,000	26,000	730,000	1,290,000
220	220,000	250.0	3.597	0.146	0.260	1,770,000	26,000	710,000	1,270,000
230	210,000	256.7	3.661	0.149	0.266	1,750,000	25,000	700,000	1,240,000
240	200,000	265.4	3.731	0.153	0.273	1,720,000	24,000	680,000	1,210,000
250	190,000	272.1	3.791	0.156	0.279	1,690,000	24,000	670,000	1,190,000
300	160,000	310.9	4.080	0.173	0.310	1,560,000	20,000	590,000	1,070,000
350	130,000	347.0	4.376	0.189	0.337	1,420,000	18,000	530,000	950,000
400	100,000	383.8	4.706	0.203	0.362	1,280,000	16,000	460,000	830,000
500	7 0,000	458.6	5.162	0.230	0.405	1,060,000	12,000	370,000	640,000

Inferred material in Sulfide:

Cutoff		Silver	Gold	Lead	Zinc	Silver	Gold	Lead	Zinc
g AgEq/t	Tonnes	(g Ag/t)	(g Au/t)	(% Pb)	(% Zn)	(oz)	(oz)	(lbs)	(lbs)
75	2,280,000	137.6	0.152	0.216	0.408	10,080,000	11,000	10,840,000	20,510,000
100	1,370,000	183.2	0.217	0.272	0.511	8,060,000	10,000	8,210,000	15,410,000
120	1,080,000	208.5	0.254	0.289	0.547	7,260,000	9,000	6,910,000	13,070,000
130	970,000	221.2	0.271	0.301	0.568	6,890,000	8,000	6,420,000	12,130,000
140	860,000	234.5	0.294	0.315	0.594	6,500,000	8,000	5,980,000	11,290,000
150	790,000	245.1	0.312	0.322	0.606	6,220,000	8,000	5,610,000	10,550,000
160	730,000	254.9	0.326	0.329	0.617	5,990,000	8,000	5,300,000	9,940,000
170	670,000	265.7	0.339	0.333	0.620	5,750,000	7,000	4,940,000	9,210,000
180	630,000	274.9	0.352	0.342	0.637	5,520,000	7,000	4,720,000	8,780,000
190	580,000	284.4	0.372	0.353	0.653	5,280,000	7,000	4,500,000	8,320,000
200	540,000	293.2	0.387	0.361	0.665	5,080,000	7,000	4,280,000	7,900,000
210	480,000	306.3	0.414	0.378	0.692	4,770,000	6,000	4,030,000	7,390,000
220	440,000	318.9	0.441	0.386	0.701	4,510,000	6,000	3,750,000	6,800,000
230	400,000	331.2	0.468	0.394	0.716	4,280,000	6,000	3,490,000	6,350,000
240	380,000	340.7	0.483	0.395	0.714	4,130,000	6,000	3,290,000	5,940,000
250	350,000	351.2	0.498	0.392	0.707	3,980,000	6,000	3,040,000	5,490,000
300	260,000	396.7	0.554	0.379	0.681	3,360,000	5,000	2,200,000	3,960,000
350	190,000	444.9	0.609	0.404	0.726	2,700,000	4,000	1,680,000	3,030,000
400	130,000	504.8	0.689	0.429	0.779	2,080,000	3,000	1,220,000	2,200,000
500	60,000	666.5	0.930	0.468	0.880	1,190,000	2,000	580,000	1,080,000

Note: prices used for silver equivalent are US$0.95/lb Zn, US$1.00/lb Pb, $22/oz Ag, and US$1,325/oz Au

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Figure 14-4: Section 806300 - Escobal Central Zone Block Model: AgEq Block Grades

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Figure 14-5: Section 806800 - Escobal Central Zone Block Model: AgEq Block Grades

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Figure 14-6: Section 807500 - Escobal East Zone Block Model: AgEq Block Grades

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Checks were made on the Escobal resource model in the following manner:

1.Cross sections with the mineral domains, drill-hole assays and geology, topography, sample coding, and block grades with classification were reviewed for reasonableness;

2.Block-model information, such as coding, number of samples, and classification were checked visually by domain and lithology on sections and level plans;

3.Cross-section mineral domain volumes to level plan mineral domain volumes, to block model mineral domain volumes were checked;

4.Nearest-neighbor and indicator kriging models were made for comparison;

5.Sectional polygonal models were calculated from the original modeled section domains; and

6.Quantile-quantile plots of assays, composites, and block-model grades were made to evaluate differences in distributions of metals throughout all domains and areas.

The resource estimate is considered reasonable, honors the geology, and is supported by the geologic model.

14.11DISCUSSION, QUANTIFICATIONS, RISK, AND RECOMMENDATIONS

The Escobal deposit’s Central zone hosts laterally continuous mineralization over a 1,500 m strike and up to 1000 m down-dip. The East zone mineralization is up to 850 m along strike and 950 m down-dip. Sulfide mineralization is dominant, with silver, lead, and zinc occurring in potentially economic grades throughout most of the sulfide mineralization. Gold distribution is more erratic within the sulfide mineralization but can be high grade (>10 g Au/t) within the oxidized portions of the East zone where gold is the dominant metal.

The Escobal resource estimate is based on sufficient drill-sample analytical and density measurements, detailed drill-hole lithology and alteration data, and metallurgical results, to support a classification of Indicated for much of the Escobal mineralization. The above factors, along with the closely-spaced nature of the underground drilling, result in a classification of Measured for some of the resource within the current underground development. The limited metallurgical testing on the oxide material and some spatial uncertainty in the model have resulted in an Inferred classification for the gold-dominant mineralization within the upper levels of the East zone.

Both the Central and East zones are open at depth, and further extensional drilling is recommended. This drilling might be better accomplished from underground drill stations than targeting from the surface.

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15MINERAL RESERVE ESTIMATES

15.1MINERAL RESERVE CLASSIFICATION

Blattman Brothers Consulting LLC (Blattman) classified the Escobal reserves in order of increasing confidence into Probable and Proven categories as defined by the CIM Definition Standards — For Mineral Resources and Mineral Reserves in 2005, in compliance with Canadian National Instrument 43-101. CIM mineral reserve definitions are given below:

15.1.1Mineral Reserve

Mineral Reserves are sub-divided in order of increasing confidence into Probable Mineral Reserves and Proven Mineral Reserves. A Probable Mineral Reserve has a lower level of confidence than a Proven Mineral Reserve.

A Mineral Reserve is the economically mineable part of a Measured and/or Indicated Mineral Resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as appropriate that include application of Modifying Factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which Mineral Reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

The public disclosure of a Mineral Reserve must be demonstrated by a Pre-Feasibility Study or Feasibility Study.

Mineral Reserves are those parts of Mineral Resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the Qualified Person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant Modifying Factors. Mineral Reserves are inclusive of diluting material that will be mined in conjunction with the Mineral Reserves and delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

15.1.2Probable Mineral Reserve

A Probable Mineral Reserve is the economically mineable part of an Indicated, and in some circumstances, a Measured Mineral Resource. The confidence in the Modifying Factors applying to a Probable Mineral Reserve is lower than that applying to a Proven Mineral Reserve.

The Qualified Person(s) may elect to convert Measured Mineral Resources to Probable Mineral Reserves if the confidence in the Modifying Factors is lower than that applied to a Proven Mineral Reserve. Probable Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study.

15.1.3Proven Mineral Reserve

A Proven Mineral Reserve is the economically mineable part of a Measured Mineral Resource. A Proven Mineral Reserve implies a high degree of confidence in the Modifying Factors.

Application of the Proven Mineral Reserve category implies that the Qualified Person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should

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be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study.

15.2MINERAL RESERVE STATEMENT

The Proven & Probable Reserve for the Escobal mine is 31.4 Mt at average grades of 347 grams of silver per tonne, 0.33 grams of gold per tonne, 0.74 percent lead and 1.21 percent zinc; containing 350.5 million ounces of silver, 335,600 ounces of gold, 232,100 tonnes of lead and 381,600 tonnes of zinc. The Mineral Reserve is based on the Mineral Resource Estimate presented in Section 14. As is the case with the Mineral Resources, the Escobal mine Mineral Reserves are reported using the CIM Code and are dated July 1, 2014. Table 15-1 presents a summary of the Mineral Reserve Estimate.

Blattman derived minable stope shapes for the deposit based solely upon the material reported as Measured and Indicated Mineral Resources. The mass of any Inferred Resources contained within a stope shape is included in the Reserve but the metal associated with the material is not (i.e. the Inferred mass has a grade of 0.0 for all metals). This has the overall effect of diluting the reserve.

Table 15-1: Mineral Reserve Estimate

	Tonnes	Ag	Au	Pb	Zn
Classification	(M)	g/t	g/t	%	%
Proven Reserves	6.0 M	457	0.37	0.86	1.51
Probable Reserves	25.4 M	321	0.32	0.71	1.14
Total Proven & Probable Reserves	31.4 M	347	0.33	0.74	1.21

	Ag Ounces	Au Ounces	Pb Tonnes	Zn Tonnes
Classification	(M)	(000s)	(000s)	(000s)
Proven Reserves	87.8	70.8	51.7	90.2
Probable Reserves	262.7	265.2	180.4	291.3
Total Proven & Probable Reserves	350.5	336.0	232.1	381.5

The effective date of the Escobal mine Mineral Reserve Estimate is July 1, 2014 and the model has been depleted to account for extraction of mineralized resources prior to this date. Mineral Reserves are inclusive of Mineral Resources.

The Proven and Probable reserves listed in Table 15-1 are comprised of 4.1 Mt of Measured Resource at average grades of 652 grams of silver per tonne, 0.49 grams of gold per tonne, 1.17 percent lead and 2.03 percent zinc; 17.4 Mt of Indicated Resource at average grades of 453 grams of silver per tonne, 0.43 grams of gold per tonne, 0.94 percent lead and 1.50 percent zinc; 9.3 Mt of waste material at average grades of 35 grams of silver per tonne, 0.11 grams of gold per tonne, 0.22 percent lead and 0.39 percent zinc and 0.6 Mt of paste fill dilution at zero grade for all metals

The tonnes and grades reported do not have any additional mine recovery factors applied beyond the creation of the minable stope shape. The secondary transverse stopes include an additional one meter of paste fill dilution (0.5m from each side wall); this additional dilution is included at a zero grade and a density of 1.6. The overall dilution rate is approximately 31% (mass of dilution/total mass). Subeconomic material internal to the stope designs and external

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mining dilution account for approximately 20% and 9% of the dilution total, respectively. Paste backfill dilution accounts for about 2%.

Though not quantifiable, Blattman acknowledges that a portion of the reported dilution results from applying realistic stope geometries to resource model blocks with dimensions of 5m(X) by 2.5m(Y) x 5.0m(Z) often creates coarse ‘stair-step’ boundaries of ore grade blocks along the footwall and hanging wall of the block model that do not accurately portray the true ore/geological boundaries. The actual dilution when mining may be lower than reported. Future resource modeling should investigate methods of more accurate depiction of the modeled grade domain boundaries in all directions.

The Mineral Reserve Estimate is based on material provided as mill feed to the Escobal Mill and does not include process recovery factors or additional plant losses.

Blattman is not aware of any metallurgical, infrastructural, environmental, legal, title, taxation, socio-economic or marketing issues that would impact the mineral reserve statement as presented.

15.3CUTOFF GRADE

The cut-off value calculation used to determine the minable portion of the Measured and Indicated Resources at Escobal is predicated on the assumption that mine production will feed the mill at capacity throughout the mine life. Production of a discretionary increment of material will extend the life of the operation and therefore increase the total amount of fixed costs generated over the life of the operation. Production of the discretionary increment defers the realization of production from other increments. All costs that are incremental with production must therefore be covered in the cut-off value calculation. Costs in the cut-off value calculation include the variable and fixed costs directly related to ore production, expensed stope access development, smelting, refining, and concentrate transportation, general and administrative costs directly related to production, royalties, and project costs related to production and the plant facilities that do not have a measurable payback.

Costs excluded from the basis include exploration, capitalized development costs, capital infrastructure costs, in-mine projects having a measurable economic benefit, and non-cash charges.

Sustaining capital and expansion capital costs are excluded from the cut-off value cost basis as these costs are not incremental to a specific unit of production but rather common to large portions of the mineral deposit.

Cut-off grades to define the Mineral Reserves were calculated from the NSR value of the ore minus the production cost to account for variability in mining method and metallurgical response. NSR value was determined using metal prices of US$22.50 per ounce silver, US$1,300.00 per ounce gold, US$0.95 per pound lead and US$0.90 per pound zinc. The silver price used in this analysis is slightly optimistic in comparison to the market prices of July 2014 (typically ranging from US$20.50 to US$21.50). The metal prices for this section of the study are primarily used in the definition of minable portions of deposit. By using the slightly optimistic value, the continuity of the ore-grade mineralization improves and more realistic stope shapes are possible. These metal prices are only used in the definition of the minable mining shapes and the determination of ore and waste. The economic analysis section of this study (Section 22) includes a more detailed analysis of metals prices, sensitivities and the economic viability of the mine.

Actual operating costs, metallurgical performance, and smelter contracts from the Escobal mine and engineering first-principles were used to derive operating costs and revenue. Note that the base case economic parameters used in the financial model in this study may vary from the NSR model inputs due to additional metallurgical and other related knowledge acquired during the study. Blattman considers the magnitude of any parameter variation to have no material impact on the reserve estimate.

Table 15-2 through Table 15-6 list the costs and other parameters utilized to calculate the Escobal cut off-value.

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Table 15-2: Cutoff Value Assumptions — Metal Prices

Ag	$	22.50/oz
Au	$	1,300.00/oz
Pb	$	0.95/lb
Zn	$	0.90/lb

Table 15-3: Cutoff Value Assumptions — Process Recoveries

	Recovery		Recovery		Total
Process Recovery	Pb Conc.		Zn Conc.		Recovery
Ag	83.1	%	3.4	%	86.5	%
Au	61.6	%	4.4	%	65.9	%
Pb	88.9	%	2.1	%	91.0	%
Zn	18.1	%	72.1	%	90.2	%

Table 15-4: Cutoff Value Assumptions — % Payable Metals in Concentrate

	% Payable in		% Payable in
Payable Metal	Pb Conc.		Zn Conc.
Ag	96	%	75	%
Au	95	%	75	%
Pb	95	%	0	%
Zn	0	%	85	%

Table 15-5: Cutoff Value Assumptions — Initial Operating Costs

Mining Cost per tonne ore	$	50.32
Processing Cost per tonne ore	$	29.81
Surface Operations Cost per tonne ore	$	2.70
G&A Costs per tonne ore	$	20.92
Subtotal	$	103.84
Less Capital Development Costs per tonne ore	$	(7.32	)
Total Operating Costs	$	96.52

Table 15-6: Cutoff Value Assumptions — Transport & Treatment Charges

Pb Concentrate Transport & Treatment ($/tonne of conc.)	$	453.87
Zn Concentrate Transport & Treatment ($/tonne of conc.)	$	413.49
Ag Refining ($/oz Ag)	$	1.04
Au Refining ($/oz Au)	$	9.08

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15.4MINING SHAPES

Blattman created practical mining shapes for each stope by first extracting the centroids of model blocks having a net value greater than $0/tonne. The ore-grade blocks located within the extents of a proposed stope were used as inputs to a plane-fitting algorithm which determined the hanging wall (north side) and foot wall (south side) faces for the stope. The extent of each stope is generally 25 m high by 10 m wide (east-west direction). The extent in the north-south varies according to the geometry of the modeled mineralization.

While accurately mimicking the local geometry of the block model, the plane-fitting algorithm by itself does not always result in a practical mining shape. For example, the mineralization may have a local dip angle less than 45° from the horizontal, an angle generally unachievable with the mining equipment in use at Escobal. Therefore, the plane-fitting routine included a second pass to modify the original planes to more practical shapes. In general, the transverse stopes were allowed less variability than the longitudinal stopes. Table 15-7 below lists the input parameters for the stope shape creation process.

Table 15-7: Stope Shape Parameters

Stope Design Parameter	Transverse Stopes	Longitudinal Stopes
Height	25 m	25 m
Width	10 m	varies
Length	varies	10 m
Minimum number of points to form plane	75 pts	75 pts
Foot wall dip range (south face)	-90° to -45°	±30° from vertical
Hangingwall dip range (north face)	±10° from vertical	±30° from vertical
Strike range (horizontal plane)	70° to 110°	55° to 125°
Minimum volume constraint	2500m3	750m3

Figure 15-1: Transverse Stope Shape Parameters

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Figure 15-2: Longitudinal Stope Shape Parameters

The preliminary shapes were individually refined where necessary to ensure stope geometry viability and to minimize the amount of sub-economic material within the shape volume that is inseparable from profitable material due to the practical constraints of mining.

The decision to design a proposed stope as a longitudinal or transverse stope was an engineering decision, mainly guided by the local continuity and horizontal width of the mineralization. Each proposed stope was reviewed to assure minimum volume restrictions; specifically, any transverse stope less than 2500 m3 or longitudinal stope less than 750m3 was subject to further processing to widen the stope to meet the minimum volume requirements.

After ensuring adherence to the minimum volume restriction, Blattman determined the average grade and tonnage for each stope. This included calculating the net value for the entire stope in terms of $/tonne. Any stope shape with a net value less than $0/tonne was discarded and not included in the final reservation tabulation.

15.5DILUTION AND RECOVERY ESTIMATES

Ore dilution in the secondary stopes includes 0.5 m of overbreak (1.0 m total per stope) into the adjacent backfilled stopes, adding paste fill materials to the mineralized and unmineralized rock in the stope. This effectively increases the total mass of material extracted from the stope and reduces the overall grade. In addition, the mass of any Inferred Resources contained within the stope design are assumed to carry no metal values in the estimation of Mineral Reserve grades.

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The external dilution (i.e. the waste rock on the hanging wall and footwall surfaces of the stope) averages 8% in the transverse stopes and 12% in the longitudinal stopes. Internal waste accounts for an additional 24% in the transverse stopes and 13% in the longitudinal stopes. Total rock dilution, prior to adding paste fill dilution, averages 33% in the transverse stopes and 26% in the longitudinal stopes.

In the evaluation of the Mineral Reserves, no additional modifying factors were applied to the tonnages and grade of the mining shapes to account for mining losses.

Approximately 17.4 Mt of Measured and Indicated Resources, grading 167 g Ag/t, 0.20 g Au/t, 0.39% Pb and 0.70% Zn are excluded from the Mineral Reserves at this time. The primary reasons for excluding these resources from the mine design include:

·Resources located within a stope that did not meet the minimum net value ($0/tonne). Although this mineralized material may be of sufficient grade and confidence to be called a resource, it cannot be economically recovered using the proposed mining methods at this time.

·Immediate proximity to the current development design, precluding mining of the resource. The primary ramp and development access designs were completed prior to finalization of the resource model. In some cases, mineralized resources are located too close to the development to allow for clearance between development and production areas.

·Distal location of the excluded resources relative to the planned mine development. In some cases, the resources are located too far away from the primary mine workings to justify the additional mine development.

·Local geometry of the mineralization in relation to the hanging wall or footwall planes results in ore losses. Because the stope shapes are based on a single plane in the hanging wall and footwall planes, there will be inevitable losses.

The first three items above account for 15.9 Mt of resource, grading 141 g Ag/t, 0.19 g Au/t, 0.36% Pb and 0.66% Zn. In general, this mineralized material is of lower grade than the rest of the resource. The location of these resources outside of the Mineral Reserve can be seen in Figure 15-3.

The last item, ore losses, accounts for 1.5 Mt grading 434 g Ag/t, 0.33 g Au/t, 0.68% Pb and 1.10 % Zn. Table 15-8 shows the ore-loss calculation due to the local geometry variability in relation to the stope plane-fitting process. In general, the current mine design extracts approximately 93% of all M&I resources above cutoff grade and within the immediate area of the proposed mine design. The stopes tend to incur nearly 7% ore losses on the hanging wall and footwall surfaces.

Table 15-9 shows the accounting by level of the resources above cutoff grade, waste rock and paste fill included in the stope designs. Note that minor differences between the various tables may be present due to reporting precision and do not represent material errors in the estimate.

If a change occurs to the cut-off value input parameters, the development design or the mining method, a complete design would be required to determine how much, if any, of the excluded resource might become part of the reserve. Blattman cannot speculate on the potential impact to the reserves.

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Figure 15-3: Generalized Location of Resources and Reserves

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Table 15-8: Mineral Resources in Relation to the Mine Design

	Available M&I Resources Above Cutoff					M&I Resources Inside Stope Shapes					Ore Losses
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	11,089	319	3.46	0.37	0.46	10,411	323	3.58	0.39	0.48	678	263	1.65	0.15	0.20
1590	43,547	510	5.06	0.46	0.51	39,742	526	5.21	0.47	0.52	3,805	351	3.54	0.34	0.41
1565	73,536	871	2.13	0.73	0.68	67,516	921	2.21	0.76	0.71	6,020	312	1.16	0.29	0.37
1540	142,589	898	1.23	0.41	0.67	129,957	937	1.30	0.42	0.70	12,632	494	0.48	0.25	0.37
1515	166,355	922	1.02	0.56	0.83	150,791	977	1.05	0.59	0.87	15,564	386	0.72	0.22	0.38
1490	216,067	1,024	0.70	0.57	0.78	197,330	1,080	0.73	0.60	0.82	18,737	429	0.33	0.22	0.34
1465	264,313	990	0.40	0.55	0.81	246,093	1,025	0.41	0.57	0.83	18,220	521	0.31	0.27	0.47
1440	228,594	867	0.31	0.53	0.82	204,400	882	0.31	0.56	0.85	24,194	735	0.34	0.34	0.53
1415	285,006	764	0.35	0.58	0.85	229,385	825	0.37	0.63	0.91	55,621	513	0.27	0.39	0.59
1390	437,839	735	0.37	0.52	0.82	391,958	752	0.38	0.53	0.82	45,881	589	0.27	0.51	0.81
1365	849,728	553	0.29	0.45	0.76	750,400	566	0.30	0.46	0.78	99,328	451	0.22	0.39	0.63
1340	1,405,740	562	0.27	0.58	0.99	1,260,793	567	0.27	0.59	1.00	144,947	520	0.23	0.56	0.97
1315	1,772,449	589	0.32	0.69	1.15	1,652,519	594	0.32	0.70	1.16	119,930	512	0.29	0.57	0.98
1290	1,790,611	558	0.31	0.64	1.12	1,640,177	557	0.31	0.64	1.13	150,434	567	0.33	0.60	0.99
1265	1,554,641	485	0.34	0.63	1.07	1,376,320	486	0.33	0.63	1.07	178,321	478	0.37	0.68	1.09
1240	1,732,567	538	0.39	0.98	1.67	1,635,602	548	0.40	1.01	1.73	96,965	365	0.21	0.34	0.69
1215	1,370,294	493	0.43	1.18	1.96	1,288,043	500	0.43	1.22	2.03	82,251	400	0.42	0.56	0.90
1190	1,185,467	445	0.36	0.74	1.29	1,102,109	451	0.37	0.77	1.33	83,358	369	0.23	0.43	0.80
1165	1,056,454	421	0.33	0.60	1.08	1,016,866	424	0.34	0.61	1.10	39,588	341	0.22	0.36	0.64
1140	948,530	375	0.34	0.70	1.43	915,895	377	0.34	0.71	1.46	32,635	320	0.29	0.44	0.86
1115	788,651	366	0.45	0.83	1.32	756,878	369	0.46	0.85	1.34	31,773	299	0.36	0.37	0.73
1090	701,409	387	0.60	1.22	2.18	679,398	389	0.60	1.24	2.22	22,011	311	0.51	0.46	0.96
1065	692,259	389	0.70	1.25	2.55	664,114	392	0.71	1.28	2.61	28,145	319	0.41	0.55	1.15
1040	709,360	379	0.71	1.14	2.00	682,608	381	0.72	1.16	2.02	26,752	321	0.44	0.76	1.55
1015	676,114	360	0.57	1.03	1.76	647,275	363	0.58	1.05	1.79	28,839	305	0.35	0.65	1.22
990	684,672	366	0.48	0.81	1.51	655,754	368	0.49	0.82	1.52	28,918	313	0.37	0.68	1.22
965	636,840	356	0.49	0.79	1.44	603,504	358	0.49	0.79	1.45	33,336	329	0.49	0.80	1.27
940	577,403	341	0.47	1.11	1.80	546,049	343	0.48	1.12	1.82	31,354	304	0.37	0.96	1.50
915	508,839	355	0.44	1.27	2.09	486,446	357	0.45	1.29	2.09	22,393	329	0.38	1.05	2.19
890	404,702	328	0.41	1.53	2.68	384,610	333	0.42	1.53	2.66	20,092	225	0.26	1.65	3.03
865	329,980	328	0.31	1.55	2.81	318,064	328	0.31	1.57	2.84	11,916	354	0.37	1.27	2.37
840	245,386	329	0.34	2.21	3.73	229,420	331	0.34	2.21	3.75	15,966	300	0.32	2.13	3.35
815	155,826	310	0.49	3.08	4.70	143,685	312	0.49	3.19	4.81	12,141	293	0.45	1.84	3.51
790	114,261	297	0.62	5.50	6.53	105,306	300	0.63	5.76	6.62	8,955	260	0.52	2.44	5.40
765	115,159	276	0.56	5.00	3.32	102,264	282	0.56	5.11	3.35	12,895	221	0.59	4.13	3.04
740	91,724	173	0.42	5.15	4.89	85,347	171	0.40	5.12	4.96	6,377	200	0.60	5.57	4.02
715	81,150	152	0.37	6.75	7.16	74,002	150	0.36	6.83	7.32	7,148	169	0.40	5.93	5.58
690	52,108	147	0.29	8.64	8.23	48,794	148	0.29	8.71	8.31	3,314	129	0.20	7.57	7.10
665	24,161	213	0.35	6.24	4.93	21,799	213	0.36	6.37	5.00	2,362	211	0.25	5.04	4.27
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	23,125,420	487	0.43	0.96	1.57	21,541,624	491	0.44	0.98	1.60	1,583,796	434	0.33	0.68	1.10

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Table 15-9: Mineral Reserve Estimate by Elevation

	M&I Resources Inside Stope Shapes					Waste Blocks					Paste	P&P Reserves
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	10,411	323	3.58	0.39	0.48	5,455	21	0.30	0.05	0.09	—	15,866	219	2.45	0.27	0.35
1590	39,742	526	5.21	0.47	0.52	14,966	18	0.35	0.05	0.09	—	54,708	387	3.88	0.36	0.40
1565	67,516	921	2.21	0.76	0.71	17,151	37	0.20	0.16	0.22	—	84,667	742	1.80	0.64	0.61
1540	129,957	937	1.30	0.42	0.70	47,701	21	0.10	0.08	0.16	—	177,658	691	0.98	0.33	0.55
1515	150,791	977	1.05	0.59	0.87	65,864	45	0.15	0.07	0.12	—	216,655	694	0.77	0.43	0.65
1490	197,330	1,080	0.73	0.60	0.82	109,197	55	0.05	0.09	0.16	—	306,527	715	0.49	0.42	0.58
1465	246,093	1,025	0.41	0.57	0.83	96,277	64	0.10	0.16	0.28	—	342,370	755	0.32	0.45	0.68
1440	204,400	882	0.31	0.56	0.85	69,595	43	0.09	0.14	0.26	—	273,995	669	0.25	0.45	0.70
1415	229,385	825	0.37	0.63	0.91	165,536	35	0.10	0.08	0.16	8,301	403,222	484	0.25	0.39	0.59
1390	391,958	752	0.38	0.53	0.82	223,818	34	0.12	0.08	0.16	15,271	631,047	479	0.28	0.36	0.57
1365	750,400	566	0.30	0.46	0.78	480,258	37	0.10	0.12	0.25	30,722	1,261,380	351	0.22	0.32	0.56
1340	1,260,793	567	0.27	0.59	1.00	746,941	41	0.09	0.12	0.23	56,545	2,064,279	361	0.20	0.40	0.69
1315	1,652,519	594	0.32	0.70	1.16	816,221	39	0.10	0.14	0.24	69,394	2,538,134	399	0.24	0.50	0.84
1290	1,640,177	557	0.31	0.64	1.13	657,432	37	0.09	0.11	0.22	65,957	2,363,566	397	0.24	0.48	0.85
1265	1,376,320	486	0.33	0.63	1.07	571,818	32	0.09	0.10	0.21	56,897	2,005,035	343	0.26	0.46	0.79
1240	1,635,602	548	0.40	1.01	1.73	608,274	41	0.11	0.13	0.26	53,174	2,297,050	401	0.31	0.76	1.30
1215	1,288,043	500	0.43	1.22	2.03	551,088	36	0.10	0.15	0.29	49,234	1,888,365	351	0.32	0.87	1.47
1190	1,102,109	451	0.37	0.77	1.33	546,275	30	0.10	0.19	0.35	36,103	1,684,487	305	0.28	0.56	0.98
1165	1,016,866	424	0.34	0.61	1.10	429,351	40	0.10	0.16	0.32	13,138	1,459,355	308	0.26	0.48	0.86
1140	915,895	377	0.34	0.71	1.46	394,683	36	0.09	0.14	0.27	12,637	1,323,215	272	0.26	0.53	1.09
1115	756,878	369	0.46	0.85	1.34	268,086	36	0.11	0.18	0.33	14,734	1,039,698	278	0.36	0.67	1.06
1090	679,398	389	0.60	1.24	2.22	153,359	43	0.15	0.19	0.33	12,813	845,570	320	0.51	1.03	1.84
1065	664,114	392	0.71	1.28	2.61	191,338	30	0.14	0.23	0.42	14,414	869,866	306	0.58	1.02	2.09
1040	682,608	381	0.72	1.16	2.02	224,736	21	0.15	0.31	0.54	15,546	922,890	287	0.57	0.93	1.63
1015	647,275	363	0.58	1.05	1.79	292,860	45	0.15	0.37	0.65	17,770	957,905	259	0.44	0.82	1.41
990	655,754	368	0.49	0.82	1.52	221,644	43	0.17	0.38	0.65	14,214	891,612	281	0.40	0.70	1.28
965	603,504	358	0.49	0.79	1.45	261,556	36	0.15	0.31	0.56	14,862	879,922	256	0.38	0.63	1.16
940	546,049	343	0.48	1.12	1.82	226,459	19	0.14	0.54	0.93	11,639	784,147	245	0.37	0.94	1.54
915	486,446	357	0.45	1.29	2.09	196,854	25	0.13	0.62	0.86	9,849	693,149	258	0.35	1.08	1.71
890	384,610	333	0.42	1.53	2.66	135,405	18	0.17	0.64	1.11	7,534	527,549	247	0.35	1.28	2.22
865	318,064	328	0.31	1.57	2.84	127,129	11	0.16	0.53	0.96	8,055	453,248	233	0.26	1.25	2.26
840	229,420	331	0.34	2.21	3.75	127,338	25	0.14	0.59	1.06	5,889	362,647	218	0.27	1.61	2.75
815	143,685	312	0.49	3.19	4.81	44,059	26	0.14	0.64	1.36	—	187,744	245	0.41	2.59	4.00
790	105,306	300	0.63	5.76	6.62	33,476	23	0.12	0.84	1.85	—	138,782	233	0.51	4.57	5.47
765	102,264	282	0.56	5.11	3.35	65,818	20	0.14	0.98	1.65	—	168,082	180	0.39	3.49	2.69
740	85,347	171	0.40	5.12	4.96	38,262	19	0.11	0.35	0.55	—	123,609	124	0.31	3.65	3.59
715	74,002	150	0.36	6.83	7.32	12,266	29	0.15	0.45	0.73	—	86,268	133	0.33	5.92	6.38
690	48,794	148	0.29	8.71	8.31	19,376	43	0.14	0.73	1.14	—	68,170	118	0.25	6.44	6.27
665	21,799	213	0.36	6.37	5.00	18,905	24	0.09	0.98	1.21	—	40,704	126	0.24	3.87	3.24
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	21,541,624	491	0.44	0.98	1.60	9,276,827	36	0.11	0.21	0.38	614,690	31,433,141	347	0.33	0.74	1.21

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16MINING METHODS

16.1DEVELOPMENT AND PRODUCTION

Active mining areas in the Escobal mine are accessed through two main portals, called the East and West portals. These two primary declines provide access to the Central Zone. A third primary ramp is being driven into the East Zone from the Central Zone. Access ramps are driven from the main ramp system to establish sublevel footwall laterals driven parallel to the vein in transverse mining areas on 25 m vertical intervals, with stopes accessed from the footwall laterals. In longitudinal mining area, development is done on vein also on 25 m vertical intervals.

The access ramps are located nominally 75 to 150 m from the vein. There are also accesses leading to ventilation ingress and exhaust raises. Ventilation raises and ore passes are strategically located throughout the mine and are included in the pre-production development schedules.

The primary development ramps are designed to be 5 m wide by 6 m high with an arched back and are typically driven at a maximum grade of -15%. Secondary development headings allowing access to the individual stopes are designed 5 m wide by 5 m high.

Production from the Escobal mine is achieved via longhole stoping methods. Two variations of this mining method are utilized. Where the vein dimension across the strike is less than 15 m, longitudinal longhole stoping is applied. This method consists of driving horizontal drifts, spaced 25 m vertically, along the strike of the vein and then blasting the ore vertically from the upper level (“over-cut”) to the lower level (“under-cut”).

Breaking slots are established at the extreme ends of the stope to provide a void space for production blasting. These breaking slots are excavated utilizing Cubex drills equipped with V-30 blind bore reaming heads to bore a 30-inch diameter raise between the upper-cut and under-cut for each stope. Once the breaking slots are complete, the stope faces retreat towards the accesses by drilling holes between the over-cut and under-cut, charging the holes with explosives, and blasting a ring or row of holes at the end of the stope. The broken material blasted from the end of the stope is excavated from the under-cut with Caterpillar R1700 or R2900 load-haul-dump (LHD) machines equipped for remote operations. The ore is then loaded into Caterpillar AD45 trucks and transported to the surface.

This process continues until the maximum hydraulic radius or design limit of the opening along strike is reached, at which time longhole mining ceases and the void is filled with paste backfill.

Filtered tails from the process plant are combined with cement and water to make a structural fill for use underground. A paste backfill plant located on the surface produces backfill for delivery via piping into the mine for placement in the mined out stopes. Backfill is required for all stopes for stability reasons and as a preferred place to store tailings. Ore produced from the stope is hauled to the surface ore stockpile by truck and development waste is trucked to the surface and used as construction material for the tailings dry stack buttress.

Once the stope is backfilled and the fill cured, a new breaking slot is required to continue longhole mining in the stope. This process continues until the entire strike length is mined and filled. Excavation lengths along strike, open prior to backfilling, vary depending on the Rock Mass Rating (RMR) of the hanging wall. In areas where the RMR of the vein does not allow excavation of the entire width of the vein in one pass, two or more panels are utilized across the dip to complete excavation of the entire vein. Mining can progress vertically once mining has been completed on the level below, the stope has been backfilled and the fill allowed sufficient curing time. If mining has already taken place below, the stope can be filled with lower strength fill and or waste rock. Figure 16-1 shows the generalized mining sequence for the longitudinal mining method.

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Figure 16-1: Generalized Longitudinal Longhole Stoping Method

In areas where the horizontal vein width exceeds 15 m, measured perpendicular to the strike of the vein, stopes are developed perpendicular to the strike of the vein. This is commonly known as transverse longhole stoping. In this case, 5 m wide by 5 m high footwall laterals are developed approximately 20 m to the south of and parallel to the vein. Access to the footwall laterals is from the primary declines. Five (5) m wide by 5 m high over-cut and under-cut drifts are developed from the south side of the vein to the north side of the vein spaced 25 m vertically. The south side of the vein is typically the footwall but due to local variations in vein geometry can sometimes become the hanging wall.

In the same manner as the longitudinal stopes, breaking slots are established at the extreme ends of the stope to provide a void space for production blasting. These breaking slots are excavated utilizing Cubex drills equipped with V-30 blind bore reaming heads to bore a 30-inch diameter raise between the upper-cut and under-cut for each stope. Once the breaking slots are complete, the stope faces retreat towards the accesses by drilling holes between the over-cut and under-cut, charging the holes with explosives, and blasting a ring or row of holes at the end of the stope. The broken material blasted from the end of the stope is excavated from the under-cut with Caterpillar R1700 or R2900 LHD machines equipped for remote operations. The material is then loaded into Caterpillar AD45 trucks and transported to the process plant. This process continues until all of the material between the hanging wall and the footwall has been excavated at which time hole mining ceases and the void is filled with paste backfill.

The transverse mining method allows for multiple stopes to be in production along strike simultaneously on any given sublevel. Stopes along strike are split into primary stopes and secondary stopes. Each primary stope is separated along strike by a secondary stope. This allows for a rock pillar to be maintained between the primary stopes while these stopes are being excavated increasing the overall stability of the stopes. Once two primary stopes are excavated, backfilled and the backfill allowed to cure, the secondary stope between them can be excavated and subsequently filled with either lower strength fill or waste rock or a combination. Mining progresses from the lower level to the next level above as the stopes on the lower level are mined and backfilled. The over-cut from the lower level becomes the

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undercut or mucking level for the next level above. Figure 16-2 shows the generalized mining sequence for the transverse longhole stoping mining method as applied at the Escobal mine.

Figure 16-2: Generalized Transverse Longhole Stoping

16.2CURRENT STATUS

Underground development of the Escobal mine commenced in May 2011, with construction of the East Central and West Central decline portals; after which ramp development began. Figure 16-3 and Figure 16-4 depict the East Central and West Central Portals, respectively. Since May 2011, approximately 17,140 meters of horizontal development have been driven, averaging nearly 15 meters per day. An additional 229 m of vertical development were accomplished during this time period. Table 16-1 summarizes the development undertaken during the period between May 2011 and July 2014.

During this same time period, the mine produced 786,551 tonnes of ore grading 566 g/t silver, 0.46 g/t gold, 1.01% lead and 1.39% zinc. Table 16-2 summarizes the quantities of ore mined during each month between May 2011 and July 2014.

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Figure 16-3: East Central Decline Portal Area

Figure 16-4: West Central Decline Portal Area

Table 16-1: Summary of Development Completed (as of July 1, 2014)

Development Type	Length
Primary Ramp	3,474 m
Auxiliary & Services Headings	1,466 m
Accesses to Footwall Laterals	703 m
Footwall Laterals	3,430 m
Stope Accesses	8,069 m
Subtotal — Horizontal Development	17,141 m
Raises	229 m
Total of Development Lengths	17,370 m

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Table 16-2: Summary of Ore Production (as of July 1st, 2014)

Production	Ore	Ag	Au	Pb	Zn	Ag	Au	Pb	Zn
Period	Tonnes	(g/t)	(g/t)	(%)	(%)	Ounces	Ounces	Tonnes	Tonnes
Oct 2013	33,725	393	0.45	0.51	0.72	425,760	488	172	243
Nov 2013	55,485	492	0.59	0.65	0.92	876,813	1,054	360	512
Dec 2013	70,002	521	0.54	0.74	1.09	1,171,499	1,209	517	766
Jan 2014	89,078	638	0.46	1.55	1.98	1,827,589	1,323	1,382	1,764
Feb 2014	84,876	455	0.37	0.68	0.88	1,241,226	998	581	747
Mar 2014	99,739	554	0.54	0.71	1.21	1,776,567	1,732	708	1,207
Apr 2014	98,261	645	0.48	1.07	1.78	2,037,374	1,518	1,055	1,747
May 2014	101,575	711	0.47	1.24	1.70	2,322,350	1,543	1,261	1,722
Jun 2014	111,256	618	0.39	1.34	1.51	2,209,458	1,379	1,492	1,676
Stockpile	42,555	303	0.24	1.00	1.22	414,557	328	426	518
Total Production	786,551	566	0.46	1.01	1.39	14,303,192	11,571	7,954	10,903

Figure 16-5 through Figure 16-12 show the current status of the underground development (brown) and production stopes (red) for each active sublevel in the mine. Note that the 1265, 1290 and 1315 m elevations are where most of the production occurred prior to July 1, 2014.

Figure 16-5: As-built Plan Map — 1365 to 1415 m Levels

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Figure 16-6: As-built Plan Map — 1340 m Level

Figure 16-7: As-built Plan Map — 1315 m Level

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Figure 16-8: As-built Plan Map — 1290 m Level

Figure 16-9: As-built Plan Map — 1265 m Level

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Figure 16-10: As-built Plan Map — 1240 m Level

Figure 16-11: As-built Plan Map — 1215 m Level

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Figure 16-12: As-built Plan Map — 1190 m Level

16.3GEOTECHNICAL CONSIDERATIONS

The purpose and scope of this section is to briefly describe conditions governing rock mass stability at the Escobal mine. The section has been divided into descriptions of rock mass quality, considerations for geologic structures, in-situ and mining-induced rock stresses, probabilistic stope stability analysis, installed ground support, backfill and ground support QA/QC.

Previous geotechnical work has been conducted by Rimas Pakalnis Ph.D. (2010, 2011-2012), Christian Caceras PhD. (2013), Langston and Associates (2013-2014), and Golder Associates (2014). Underground geotechnical consulting is currently being conducted by Langston and Associates and SRK Consulting.

Stability of underground excavations is primarily controlled by rock mass quality, orientation of geologic structures, and interaction with the local and regional stress field. Interaction with the stress field aggravates any poor rock mass conditions or adversely-oriented geologic structure that may be encountered.

16.3.1Rock Mass Rating

Rock mass quality data for the Escobal mine have been obtained from geotechnical logging of definition drill core and from direct assessments of the underground headings. Rock mass quality is typically reported from the definition diamond drill hole database in terms of Rock Mass Rating (RMR76) developed by Bieniawski (1976).

Table 16-3 provides a brief summary of the rock mass ratings derived from geotechnical logging of the definition drilling core samples. The minimum, average, and maximum RMR values for all intervals in the diamond drill core database are 0, 40, and 79 respectively.

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Table 16-3: Summary of Drill Core RMR Values

	RMR Values
Location/ Material Type	P25	P50	P75
Waste	31	40	49
Footwall	32	41	52
Ore	33	41	51
Hanging wall	30	36	45

Results reported by the diamond drill data show that the variations in rock mass quality encompass the entire range of geomechanical classifications from I (Good Ground) to IV (Extremely Poor Ground) as defined by the Rock Mass Rating method.

Escobal technical staff assesses the rock mass quality in the mine development headings as mining progresses. Thus far, RMR values obtained during these underground observations generally ranges from 22 to 44 with an average value of 30.

Drill core is typically logged on intervals defined by the drill rod lengths or at major geologic subdivisions. The drill core-based RMR assessments are made on limited and discrete volumes of rock over a constant length, potentially introducing bias to the data. In contrast, assessments of the development headings are generally based on an entire heading within a defined area, resulting in a larger volume and a more generalized RMR value.

The nature of the data collection can result in a bias, creating a differential between the RMR values from drill core and the results found during assessments of the underground excavations.

16.3.2Geologic Structure

Geologic structure interacting with the underground excavations plays a very important role in the stability of all headings observed. Root cause analysis of several problematic areas has shown a direct correlation between the size and orientation of geologic structures and excavation instability.

Underground geologic mapping shows a predominance of northeast trending southeast dipping geologic structures and north trending features that dip moderately to steeply east. Additional structural trends strike E-W and dip north/south. This combination of orientations has the potential to form kinematic wedge failures in headings with azimuths of 270° and 000°.

16.3.3Rock Strength Testing

The unconfined compressive rock strength (UCS) and triaxial rock strength of intact cores have been determined for andesite, capas rojas, and vein/breccia. UCS values ranged from 23 to 69 MPa with an average of 49 MPa over all rock types tested.

Triaxial strength tests were conducted using confining pressures of 3, 7, and 12 MPa. Eight test specimens were evaluated from the capas rojas unit and six from the vein/breccia sequence. The average strength values derived from that analysis were 41 and 52 MPa respectively.

16.3.4Stress Regime

The main mining areas are presently under approximately 200 to 400 m of cover or overburden. Vertical stress at this depth is estimated at approximately 5 to 11 MPa. According to relationships of rock elastic moduli and depth

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determined by Sheorey (1994), the horizontal to vertical stress ratio could range from 1.5 to 2.0, making the in-plane horizontal stress from 7.5 to 22 MPa. Mining an opening in this stress field would then result in an induced stress ranging from 15 to 65 MPa. Given the UCS of the rock and rock mass, the upper end induced stress would be sufficient to cause significant fracturing and deformation in the rock mass around the excavations.

Preliminary FLAC three-dimensional finite difference numerical modeling had been conducted by SRK on the current 15-month mine plan and two proposed mining sequences in the 1190 to 1265 elevation sill pillar. Their results show that induced compressive stress in the rock mass immediately adjacent to open stopes in the early or as-built mining stages are about 15 to 20 MPa which matches relatively well to observed conditions in and around stopes extracted since 2013.

Simple close-form calculations and numerical modeling results show that despite the relatively shallow depth, in situ and mining induced stress could be relatively high. Current observation of the primary stopes indicates a significant influence of the stress field on the rock mass as shown by the presence of deformation and stress induced fracturing and elevated over break into the secondary stopes.

16.3.5             Probabilistic Stope Stability Analysis

The initial transverse stope layout was 20 m wide along strike. Level spacing currently is 25 m which ultimately gives a 30 m high opening between the upper and lower stope accesses. This configuration resulted in a 20 m wide by 30 m vertical high wall on the hanging wall, which would actually be approximately 33 m high (long) with a dip of 65°. Back span in the open stope is then also 20 m. The long dimension of the stope perpendicular to the strike of the orebody is variable and dependent on the width of the ore.

Modified stability graph analyses with a probabilistic basis have been conducted over the course of initial mining to help calibrate optimal stope dimensions. In the analysis, stope wall dimension is plotted against the rock mass quality/stability factor. By comparing cases from other mines, a probability of failure has been ascribed to the point where the stope wall shape factor and rock quality intersect.

The analysis revealed that for the open stope back in the poorest conditions encountered, the probability of failure exceeded 100%. When the back span was 13.3 m for this case the probability of failure became 40% and by the time the back span was reduced to 10 m, the probability of failure decreased to 8%. Analysis of the stope hanging walls showed that the probability of failure for a 20 m span was 70% or very high to certain but dropped to 9% for a 13.3 m span. Reducing the hanging wall span to 10 m decreased the probability of failure to 5%.

Having a probability of failure greater than 10% or one chance in 10 was thought to be too high. Therefore, the decision was made to reduce the stope dimension to 10 m so that the probability of failure in both the open stope back and hanging wall was less than 10%. Additionally, application of the 8 m long cable bolt support was sought to be more effective in the narrower span.

16.3.6             Installed Ground Support

Waste development headings such as footwall laterals and ramps have a nominal designed dimension of 5 m wide by 6 m high with a 2.5 m radius arch in the back. Stope access headings are designed to be 5 m wide by 5 m high with a similar arch in the back. The mine ground support standard is applied to both waste development and stope access headings. Specialized excavations such as vertical ventilation raises have ground support design applied on a case-by-case basis.

Ground support pattern basically consists of 2.4 m standard Swellex bolts on 1.2 m centers with 6 gauge 100 mm aperture welded wire mesh panels as surface support. As ground conditions deteriorate, a 50 — 75 mm shotcrete layer and spiling are incorporated into the pattern.

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Secondary ground support in the form of 3.6 m super Swellex bolts may be added as needed in the intersections of development headings on a 1.5 to 1.8 m spacing. Eight meter long 15 mm diameter double strand cable bolts are installed as secondary support in the stope accesses on a nominal 1.5 m spacing

16.3.7             Ground Support QA/QC

Quality assurance and quality control of ground support elements are systematically tested at least once per quarter. In 2014 the UCS of shotcrete cores has yielded an average of 27 Mpa from a range of 12 to 38 MPa. Pull tests to determine bond strength of standard Swellex rock bolts from 2014 average 9.3 tons from a range of 0 to 14 tons.

16.3.8             Backfill

Cemented paste backfill is placed in the open primary stopes after mining. Cement addition is typically on the order of 6 to 15% depending on the application. Recent strength testing of cemented paste backfill showed an average strength of 0.4 MPa for binder contents of 10 to 15% and 80 test age of 20 to 27 days. More samples with the same binder content and a test age of 28 days or more, the average compressive strength was 0.7 MPa.

Required cemented paste backfill compressive strength for an open stope side wall of 30 by 30 m is 0.4 MPa. The compressive strength required for underhand mining is quite variable and depends greatly on the back span of the opening.

16.4                PUMPING

Drilling and underground operations to date have encountered varying quantities of ground water, in some cases in excess of 500 gallons per minute (gpm) over short periods of time. Overall, the mine currently generates between 400 and 600 gpm of ground water and pumping is required to transport this water to the surface. The pumping system utilizes small pumps in the development headings and stopes to transport water to a central location. Permanent sumps and pump stations have been constructed below the active production levels. From the main sump, water is pumped to the mine water thickener on the surface where the thickener overflow reports to the process water pond. Where possible, clean water (i.e., water not ‘impacted’ by mining activities) intercepted underground is collected and pumped in a separate system and delivered to holding ponds on the surface where the pH is continually monitored and adjusted prior to discharge.

Permanent mine pumping stations will be able to handle approximately 2,400 gpm of ‘dirty’ mine water. There will be five main pumping station locations, three in the Central Zone and two in the East Zone. The stations will be placed approximately 200 vertical meters apart. It is expected that a recharge rate of 1,200 gpm will require the system to operate 50% of the time. Pump stations are added as the mine expands. Mine water is collected on each level into a sump. The sumps then drain to pipes which feed to the closest pump station. Discharge out of the mine is via a 10” discharge line. Temporary pump skids are used in the declines until the next pump station is constructed.

The permanent pumping system, once activated, will operate at a relatively high flow rate so that slimes do not settle in the discharge pipes. This design also prevents excessive wear on the 10” diameter discharge pipe. The pumps and sumps are sized such that the system will operate for a period of 30 minutes, approximately 24 times per day.

The design intent of the system is to allow handling of the slimes to occur at the mine water thickener rather than in the underground operations. Transporting the slimes out of the mine via the ramps is not as efficient as pumping the slimes out of the mine. Larger material that finds its way into the mine drainage system will be trapped at the smaller production level sumps. These sumps are designed to be cleaned out with a loader and will be maintained for the life of mine. The pump station sumps include submersible pumps that will be used to agitate the water during the 30 minute recharge time.

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16.5                VENTILATION

The overall ventilation system design in operation at Escobal is a negative or “pull” system with supply air delivered via the two main declines and a fresh air raise to be located in the East Zone. A single return air raise (6 m diameter) is located at the center of the Central Zone with the primary return air fans located on surface. Two 1,250-hp fans (one currently installed) sit atop of the return air raise and pull approximately 500 m3/s of air out of the mine. Fresh air is then drawn from the ramps onto the production levels where it sweeps the working areas and then exhausts out raises located at the end of the levels. The exhaust raises are located at the ends in order to establish flow through ventilation which is used as a source of air for smaller auxiliary fans that push air into the work areas via ventilation ducting.

The estimated air volume requirements are 350 m³/s for the Central Zone and 150 m³/s for the East Zone. The overall ventilation air flow distribution will vary throughout the mine life with the completion of the East Zone and start-up of the lower Central Zone. A generalized view of the surface fan layout is shown in Figure 16-13 and the schematic of the overall ventilation network is shown in Figure 16-14.

Figure 16-13: Generalized Surface Fan Installation

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Figure 16-14: Generalized Ventilation Network

16.6                STOPE DESIGN

Practical mining shapes for each stope were created by first extracting the centroids of resource model blocks having a net value greater than $0/tonne. The ore-grade blocks located within the extents of a proposed stope were used as inputs to a plane-fitting algorithm which determined the hanging wall (north side) and footwall (south side) faces for the stope. The extent of each stope is generally 25 m high by 10 m wide (east-west direction). The extent in the north-south varies according to the geometry of the modeled mineralization.

While accurately mimicking the local geometry of the block model, the plane-fitting algorithm by itself does not always result in a practical mining shape. For example, the mineralization may have a local foot wall dip angle less than 45° from the horizontal, an angle generally unachievable with the mining method in use at Escobal. Therefore, the plane-fitting routine included a second pass to modify the original planes to more practical shapes. In general, the transverse stopes were allowed less variability than the longitudinal stopes. Table 16-4 below lists the input parameters for the stope model creation process.

Table 16-4: Stope Shape Parameters

Stope Design Parameter	Transverse Stopes	Longitudinal Stopes
Height	25 m	25 m
Width	10 m	varies
Length	varies	10 m
Minimum number of points to form plane	75 pts	75 pts
Footwall dip range (south face)	-90° to -45°	±30° from vertical
Hanging wall dip range (north face)	±10° from vertical	±30° from vertical
Strike range (horizontal plane)	70° to 110°	55° to 125°
Minimum volume constraint	2500m3	750m3

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Figure 16-15: Transverse Stope Shape Parameters

Figure 16-16: Longitudinal Stope Shape Parameters

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The preliminary stope shapes were individually refined where necessary to ensure stope geometry viability and to minimize the amount of sub-economic material within the shape volume that is inseparable from profitable material due to the practical constraints of mining. The stopes were also adjusted to account for overbreak from the primary stopes that were excavated prior to July 2014.

The decision to design a proposed stope as a longitudinal or transverse stope was an engineering decision, mainly guided by the local continuity and horizontal width of the mineralization. Each proposed stope was reviewed to assure minimum volume restrictions; specifically, any transverse stope less than 2,500 m3 or longitudinal stope less than 750 m3 was subject to further processing to widen the stope to meet the minimum volume requirements.

After ensuring adherence to the minimum volume restriction, Blattman determined the average grade and tonnage for each stope. This included calculating the net value for the entire stope in terms of $/tonne. Any stope shape with a net value less than $0/tonne was discarded and not included in the final reserve tabulation.

In total, 3,926 stopes were designed including 1,319 transverse stopes and 2,607 longitudinal. Despite outnumbering the transverse stopes by almost 2 to 1, the longitudinal stopes account for only 35% of the total reserve mass. The average transverse stope contains roughly 15,600 tonnes at 350 g Ag/t while the average longitudinal stope contains 4,200 tonnes at 341 g Ag/t.

Figure 16-17 and Figure 16-18 show the complete set of stopes designed for the Escobal mine including transverse and longitudinal stopes.

Figure 16-17: Escobal Reserve Stopes — Looking North

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Figure 16-18: Escobal Reserve Stopes — Oblique View

16.7                DILUTION AND RECOVERY ESTIMATES

Ore dilution in the secondary stopes includes 0.5 m of overbreak (1.0 m total per stope) into the adjacent backfilled stopes, adding paste fill to the mineralized rock in the stope. This effectively increases the total mass of material extracted from the stope and reduces the overall grade. In addition, the mass of any Inferred resources contained within the stope design are assumed to carry no metal values in the estimation of Mineral Reserve grades. In the evaluation of the Mineral Reserves, no additional modifying factors were applied to the tonnages and grade of the mining shapes to account for mining losses.

Table 16-5 shows the ore-loss calculation due to the local geometry variability in relation to the stope plane-fitting process. In general, the current mine design extracts approximately 93% of all Measured and Indicated resources above cutoff grade and within the immediate area of the proposed mine design. The stopes tend to incur nearly 7% ore losses on the hanging wall and footwall surfaces. Table 16-6 and Table 16-7 breakdown the same information by stope type. Generally the longitudinal stopes have a lower ore-loss factor (~6%) than the transverse stopes (~7.5%) but the overall rates are not dissimilar.

Table 16-8 shows the accounting by level of Measured and Indicated resources above cutoff grade and waste rock and paste fill dilution included in the stope designs. Table 16-9 provides similar information for the transverse stopes only and Table 16-10 for the longitudinal stopes. Note minor differences between the various tables may be present and are due to reporting precision and do not represent material errors in the estimate.

In general, the dilution rate for the overall reserve is approximately 31%. The average dilution rate for the transverse stopes is 35% while the longitudinal stopes average 26%. Though not quantifiable, Blattman acknowledges that a portion of the reported dilution results from applying realistic stope geometries to resource model blocks with dimensions of 5m x 5m x 2.5m that often create coarse ‘stair-step’ boundaries of ore grade blocks along the footwall and hanging wall of the block model that do not accurately portray the true ore boundaries. The actual dilution when mining may be lower than reported in this Study. Future resource modeling efforts should incorporate methods to better depict grade domain boundaries in the resulting model blocks to allow for a more accurate estimate of mining dilution.

132

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 16-5: Mineral Resources in Relation to the Mine Design

	Available M&I Resources Above Cutoff					M&I Resources Inside Stope Shapes					Ore Losses
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	11,089	319	3.46	0.37	0.46	10,411	323	3.58	0.39	0.48	678	263	1.65	0.15	0.20
1590	43,547	510	5.06	0.46	0.51	39,742	526	5.21	0.47	0.52	3,805	351	3.54	0.34	0.41
1565	73,536	871	2.13	0.73	0.68	67,516	921	2.21	0.76	0.71	6,020	312	1.16	0.29	0.37
1540	142,589	898	1.23	0.41	0.67	129,957	937	1.30	0.42	0.70	12,632	494	0.48	0.25	0.37
1515	166,355	922	1.02	0.56	0.83	150,791	977	1.05	0.59	0.87	15,564	386	0.72	0.22	0.38
1490	216,067	1,024	0.70	0.57	0.78	197,330	1,080	0.73	0.60	0.82	18,737	429	0.33	0.22	0.34
1465	264,313	990	0.40	0.55	0.81	246,093	1,025	0.41	0.57	0.83	18,220	521	0.31	0.27	0.47
1440	228,594	867	0.31	0.53	0.82	204,400	882	0.31	0.56	0.85	24,194	735	0.34	0.34	0.53
1415	285,006	764	0.35	0.58	0.85	229,385	825	0.37	0.63	0.91	55,621	513	0.27	0.39	0.59
1390	437,839	735	0.37	0.52	0.82	391,958	752	0.38	0.53	0.82	45,881	589	0.27	0.51	0.81
1365	849,728	553	0.29	0.45	0.76	750,400	566	0.30	0.46	0.78	99,328	451	0.22	0.39	0.63
1340	1,405,740	562	0.27	0.58	0.99	1,260,793	567	0.27	0.59	1.00	144,947	520	0.23	0.56	0.97
1315	1,772,449	589	0.32	0.69	1.15	1,652,519	594	0.32	0.70	1.16	119,930	512	0.29	0.57	0.98
1290	1,790,611	558	0.31	0.64	1.12	1,640,177	557	0.31	0.64	1.13	150,434	567	0.33	0.60	0.99
1265	1,554,641	485	0.34	0.63	1.07	1,376,320	486	0.33	0.63	1.07	178,321	478	0.37	0.68	1.09
1240	1,732,567	538	0.39	0.98	1.67	1,635,602	548	0.40	1.01	1.73	96,965	365	0.21	0.34	0.69
1215	1,370,294	493	0.43	1.18	1.96	1,288,043	500	0.43	1.22	2.03	82,251	400	0.42	0.56	0.90
1190	1,185,467	445	0.36	0.74	1.29	1,102,109	451	0.37	0.77	1.33	83,358	369	0.23	0.43	0.80
1165	1,056,454	421	0.33	0.60	1.08	1,016,866	424	0.34	0.61	1.10	39,588	341	0.22	0.36	0.64
1140	948,530	375	0.34	0.70	1.43	915,895	377	0.34	0.71	1.46	32,635	320	0.29	0.44	0.86
1115	788,651	366	0.45	0.83	1.32	756,878	369	0.46	0.85	1.34	31,773	299	0.36	0.37	0.73
1090	701,409	387	0.60	1.22	2.18	679,398	389	0.60	1.24	2.22	22,011	311	0.51	0.46	0.96
1065	692,259	389	0.70	1.25	2.55	664,114	392	0.71	1.28	2.61	28,145	319	0.41	0.55	1.15
1040	709,360	379	0.71	1.14	2.00	682,608	381	0.72	1.16	2.02	26,752	321	0.44	0.76	1.55
1015	676,114	360	0.57	1.03	1.76	647,275	363	0.58	1.05	1.79	28,839	305	0.35	0.65	1.22
990	684,672	366	0.48	0.81	1.51	655,754	368	0.49	0.82	1.52	28,918	313	0.37	0.68	1.22
965	636,840	356	0.49	0.79	1.44	603,504	358	0.49	0.79	1.45	33,336	329	0.49	0.80	1.27
940	577,403	341	0.47	1.11	1.80	546,049	343	0.48	1.12	1.82	31,354	304	0.37	0.96	1.50
915	508,839	355	0.44	1.27	2.09	486,446	357	0.45	1.29	2.09	22,393	329	0.38	1.05	2.19
890	404,702	328	0.41	1.53	2.68	384,610	333	0.42	1.53	2.66	20,092	225	0.26	1.65	3.03
865	329,980	328	0.31	1.55	2.81	318,064	328	0.31	1.57	2.84	11,916	354	0.37	1.27	2.37
840	245,386	329	0.34	2.21	3.73	229,420	331	0.34	2.21	3.75	15,966	300	0.32	2.13	3.35
815	155,826	310	0.49	3.08	4.70	143,685	312	0.49	3.19	4.81	12,141	293	0.45	1.84	3.51
790	114,261	297	0.62	5.50	6.53	105,306	300	0.63	5.76	6.62	8,955	260	0.52	2.44	5.40
765	115,159	276	0.56	5.00	3.32	102,264	282	0.56	5.11	3.35	12,895	221	0.59	4.13	3.04
740	91,724	173	0.42	5.15	4.89	85,347	171	0.40	5.12	4.96	6,377	200	0.60	5.57	4.02
715	81,150	152	0.37	6.75	7.16	74,002	150	0.36	6.83	7.32	7,148	169	0.40	5.93	5.58
690	52,108	147	0.29	8.64	8.23	48,794	148	0.29	8.71	8.31	3,314	129	0.20	7.57	7.10
665	24,161	213	0.35	6.24	4.93	21,799	213	0.36	6.37	5.00	2,362	211	0.25	5.04	4.27
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	23,125,420	487	0.43	0.96	1.57	21,541,624	491	0.44	0.98	1.60	1,583,796	434	0.33	0.68	1.10

133

ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 16-6: Mineral Resources in Relation to the Mine Design (Transverse Stopes Only)

	Available M&I Resources Above Cutoff					M&I Resources Inside Stope Shapes					Ore Losses
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1590	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1565	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1540	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1515	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1490	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1465	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1440	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1415	206,082	724	0.30	0.54	0.73	159,591	789	0.31	0.59	0.79	46,491	499	0.25	0.36	0.52
1390	352,870	741	0.37	0.52	0.79	312,176	765	0.38	0.52	0.79	40,694	554	0.26	0.49	0.78
1365	646,073	523	0.26	0.42	0.73	566,777	537	0.27	0.43	0.74	79,296	428	0.20	0.35	0.63
1340	1,262,364	556	0.27	0.56	0.99	1,129,027	560	0.27	0.56	0.99	133,337	524	0.23	0.55	0.96
1315	1,532,553	597	0.32	0.71	1.19	1,435,513	600	0.33	0.71	1.20	97,040	558	0.30	0.63	1.09
1290	1,570,727	583	0.32	0.68	1.18	1,438,207	581	0.32	0.68	1.19	132,520	601	0.35	0.65	1.06
1265	1,344,856	511	0.37	0.69	1.16	1,183,034	513	0.37	0.69	1.16	161,822	495	0.39	0.72	1.16
1240	1,381,583	599	0.45	1.15	1.95	1,305,568	613	0.46	1.19	2.02	76,015	374	0.22	0.37	0.74
1215	1,221,994	516	0.45	1.28	2.12	1,150,178	522	0.46	1.32	2.19	71,816	410	0.42	0.59	0.92
1190	842,030	460	0.35	0.76	1.43	783,758	466	0.36	0.79	1.47	58,272	380	0.19	0.45	0.86
1165	326,188	392	0.57	0.97	1.92	316,863	394	0.58	0.99	1.95	9,325	348	0.26	0.33	0.79
1140	338,344	388	0.61	1.34	2.87	321,281	391	0.62	1.39	2.96	17,063	321	0.42	0.55	1.10
1115	398,281	379	0.71	1.25	1.94	381,100	384	0.72	1.29	1.99	17,181	278	0.50	0.43	0.81
1090	364,037	429	0.88	1.87	3.31	353,389	432	0.89	1.90	3.36	10,648	336	0.74	0.64	1.38
1065	373,158	423	0.91	1.82	3.85	357,655	427	0.93	1.87	3.95	15,503	343	0.56	0.72	1.52
1040	384,334	408	0.87	1.62	2.81	370,532	410	0.88	1.64	2.83	13,802	340	0.57	1.14	2.21
1015	390,368	401	0.74	1.41	2.35	376,939	403	0.75	1.42	2.36	13,429	349	0.46	0.99	1.89
990	344,496	378	0.59	1.09	2.00	331,003	380	0.60	1.10	2.00	13,493	331	0.49	0.92	1.77
965	309,903	376	0.59	0.93	1.71	292,113	377	0.58	0.93	1.72	17,790	358	0.67	1.04	1.57
940	261,225	368	0.48	1.56	2.27	244,909	372	0.50	1.57	2.29	16,316	306	0.29	1.36	2.01
915	200,646	417	0.51	2.00	3.00	192,083	420	0.51	2.00	2.97	8,563	348	0.42	1.94	3.72
890	174,615	402	0.43	2.25	3.76	166,508	411	0.44	2.24	3.74	8,107	229	0.24	2.41	4.11
865	180,335	382	0.35	2.30	3.95	174,697	382	0.36	2.31	3.98	5,638	383	0.33	1.94	3.13
840	117,835	376	0.39	2.98	5.02	111,670	374	0.39	3.03	5.10	6,165	412	0.40	2.14	3.60
815	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
790	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
765	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
740	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
715	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
690	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
665	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	14,524,897	514	0.44	1.00	1.74	13,454,571	517	0.45	1.03	1.79	1,070,326	468	0.32	0.63	1.09

134

ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 16-7: Mineral Resources in Relation to the Mine Design (Longitudinal Stopes Only)

	Available M&I Resources Above Cutoff					M&I Resources Inside Stope Shapes					Ore Losses
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	11,089	319	3.46	0.37	0.46	10,411	323	3.58	0.39	0.48	678	263	1.65	0.15	0.20
1590	43,547	510	5.06	0.46	0.51	39,742	526	5.21	0.47	0.52	3,805	351	3.54	0.34	0.41
1565	73,536	871	2.13	0.73	0.68	67,516	921	2.21	0.76	0.71	6,020	312	1.16	0.29	0.37
1540	142,589	898	1.23	0.41	0.67	129,957	937	1.30	0.42	0.70	12,632	494	0.48	0.25	0.37
1515	166,355	922	1.02	0.56	0.83	150,791	977	1.05	0.59	0.87	15,564	386	0.72	0.22	0.38
1490	216,067	1,024	0.70	0.57	0.78	197,330	1,080	0.73	0.60	0.82	18,737	429	0.33	0.22	0.34
1465	264,313	990	0.40	0.55	0.81	246,093	1,025	0.41	0.57	0.83	18,220	521	0.31	0.27	0.47
1440	228,594	867	0.31	0.53	0.82	204,400	882	0.31	0.56	0.85	24,194	735	0.34	0.34	0.53
1415	78,924	870	0.49	0.70	1.17	69,794	907	0.50	0.72	1.20	9,130	587	0.39	0.54	0.94
1390	84,969	710	0.37	0.54	0.93	79,782	700	0.37	0.53	0.93	5,187	865	0.30	0.61	1.00
1365	203,655	647	0.38	0.57	0.88	183,623	658	0.39	0.57	0.91	20,032	543	0.32	0.52	0.63
1340	143,376	611	0.24	0.80	1.04	131,766	623	0.24	0.81	1.04	11,610	473	0.26	0.71	1.10
1315	239,896	533	0.27	0.59	0.89	217,006	556	0.27	0.62	0.93	22,890	318	0.27	0.32	0.53
1290	219,884	379	0.19	0.35	0.69	201,970	384	0.19	0.36	0.71	17,914	313	0.19	0.22	0.46
1265	209,785	323	0.14	0.26	0.51	193,286	324	0.13	0.27	0.52	16,499	311	0.15	0.21	0.45
1240	350,984	296	0.13	0.30	0.59	330,034	293	0.13	0.31	0.59	20,950	332	0.18	0.26	0.50
1215	148,300	308	0.22	0.35	0.66	137,865	309	0.20	0.35	0.65	10,435	328	0.44	0.35	0.75
1190	343,437	408	0.40	0.70	0.96	318,351	414	0.40	0.72	0.98	25,086	343	0.31	0.39	0.67
1165	730,266	434	0.23	0.44	0.71	700,003	438	0.23	0.44	0.71	30,263	339	0.20	0.37	0.59
1140	610,186	368	0.19	0.35	0.64	594,614	369	0.19	0.35	0.64	15,572	319	0.15	0.31	0.61
1115	390,370	353	0.19	0.40	0.68	375,778	354	0.19	0.41	0.68	14,592	323	0.19	0.29	0.63
1090	337,372	341	0.30	0.52	0.97	326,009	343	0.30	0.53	0.98	11,363	288	0.30	0.29	0.56
1065	319,101	350	0.46	0.57	1.03	306,459	352	0.47	0.58	1.05	12,642	291	0.22	0.33	0.70
1040	325,026	345	0.51	0.58	1.06	312,076	346	0.52	0.59	1.06	12,950	301	0.31	0.34	0.84
1015	285,746	305	0.35	0.52	0.97	270,336	307	0.35	0.53	0.99	15,410	267	0.26	0.35	0.63
990	340,176	354	0.37	0.54	1.02	324,751	356	0.38	0.54	1.04	15,425	296	0.26	0.46	0.73
965	326,937	337	0.40	0.65	1.18	311,391	339	0.41	0.66	1.20	15,546	295	0.28	0.52	0.93
940	316,178	318	0.46	0.75	1.41	301,140	320	0.46	0.76	1.44	15,038	301	0.45	0.52	0.94
915	308,193	315	0.40	0.79	1.50	294,363	316	0.40	0.82	1.52	13,830	317	0.35	0.50	1.24
890	230,087	271	0.39	0.99	1.86	218,102	274	0.40	0.98	1.83	11,985	222	0.27	1.14	2.30
865	149,645	263	0.26	0.65	1.44	143,367	262	0.26	0.67	1.45	6,278	329	0.41	0.67	1.69
840	127,551	285	0.30	1.49	2.53	117,750	290	0.30	1.44	2.47	9,801	229	0.28	2.12	3.19
815	155,826	310	0.49	3.08	4.70	143,685	312	0.49	3.19	4.81	12,141	293	0.45	1.84	3.51
790	114,261	297	0.62	5.50	6.53	105,306	300	0.63	5.76	6.62	8,955	260	0.52	2.44	5.40
765	115,159	276	0.56	5.00	3.32	102,264	282	0.56	5.11	3.35	12,895	221	0.59	4.13	3.04
740	91,724	173	0.42	5.15	4.89	85,347	171	0.40	5.12	4.96	6,377	200	0.60	5.57	4.02
715	81,150	152	0.37	6.75	7.16	74,002	150	0.36	6.83	7.32	7,148	169	0.40	5.93	5.58
690	52,108	147	0.29	8.64	8.23	48,794	148	0.29	8.71	8.31	3,314	129	0.20	7.57	7.10
665	24,161	213	0.35	6.24	4.93	21,799	213	0.36	6.37	5.00	2,362	211	0.25	5.04	4.27
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	8,600,523	441	0.41	0.90	1.28	8,087,053	446	0.41	0.90	1.29	513,470	363	0.35	0.79	1.12

135

ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 16-8: Mineral Reserve Estimate by Elevation

	M&I Resources Inside Stope Shapes					Waste Blocks					Paste	P&P Reserves
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	10,411	323	3.58	0.39	0.48	5,455	21	0.30	0.05	0.09	—	15,866	219	2.45	0.27	0.35
1590	39,742	526	5.21	0.47	0.52	14,966	18	0.35	0.05	0.09	—	54,708	387	3 88	0.36	0.40
1565	67,516	921	2.21	0.76	0.71	17,151	37	0.20	0.16	0.22	—	84,667	742	1.80	0.64	0.61
1540	129,957	937	1.30	0.42	0.70	47,701	21	0.10	0.08	0.16	—	177,658	691	0.98	0.33	0.55
1515	150,791	977	1.05	0.59	0.87	65,864	45	0.15	0.07	0.12	—	216,655	694	0.77	0.43	0.65
1490	197,330	1,080	0.73	0.60	0.82	109,197	55	0.05	0.09	0.16	—	306,527	715	0.49	0.42	0.58
1465	246,093	1,025	0.41	0.57	0.83	96,277	64	0.10	0.16	0.28	—	342,370	755	0.32	0.45	0.68
1440	204,400	882	0.31	0.56	0.85	69,595	43	0.09	0.14	0.26	—	273,995	669	0.25	0.45	0.70
1415	229,385	825	0.37	0.63	0.91	165,536	35	0.10	0.08	0.16	8,301	403,222	484	0.25	0.39	0.59
1390	391,958	752	0.38	0.53	0.82	223,818	34	0.12	0.08	0.16	15,271	631,047	479	0.28	0.36	0.57
1365	750,400	566	0.30	0.46	0.78	480,258	37	0.10	0.12	0.25	30,722	1,261,380	351	0.22	0.32	0.56
1340	1,260,793	567	0.27	0.59	1.00	746,941	41	0.09	0.12	0.23	56,545	2,064,279	361	0.20	0.40	0.69
1315	1,652,519	594	0.32	0.70	1.16	816,221	39	0.10	0.14	0.24	69,394	2,538,134	399	0.24	0.50	0.84
1290	1,640,177	557	0.31	0.64	1.13	657,432	37	0.09	0.11	0.22	65,957	2,363,566	397	0.24	0.48	0.85
1265	1,376,320	486	0.33	0.63	1.07	571,818	32	0.09	0.10	0.21	56,897	2,005,035	343	0.26	0.46	0.79
1240	1,635,602	548	0.40	1.01	1.73	608,274	41	0.11	0.13	0.26	53,174	2,297,050	401	0.31	0.76	1.30
1215	1,288,043	500	0.43	1.22	2.03	551,088	36	0.10	0.15	0.29	49,234	1,888,365	351	0.32	0.87	1.47
1190	1,102,109	451	0.37	0.77	1.33	546,275	30	0.10	0.19	0.35	36,103	1,684,487	305	0.28	0.56	0.98
1165	1,016,866	424	0.34	0.61	1.10	429,351	40	0.10	0.16	0.32	13,138	1,459,355	308	0.26	0.48	0.86
1140	915,895	377	0.34	0.71	1.46	394,683	36	0.09	0.14	0.27	12,637	1,323,215	272	0.26	0.53	1.09
1115	756,878	369	0.46	0.85	1.34	268,086	36	0.11	0.18	0.33	14,734	1,039,698	278	0.36	0.67	1.06
1090	679,398	389	0.60	1.24	2.22	153,359	43	0.15	0.19	0.33	12,813	845,570	320	0.51	1.03	1.84
1065	664,114	392	0.71	1.28	2.61	191,338	30	0.14	0.23	0.42	14,414	869,866	306	0.58	1.02	2.09
1040	682,608	381	0.72	1.16	2.02	224,736	21	0.15	0.31	0.54	15,546	922,890	287	0.57	0.93	1.63
1015	647,275	363	0.58	1.05	1.79	292,860	45	0.15	0.37	0.65	17,770	957,905	259	0.44	0.82	1.41
990	655,754	368	0.49	0.82	1.52	221,644	43	0.17	0.38	0.65	14,214	891,612	281	0.40	0.70	1.28
965	603,504	358	0.49	0.79	1.45	261,556	36	0.15	0.31	0.56	14,862	879,922	256	0.38	0.63	1.16
940	546,049	343	0.48	1.12	1.82	226,459	19	0.14	0.54	0.93	11,639	784,147	245	0.37	0.94	1.54
915	486,446	357	0.45	1.29	2.09	196,854	25	0.13	0.62	0.86	9,849	693,149	258	0.35	1.08	1.71
890	384,610	333	0.42	1.53	2.66	135,405	18	0.17	0.64	1.11	7,534	527,549	247	0.35	1.28	2.22
865	318,064	328	0.31	1.57	2.84	127,129	11	0.16	0.53	0.96	8,055	453,248	233	0.26	1.25	2.26
840	229,420	331	0.34	2.21	3.75	127,338	25	0.14	0.59	1.06	5,889	362,647	218	0.27	1.61	2.75
815	143,685	312	0.49	3.19	4.81	44,059	26	0.14	0.64	1.36	—	187,744	245	0.41	2.59	4.00
790	105,306	300	0.63	5.76	6.62	33,476	23	0.12	0.84	1.85	—	138,782	233	0.51	4.57	5.47
765	102,264	282	0.56	5.11	3.35	65,818	20	0.14	0.98	1.65	—	168,082	180	0.39	3.49	2.69
740	85,347	171	0.40	5.12	4.96	38,262	19	0.11	0.35	0.55	—	123,609	124	0.31	3.65	3.59
715	74,002	150	0.36	6.83	7.32	12,266	29	0.15	0.45	0.73	—	86,268	133	0.33	5.92	6.38
690	48,794	148	0.29	8.71	8.31	19,376	43	0.14	0.73	1.14	—	68,170	118	0.25	6.44	6.27
665	21,799	213	0.36	6.37	5.00	18,905	24	0.09	0.98	1.21	—	40,704	126	0.24	3.87	3.24
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	21,541,624	491	0.44	0.98	1.60	9,276,827	36	0.11	0.21	0.38	614,690	31,433,141	347	0.33	0.74	1.21

136

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 16-9: Transverse Stope Reserves by Elevation

	M&I Resources Inside Stope Shapes					Waste Blocks					Paste	P&P Reserves
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1590	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1565	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1540	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1515	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1490	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1465	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1440	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1415	159,591	789	0.31	0.59	0.79	119,749	33	0.10	0.07	0.14	8,301	287,641	451	0.21	0.36	0.50
1390	312,176	765	0.38	0.52	0.79	176,896	33	0.12	0.08	0.14	15,271	504,343	485	0.28	0.35	0.54
1365	566,777	537	0.27	0.43	0.74	443,365	38	0.10	0.12	0.25	30,722	1,040,864	308	0.19	0.29	0.51
1340	1,129,027	560	0.27	0.56	0.99	726,110	41	0.09	0.12	0.23	56,545	1,911,682	346	0.20	0.38	0.67
1315	1,435,513	600	0.33	0.71	1.20	743,995	40	0.11	0.14	0.25	69,394	2,248,902	396	0.24	0.50	0.85
1290	1,438,207	581	0.32	0.68	1.19	573,670	39	0.09	0.11	0.21	65,957	2,077,834	413	0.25	0.50	0.88
1265	1,183,034	513	0.37	0.69	1.16	486,569	34	0.10	0.11	0.22	56,897	1,726,500	361	0.28	0.50	0.86
1240	1,305,568	613	0.46	1.19	2.02	473,362	47	0.13	0.14	0.27	53,174	1,832,104	449	0.36	0.89	1.51
1215	1,150,178	522	0.46	1.32	2.19	510,904	37	0.10	0.15	0.30	49,234	1,710,316	362	0.34	0.93	1.56
1190	783,758	466	0.36	0.79	1.47	424,241	28	0.11	0.20	0.37	36,103	1,244,102	303	0.26	0.56	1.05
1165	316,863	394	0.58	0.99	1.95	141,661	43	0.16	0.20	0.43	13,138	471,662	277	0.44	0.73	1.44
1140	321,281	391	0.62	1.39	2.96	109,239	50	0.17	0.21	0.42	12,637	443,157	296	0.49	1.06	2.25
1115	381,100	384	0.72	1.29	1.99	116,050	46	0.18	0.23	0.45	14,734	511,884	296	0.58	1.01	1.59
1090	353,389	432	0.89	1.90	3.36	79,507	48	0.19	0.20	0.36	12,813	445,709	351	0.74	1.54	2.73
1065	357,655	427	0.93	1.87	3.95	115,320	27	0.15	0.24	0.43	14,414	487,389	319	0.71	1.43	3.00
1040	370,532	410	0.88	1.64	2.83	132,099	17	0.15	0.27	0.45	15,546	518,177	297	0.67	1.24	2.14
1015	376,939	403	0.75	1.42	2.36	211,137	47	0.15	0.35	0.66	17,770	605,846	267	0.52	1.01	1.70
990	331,003	380	0.60	1.10	2.00	154,208	47	0.17	0.39	0.66	14,214	499,425	266	0.45	0.85	1.53
965	292,113	377	0.58	0.93	1.72	203,695	33	0.14	0.30	0.54	14,862	510,670	229	0.39	0.65	1.20
940	244,909	372	0.50	1.57	2.29	152,238	13	0.13	0.59	0.97	11,639	408,786	228	0.35	1.16	1.73
915	192,083	420	0.51	2.00	2.97	139,312	18	0.12	0.71	0.89	9,849	341,244	244	0.34	1.42	2.03
890	166,508	411	0.44	2.24	3.74	88,162	6	0.18	0.72	1.18	7,534	262,204	263	0.34	1.66	2.77
865	174,697	382	0.36	2.31	3.98	93,494	3	0.16	0.59	1.04	8,055	276,246	243	0.28	1.66	2.87
840	111,670	374	0.39	3.03	5.10	91,988	24	0.15	0.66	1.17	5,889	209,547	210	0.27	1.91	3.23
815	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
790	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
765	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
740	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
715	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
690	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
665	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	13,454,571	517	0.45	1.03	1.79	6,506,971	36	0.12	0.21	0.37	614,690	20,576,232	350	0.33	0.74	1.29

137

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 16-10: Longitudinal Stope Reserves by Elevation

	M&I Resources Inside Stope Shapes					Waste Blocks					Paste	P&P Reserves
	Mass	Ag	Au	Pb	Zn	Mass	Ag	Au	Pb	Zn	Mass	Mass	Ag	Au	Pb	Zn
Level	(tonnes)	g/t	g/t	%	%	(tonnes)	g/t	g/t	%	%	(tonnes)	(tonnes)	g/t	g/t	%	%
1640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
1615	10,411	323	3.58	0.39	0.48	5,455	21	0.30	0.05	0.09	—	15,866	219	2.45	0.27	0.35
1590	39,742	526	5.21	0.47	0.52	14,966	18	0.35	0.05	0.09	—	54,708	387	3.88	0.36	0.40
1565	67,516	921	2.21	0.76	0.71	17,151	37	0.20	0.16	0.22	—	84,667	742	1.80	0.64	0.61
1540	129,957	937	1.30	0.42	0.70	47,701	21	0.10	0.08	0.16	—	177,658	691	0.98	0.33	0.55
1515	150,791	977	1.05	0.59	0.87	65,864	45	0.15	0.07	0.12	—	216,655	694	0.77	0.43	0.65
1490	197,330	1,080	0.73	0.60	0.82	109,197	55	0.05	0.09	0.16	—	306,527	715	0.49	0.42	0.58
1465	246,093	1,025	0.41	0.57	0.83	96,277	64	0.10	0.16	0.28	—	342,370	755	0.32	0.45	0.68
1440	204,400	882	0.31	0.56	0.85	69,595	43	0.09	0.14	0.26	—	273,995	669	0.25	0.45	0.70
1415	69,794	907	0.50	0.72	1.20	45,787	42	0.10	0.10	0.22	—	115,581	564	0.34	0.47	0.81
1390	79,782	700	0.37	0.53	0.93	46,922	37	0.09	0.09	0.22	—	126,704	454	0.27	0.37	0.66
1365	183,623	658	0.39	0.57	0.91	36,893	29	0.07	0.13	0.24	—	220,516	553	0.34	0.50	0.80
1340	131,766	623	0.24	0.81	1.04	20,831	30	0.04	0.16	0.29	—	152,597	542	0.21	0.72	0.93
1315	217,006	556	0.27	0.62	0.93	72,226	32	0.06	0.10	0.19	—	289,232	425	0.22	0.49	0.75
1290	201,970	384	0.19	0.36	0.71	83,762	21	0.05	0.14	0.29	—	285,732	278	0.15	0.30	0.59
1265	193,286	324	0.13	0.27	0.52	85,249	20	0.03	0.07	0.15	—	278,535	231	0.10	0.21	0.40
1240	330,034	293	0.13	0.31	0.59	134,912	21	0.03	0.09	0.19	—	464,946	214	0.10	0.24	0.47
1215	137,865	309	0.20	0.35	0.65	40,184	31	0.08	0.10	0.20	—	178,049	246	0.17	0.29	0.55
1190	318,351	414	0.40	0.72	0.98	122,034	37	0.10	0.17	0.29	—	440,385	309	0.32	0.57	0.79
1165	700,003	438	0.23	0.44	0.71	287,690	39	0.07	0.14	0.26	—	987,693	322	0.18	0.36	0.58
1140	594,614	369	0.19	0.35	0.64	285,444	31	0.06	0.11	0.21	—	880,058	259	0.15	0.27	0.50
1115	375,778	354	0.19	0.41	0.68	152,036	27	0.06	0.14	0.24	—	527,814	260	0.16	0.33	0.55
1090	326,009	343	0.30	0.53	0.98	73,852	37	0.12	0.17	0.30	—	399,861	287	0.26	0.46	0.86
1065	306,459	352	0.47	0.58	1.05	76,018	34	0.13	0.21	0.42	—	382,477	289	0.40	0.51	0.92
1040	312,076	346	0.52	0.59	1.06	92,637	28	0.15	0.36	0.65	—	404,713	274	0.44	0.53	0.97
1015	270,336	307	0.35	0.53	0.99	81,723	40	0.17	0.43	0.63	—	352,059	245	0.31	0.51	0.90
990	324,751	356	0.38	0.54	1.04	67,436	33	0.17	0.36	0.61	—	392,187	301	0.34	0.51	0.96
965	311,391	339	0.41	0.66	1.20	57,861	44	0.18	0.31	0.61	—	369,252	293	0.37	0.60	1.11
940	301,140	320	0.46	0.76	1.44	74,221	33	0.17	0.43	0.85	—	375,361	263	0.40	0.69	1.32
915	294,363	316	0.40	0.82	1.52	57,542	42	0.16	0.39	0.80	—	351,905	271	0.36	0.75	1.40
890	218,102	274	0.40	0.98	1.83	47,243	40	0.16	0.51	0.97	—	265,345	232	0.36	0.90	1.68
865	143,367	262	0.26	0.67	1.45	33,635	32	0.13	0.35	0.76	—	177,002	218	0.24	0.61	1.32
840	117,750	290	0.30	1.44	2.47	35,350	27	0.13	0.41	0.77	—	153,100	229	0.26	1.20	2.08
815	143,685	312	0.49	3.19	4.81	44,059	26	0.14	0.64	1.36	—	187,744	245	0.41	2.59	4.00
790	105,306	300	0.63	5.76	6.62	33,476	23	0.12	0.84	1.85	—	138,782	233	0.51	4.57	5.47
765	102,264	282	0.56	5.11	3.35	65,818	20	0.14	0.98	1.65	—	168,082	180	0.39	3.49	2.69
740	85,347	171	0.40	5.12	4.96	38,262	19	0.11	0.35	0.55	—	123,609	124	0.31	3.65	3.59
715	74,002	150	0.36	6.83	7.32	12,266	29	0.15	0.45	0.73	—	86,268	133	0.33	5.92	6.38
690	48,794	148	0.29	8.71	8.31	19,376	43	0.14	0.73	1.14	—	68,170	118	0.25	6.44	6.27
665	21,799	213	0.36	6.37	5.00	18,905	24	0.09	0.98	1.21	—	40,704	126	0.24	3.87	3.24
640	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—	—
Total	8,087,053	446	0.41	0.90	1.29	2,769,856	34	0.10	0.23	0.42	—	10,856,909	341	0.33	0.73	1.07

A grade-tonnage curve for all stopes in the reserve is shown in Figure 16-19. The values used to generate the curve are based on the diluted average grade for each stope. The general conclusion drawn from this graph is that the reserve is most sensitive to changes in silver grades between 160 and 500 g Ag/t.

138

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Figure 16-19: Reserve Grade-Tonnage Curve

16.8                PASTE BACKFILL

The mining methods in use at the Escobal mine are transverse and longitudinal longhole stoping. Both require use of cemented backfill as an integral part of the mining cycle to ensure stability before mining adjacent pillars.

Transverse stoping is used in wider parts of the deposit. Typical stope dimensions are 10 m in width measured along strike, 25 m high and mined to the full width of the vein which is expected to reach 35 to, locally, 50 m. A mining width of 15 to 35 m is expected to be typical. Mining takes place in a series of primary and secondary stopes. The primary stopes have two exposures of rock walls along the vein strike direction, both of which are the width of the stope and the secondary stopes have two exposures of paste backfill, both of which are the width of the vein. At the base of each production panel, high strength fill is placed to enable removal of the underlying sill pillar.

Longitudinal stoping is used in narrower parts of the deposit up to 15 m wide. Typical stope dimensions are 10 m in length measured along strike, 25 m high and mined the full width of the vein. The longitudinal stopes have one wall of backfill exposure, the width of the mined opening. Both transverse and longitudinal stope mining methods will require paste fill to be undercut at some stage within the mining schedule.

The transverse stope mining schedule requires the vertical paste fill exposures in the primary stopes to achieve target strength after 28 days of curing. The secondary stopes will not be exposed and therefore only require a minimum strength to be achieved after 56 days. The longitudinal stope mining schedule requires the vertical paste fill exposures of the end walls to achieve target strength after 14 days of curing.

The average cement addition rate for the paste fill is currently around 8.65%. Company personnel believe this number will decrease slightly as improved operating procedures and equipment upgrades are implemented. The operating cost estimate in this study assumes 7.65% cement addition for the primary transverse stopes and 6% for all other backfilled openings.

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The Company’s analysis of the existing paste backfill plant’s performance demonstrates that the plant is inadequate to deliver sufficient backfill necessary to sustain mining at increased throughput rates beyond the current rate of 3,500 t/d. The Company identified several constraints with the existing plant that preclude it from delivering additional paste backfill without compromising backfill strength, including insufficient mixing times, lack of buffering capacity, and compromised paste mix control.

Options considered to increase paste backfill capacity were to either renovate the existing plant or construct a new plant and repurpose, where practical, some of the equipment from the existing plant. Renovating the existing plant was determined to be unworkable due the plant’s space constraints which did not allow for increasing the size of the mixer, paste hopper and other equipment without substantial redesign and structural modifications. Rebuilding the plant proved to be the better option to meet planned production goals and provide for additional paste backfill capacity. The rebuilt plant incorporates the existing paste pumps and related systems and includes a tailings breaker to break down the filtered tailings cake prior to entering the paste mixing circuit.

The capital cost for paste backfill plant expansion is estimated at $6.9 million, which includes engineering, equipment procurement, structural steel, construction and installation, and commissioning. The Company’s Board of Directors approved expenditures for the paste plant upgrade in March 2014, at which time engineering and equipment selection and procurement was initiated. The plant is currently under construction with approximately $4.1 million of the capital cost expended as of June 30, 2014. Construction of the plant is expected to be finished in the fourth quarter of 2014, with commissioning scheduled for the first quarter of 2015.

16.9                DEVELOPMENT DESIGN

The active mining areas are accessed through two main portals, called the East Central and West Central portals. These two primary declines provide access to the Central Zone. A third primary ramp is being driven into the East Zone from the Central Zone. The three primary ramps will connect to a system of secondary access spirals and attack ramps to access stoping areas in the East Zone and the East Extension area. Footwall laterals will be driven parallel to the vein on 25 m vertical intervals and will be accessed from the primary ramps. The stopes are accessed from the footwall laterals.

The access ramps are located nominally 75 to 150 m from the vein. There are also accesses leading to ventilation ingress and exhaust raises. Ventilation raises and ore passes are strategically located throughout the mine and included in the pre-production development schedule.

The primary development ramps are designed to be 5 m wide by 6 m high with an arched back and are typically driven at a maximum incline of 15%. Secondary development headings allowing access to the individual stopes are designed 5 m wide by 5 m high.

A summary of the final development requirements is given in Table 16-11 while Figure 16-20 and Figure 16-21 provide a schematic view of the location of these headings. Figure 16-22 combines the development and stope designs for an overall view.

Note that the largest single quantity of development in Table 16-11 is the stope accesses, representing the over-cut and under-cut for each stope. This quantity is driven primarily by the width of the stopes, currently at 10 m. Increasing the width of the stope will reduce the number of stopes and reduce the total number of meters of development required. Approximately 26% of the stope accesses are driven in ore which helps to offset the costs but the excavation costs are significantly higher than if this mass could be excavated through the bulk stoping methods. This is a potential area for decreasing operating costs given suitable ground conditions.

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Table 16-11: Required Development Quantities

Development Type	Length (m)
Primary Ramp	13,335
Auxiliary & Services Headings	7,321
Accesses to Footwall Laterals	3,204
Footwall Laterals	33,945
Stope Accesses	118,640
Subtotal — Horizontal Development	176,444
Raises	2,594
Total Development Required	179,038

Figure 16-20: Development Design — Looking North

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Figure 16-21: Development Design — Oblique View

Figure 16-22: Complete Mine Design — Oblique View

16.10              PRODUCTION SCHEDULE AND MINING RATES

Development sequences are based on average meters of advance per day per heading. These rates vary by location in the mine and the availability of multiple headings. The rates used in the schedule are shown in Table 16-12.

Figure 16-23 illustrates the development quantities produced in each year over the life of the mine by heading type while Table 16-13 through Table 16-15 provide a more detailed description of the development schedule.

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Table 16-12: Development Advance Rates

Development Type	Zone	Single Heading Rate (m/day)	Dual Heading Rate (m/day)
Ramp — Above 1265 Level	Central	1.7	2.5
Ramp — Below 1265 level	Central	1.5	2.5
Ramp	East	2.0	3.0
Level Development — Above 1265 level	Central	3.5	5.25
Level Development — Below 1265 level	Central	1.5	3.0
Level Development	East	3.5	5.25
Raises	All	1	n/a

Figure 16-23: Development Quantities by Type and Period

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Table 16-13: Total Development Quantities Scheduled

		Totals
Primary ramp	meters	13,335
Sublevel Access	meters	3,204
Diamond Drill Station	meters	810
Connection Drift	meters	926
Muckbay	meters	62
Service cutout	meters	1,534
Sump	meters	1,081
Dewatering cutout	meters	101
Vent raise access	meters	2,542
Vent raise	meters	2,402
Footwall Lateral (contrafrente)	meters	33,945
Ore pass access	meters	266
Ore pass raise	meters	192
Stope Access (waste)	meters	87,549
Stope Access (ore)	meters	31,091
Totals		179,038

Table 16-14: Development Quantities Scheduled — 2014 through 2023

		2014	2015	2016	2017	2018	2019	2020	2021	2022	2023
Primary ramp	meters	314	521	1,603	1,120	1,036	1,045	719	977	1,017	704
Sublevel Access	meters	189	76	466	294	247	167	116	192	231	205
Diamond Drill Station	meters	12	37	62	100	76	—	50	63	75	74
Connection Drift	meters	—	—	—	88	—	33	121	296	—	25
Muckbay	meters	13	—	25	—	—	—	—	12	—	—
Service cutout	meters	—	43	102	102	109	70	139	160	151	110
Sump	meters	—	30	105	66	85	30	60	105	146	50
Dewatering cutout	meters	37	63	—	—	—	—	—	—	—	—
Vent raise access	meters	310	212	258	255	150	45	56	30	241	142
Vent raise	meters	115	453	194	213	69	100	41	28	84	162
Footwall Lateral (contrafrente)	meters	656	1,932	3,271	2,448	4,757	2,646	2,287	1,455	2,894	2,596
Ore pass access	meters	—	6	71	33	24	—	—	—	—	—
Ore pass raise	meters	—	—	69	20	—	—	—	—	—	—
Stope Access (waste)	meters	1,495	4,587	5,156	6,403	5,818	7,130	6,918	5,103	4,877	5,467
Stope Access (ore)	meters	1,412	3,706	3,915	2,705	2,447	2,589	1,909	1,954	1,640	1,565
Totals		4,552	11,667	15,297	13,846	14,817	13,855	12,415	10,374	11,356	11,099

Table 16-15: Development Quantities Scheduled — 2024 through 2033

		2024	2025	2026	2027	2028	2029	2030	2031	2032	2033
Primary ramp	meters	637	1,150	1,040	308	596	550	—	—	—	—
Sublevel Access	meters	115	235	402	68	89	114	—	—	—	—
Diamond Drill Station	meters	50	90	46	37	37	—	—	—	—	—
Connection Drift	meters	—	—	363	—	—	—	—	—	—	—
Muckbay	meters	—	—	—	12	—	—	—	—	—	—
Service cutout	meters	106	137	85	105	42	73	—	—	—	—
Sump	meters	75	75	60	105	30	60	—	—	—	—
Dewatering cutout	meters	—	—	—	—	—	—	—	—	—	—
Vent raise access	meters	85	227	138	91	28	143	90	15	7	20
Vent raise	meters	105	246	73	219	25	125	95	55	—	—
Footwall Lateral (contrafrente)	meters	1,136	2,350	1,612	777	767	934	567	144	420	295
Ore pass access	meters	—	7	56	69	—	—	—	—	—	—
Ore pass raise	meters	—	—	54	49	—	—	—	—	—	—
Stope Access (waste)	meters	6,550	5,252	4,024	5,788	3,532	2,672	2,819	2,310	781	866
Stope Access (ore)	meters	1,577	687	1,304	739	935	809	529	523	145	—
Totals		10,435	10,456	9,257	8,367	6,082	5,479	4,101	3,047	1,353	1,181

Stope production and development advance rates are based on actual values encountered since the start of construction in May 2011. Production sequencing includes the time required for on-vein development of over- and under-cuts, followed by longhole mining to the maximum safe strike length, preparation time for backfilling, backfilling and re-entry. Prior to extraction of ore from each stope, approximately 37 days of preparation time is required, including cable bolting, slot raise development and production drilling. Loading and haulage is done at a maximum rate of 4,000

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tonnes per day for any single stope. Six days of preparation time is required prior to backfilling which is accomplished at a rate of 2,450 tonnes per day of paste fill; the actual time per stope is relative to the size of the stope. A total of 28 days of cure time is required prior to mining beside the newly backfilled stope.

Figure 16-24 illustrates the production quantities produced in each year over the life of the mine by stoping while Table 16-16 through Table 16-18 provide a more detailed description of the production schedule. Figure 16-25 through Figure 16-29 depicts the evolution of the underground workings over the life of the mine.

Figure 16-24: Production Quantities by Type and Period

Table 16-16: Total Production Quantities Scheduled

Mine		Totals
Ore	ktonnes	31,433
Silver Grade (g/t)	g/t	347
Lead Grade (%)	%	0.74
Zinc Grade (%)	%	1.21
Gold Grade (g/t)	g/t	0.33

Transverse Stopes	ktonnes	18,916
Silver Grade (g/t)	g/t	349
Lead Grade (%)	%	0.74
Zinc Grade (%)	%	1.28
Gold Grade (g/t)	g/t	0.33

Longitudinal Stopes	ktonnes	10,668
Silver Grade (g/t)	g/t	332
Lead Grade (%)	%	0.74
Zinc Grade (%)	%	1.09
Gold Grade (g/t)	g/t	0.33

Development Ore	ktonnes	1,849
Silver Grade (g/t)	g/t	409
Lead Grade (%)	%	0.72
Zinc Grade (%)	%	1.24
Gold Grade (g/t)	g/t	0.34

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Table 16-17: Production Quantities Scheduled – 2014 through 2023

Mine		2014	2015	2016	2017	2018	2019	2020	2021	2022	2023
Ore	ktonnes	658.3	1,529.2	1,624.3	1,657.9	1,642.5	1,632.7	1,632.8	1,631.3	1,618.7	1,642.5
Silver Grade (g/t)	g/t	542	482	441	442	442	442	442	442	398	266
Lead Grade (%)	%	0.74	0.68	0.65	0.67	0.67	0.66	0.82	0.46	0.61	0.56
Zinc Grade (%)	%	1.27	1.19	1.13	1.12	1.13	1.11	1.39	0.78	1.01	1.03
Gold Grade (g/t)	g/t	0.38	0.36	0.31	0.32	0.32	0.40	0.61	0.27	0.34	0.27
Transverse Stopes	ktonnes	580.7	1,308.1	1,245.3	1,204.2	977.8	1,141.2	1,121.1	1,313.1	1,207.2	1,128.4
Silver Grade (g/t)	g/t	549	482	402	421	345	369	400	430	382	275
Lead Grade (%)	%	0.76	0.69	0.65	0.69	0.81	0.71	0.94	0.45	0.62	0.62
Zinc Grade (%)	%	1.30	1.21	1.17	1.20	1.39	1.22	1.61	0.75	1.03	1.15
Gold Grade (g/t)	g/t	0.39	0.37	0.32	0.32	0.31	0.32	0.38	0.24	0.34	0.28
Longitudinal Stopes	ktonnes	—	—	145.5	292.3	518.7	337.1	397.8	201.6	312.4	420.8
Silver Grade (g/t)	g/t	—	—	634	522	623	716	564	533	453	239
Lead Grade (%)	%	—	—	0.70	0.61	0.39	0.43	0.44	0.52	0.58	0.40
Zinc Grade (%)	%	—	—	0.89	0.77	0.58	0.63	0.69	0.93	0.94	0.66
Gold Grade (g/t)	g/t	—	—	0.28	0.31	0.32	0.70	1.31	0.41	0.32	0.26
Development Ore	ktonnes	77.7	221.1	233.5	161.4	146.0	154.4	113.9	116.5	99.1	93.3
Silver Grade (g/t)	g/t	487	486	527	453	448	383	429	422	417	282
Lead Grade (%)	%	0.60	0.64	0.64	0.68	0.77	0.81	0.98	0.46	0.64	0.63
Zinc Grade (%)	%	1.03	1.10	1.09	1.12	1.30	1.39	1.68	0.78	1.05	1.17
Gold Grade (g/t)	g/t	0.33	0.33	0.30	0.32	0.35	0.36	0.42	0.26	0.35	0.28

Table 16-18: Production Quantities Scheduled – 2024 through 2033

Mine		2024	2025	2026	2027	2028	2029	2030	2031	2032	2033
Ore	ktonnes	1,649.5	1,635.5	1,647.3	1,638.6	1,641.6	1,654.2	1,637.6	1,637.2	1,659.9	1,361.6
Silver Grade (g/t)	g/t	278	283	332	312	272	266	243	251	246	224
Lead Grade (%)	%	0.63	0.63	0.75	0.81	0.86	0.75	0.87	0.80	1.07	1.08
Zinc Grade (%)	%	1.04	1.11	1.31	1.50	1.47	1.27	1.41	1.02	1.27	1.80
Gold Grade (g/t)	g/t	0.34	0.33	0.41	0.44	0.37	0.32	0.28	0.17	0.17	0.25
Transverse Stopes	ktonnes	1,098.0	1,105.7	787.7	806.0	923.9	994.4	907.7	771.0	229.1	65.2
Silver Grade (g/t)	g/t	297	290	292	295	271	264	243	245	220	220
Lead Grade (%)	%	0.67	0.62	0.93	1.12	0.97	0.77	0.69	0.59	1.50	1.70
Zinc Grade (%)	%	1.14	1.16	1.66	2.13	1.61	1.24	1.13	0.98	2.58	2.94
Gold Grade (g/t)	g/t	0.37	0.33	0.47	0.55	0.39	0.27	0.25	0.15	0.26	0.29
Longitudinal Stopes	ktonnes	457.5	488.8	781.8	788.5	661.9	611.6	698.3	835.0	1,422.2	1,296.3
Silver Grade (g/t)	g/t	225	267	376	329	274	268	244	255	250	224
Lead Grade (%)	%	0.53	0.64	0.55	0.48	0.69	0.72	1.11	1.00	1.00	1.05
Zinc Grade (%)	%	0.76	0.98	0.92	0.82	1.27	1.31	1.78	1.05	1.05	1.74
Gold Grade (g/t)	g/t	0.24	0.33	0.35	0.32	0.34	0.41	0.33	0.18	0.15	0.25
Development Ore	ktonnes	94.0	41.0	77.8	44.1	55.8	48.2	31.6	31.2	8.7	—
Silver Grade (g/t)	g/t	306	303	300	304	277	271	253	268	239	—
Lead Grade (%)	%	0.69	0.64	0.95	1.16	1.00	0.79	0.73	0.60	1.59	—
Zinc Grade (%)	%	1.19	1.20	1.69	2.20	1.66	1.28	1.19	1.00	2.73	—
Gold Grade (g/t)	g/t	0.39	0.35	0.48	0.57	0.41	0.28	0.26	0.15	0.28	—

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Figure 16-25: Planned Mine Status – 2015

Figure 16-26: Planned Mine Status – 2020

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Figure 16-27: Planned Mine Status – 2025

Figure 16-28: Planned Mine Status – 2030

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Figure 16-29: Planned Mine Status — End of Mine Life (2033)

16.11MINING EQUIPMENT AND INFRASTRUCTURE

The major mine equipment includes R1700 and R2900 Caterpillar LHDs with 6.0 and 8.3 m3 buckets, respectively, and equipped for remote operation. A fleet of 45-tonne Caterpillar trucks are used for hauling ore and waste out of the mine. Atlas Copco two-boom electric hydraulic jumbos are utilized to drive the development headings and stope development headings. Ground support is installed in all headings with a fleet of Atlas Copco electric-hydraulic jumbos capable of installing split set, and swellex bolts. Cable bolting can be done using Atlas Copco Simba drills or dedicated MacClean cable bolters; the Simba’s are also the primary production drill in the longhole stopes. Cubex track-mounted drills equipped with a V-30 reaming head drill the breaking slots in the stopes. These drills are equipped with top hammers and are capable of drilling larger diameter holes for utilities as well as production drilling where larger diameter holes are desirable. Diamond drills are utilized for stope definition to enhance production planning prior to stope production. The fresh air intake in the East Zone will be initially developed using a contractor. Caterpillar 120 AWD graders will be used for road maintenance in the mine. Support equipment includes shotcrete remote spray jumbos and mixer trucks, scissor lifts, explosives trucks, and various materials handling vehicles. A list of the equipment fleet and unit requirements is shown below in Table 16-19.

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Table 16-19: Mine Mobile Equipment List

Equipment Type	Typical Quantity Required (LOM)
R 2900 CAT LHD	3
R 1700 CAT LHD	6
AD 45 CAT Truck	7 to 12
AC Jumbo Boomer 282 (2-boom)	7 to 10
AC Boltec MD Rock Bolt Jumbo	5 to 7
AC Simba Longhole Drill	5
Cubex Longhole Drill w/V-30 head	3
Shotcrete Spray Jumbo	2
McClean Cable Bolter	3
Mechanical Scaler	2
Underground Transit Mixer Truck	3
AWD 120 CAT Motor Grader	1
LM55 and LM75 Diamond Drills	3
Scissor Lift	3
ANFO Truck	2
Telehandler	3
Boom Truck	3
Fan Truck	1
Lube Truck	2
HD Pickups	30
Personnel Transport	12

The mine requires compressed air to run certain pieces of mining equipment such as drills and pumps. It is estimated that approximately 4,300 cfm of compressed air is required. This includes allowance for line losses (leaks) of 700 cfm. The compressor plant is located on surface central to the two mine portals. The plant includes compressors and receiver tanks. The plant has a master control that will stop and start the compressors as demanded by the mine. Compressed air is supplied to the working areas underground via 8” and 10” steel pipe. The compressed air distribution system consists of two pipelines, the first of which runs down the East Central ramp and the second which runs down the West central ramp. The two pipelines are tied in at drifts connecting the two ramps in order to provide balance and backup to the system.

The line running down the East Central ramp is sized at 10 inches where it enters the portal and remains as such up to the East Zone ramp future connection, where it reduces to 8 inch. The reason for the increased initial size is to provide capacity for a future connection with the East Central ramp. Air is provided to active mining areas through sixinch lines; two-inch drop downs are also provided in the ramp.

The compressed air system has a stench release warning system that can be activated in the event the mine must be evacuated.

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For the service water system, a six-inch main line runs down the West Central and East Central ramps, beginning at the portals and ending at the mine bottom. To deliver water to active mining areas, four- inch take offs are installed at the access drifts.

16.12MINING WORK FORCE

The mine operates two 11-hour shifts per day, 350 days per year, utilizing three mine crews working a rotating schedule. The total mine personnel requirement is 496, including all management and supervisory personnel. Currently some 457 people are employed in mine operations and that total will increase to 496 by the end of 2016. Table 16-20 provides a breakdown to the mine personnel by primary work area while Table 16-21 through Table 16-30 provide additional detail within each work area.

Table 16-20: Summary of Mine Personnel

Work Area	Quantity Required
Mine Operations	182
Mine Services	41
Training	4
Backfill Services	44
Shotcrete	33
Mine & Surface Equipment Shop	85
Mine Electrical Services	27
Mine Engineering	20
Geology	20
Underground Drilling	40
Total Mine Personnel	496

Table 16-21: Summary of Mine Operations Personnel

Job Description	Quantity Required
Administrative	4
Supervisors	13
Miners & Equipment Operators	141
Helpers	20
Mining Lamp	3
Drilling	1
Total Mine Operations Personnel	182

Table 16-22: Summary of Mine Services Personnel

Job Description	Quantity Required
Supervisors	4
Miners & Equipment Operators	33
Helpers	4
Total Mine Services Personnel	41

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Table 16-23: Summary of Training Personnel

Job Description	Quantity Required
Administrative	2
Supervisors	2
Total Training Personnel	4

Table 16-24: Summary of Backfill Services Personnel

Job Description	Quantity Required
Supervisors	5
Miners & Equipment Operators	11
Helpers	3
Plant Operators	9
Instrumentation & Technicians	11
Welders & Mechanics	5
Total Backfill Services Personnel	44

Table 16-25: Summary of Shotcrete Services Personnel

Job Description	Quantity Required
Supervisors	3
Miners & Equipment Operators	21
Helpers	9
Total Shotcrete Services Personnel	33

Table 16-26: Summary of Mine & Surface Equipment Shop Personnel

Job Description	Quantity Required
Administrative	5
Supervisors	7
Mechanics	45
Electricians	4
Welders	15
Helpers	8
Parts Runner	1
Total Mine & Surface Equipment Shop Personnel	85

Table 16-27: Summary of Mine Electrical Services Personnel

Job Description	Quantity Required
Administrative	3
Supervisors	4
Electricians	17
Helpers	3
Total Mine Electrical Services Personnel	27

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Table 16-28: Summary of Mine Engineering Personnel

Job Description	Quantity Required
Planning	6
Ventilation & Services	2
Geotechnical	2
Costs	1
Surveying	8
Lab technician	1
Total Mine Engineering Personnel	20

Table 16-29: Summary of Geology Personnel

Job Description	Quantity Required
Geologists	10
Samplers & Core Shed	8
Draftsman	1
Admin. Assistant	1
Total Geology Personnel	20

Table 16-30: Summary of Underground Drilling Personnel

Job Description	Quantity Required
Administrative	2
Supervisors	4
Drillers	12
Mechanics	3
Helpers	19
Total Underground Drilling Personnel	40

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17RECOVERY METHODS

17.1INTRODUCTION

Ore from the Escobal mine is treated using conventional differential flotation processes producing lead concentrates with high precious metal (silver+gold) grades and zinc concentrates with a lesser precious metal component. Lead and zinc concentrates are sold and shipped to multiple international smelter clients.

The original design basis for the processing facility is 3,500 tonnes of ore per day (t/d) or 1.28 million tonnes per year; though the installed crushing, grinding, flotation and concentrate processing components were sized for the contemplated increased throughput rate of 4,500 t/d discussed in this report.

Mill commissioning was initiated in the second half of 2013. The first metal concentrates were produced on September 30, 2013, with the inaugural concentrate shipment on October 15, 2013. The Company declared commercial production in January 2014 with the completion of mill commissioning and continued to ramp up the mill throughput rate through the first half of 2014. The Escobal ore processing facility is now operating above the 3,500 t/d design rate and averaged 3,708 t/d in June 2014 with metal recoveries generally meeting or exceeding design expectations.

17.2PROCESS SUMMARY

The process flow sheet for the Escobal ore and subsequent process plant construction design was based on the results of metallurgical and flotation test work and pilot plant studies conducted from 2009 through 2013, as discussed in Section 13, Mineral Processing and Metallurgical Testing. The process selected for recovering the silver, gold, lead and zinc is differential flotation, where the ore is crushed and ground to a fine size and processed through mineral flotation circuits. The following items summarize the process operations required to extract silver, gold, lead and zinc from the Escobal ore to create lead and zinc concentrates.

·Primary Crushing — size reduction by primary jaw crusher to reduce the ore from run-of-mine (ROM) to minus 150 mm.

·Secondary and Tertiary Crushing — size reduction of the primary crushed material by secondary and tertiary crushing to reduce the ore particle size from 150 mm to minus 9 mm.

·Grinding — grinding crushed material in a ball mill circuit to a size suitable for processing in a flotation circuit. The ball mill operates in closed circuit with hydrocyclones to deliver a design ore size of 80 percent passing 105 microns (P80105) to the flotation circuit.

·Flotation — the flotation plant consists of selective lead and zinc flotation circuits. The ore slurry from the ball mill is first routed to the lead flotation circuit, then through the zinc flotation circuit. Each flotation circuit consists of rougher flotation cells, cleaner flotation cells, scavenger circuits, and regrind circuits to maximize metal recovery.

·Concentrates — final lead and zinc concentrates are thickened, filtered to reduce moisture, sampled, loaded in one- or two-tonne super sacks and placed in 20-tonne shipping containers for transport to a local port where they are placed on vessels for delivery to smelter clients.

·Flotation tailings are thickened and filtered to reduce moisture and conveyed to either the dry stack tailing facility or to the paste backfill plant.

Water from tailing and concentrate dewatering is treated and recycled for reuse in the flotation process. The Escobal mine design maximizes water recycling and the reuse of process water in order to minimize treatment and discharge.

A flow sheet illustrating the mineral processing at the Escobal mine is shown in Figure 17-1.

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Figure 17-1: Escobal Process Flowsheet

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17.3PROCESS PLANT DESCRIPTION

The following description of the Escobal mine process plant is summarized from Performance Associates International, Inc. (2013).

17.3.1Primary Crushing

Ore from the mine is trucked to the surface and dumped on an ore stockpile proximal to the primary crusher. The ore is picked up by a front-end loader and dumped into the crusher feed hopper, where it passes through a stationary grizzly to a variable speed vibrating grizzly for initial separation, with oversized material going to the primary crusher and undersized material by-passing the primary crusher.

The stationary grizzly retains ore larger than 600 mm as oversize. The oversize material is broken with a pedestal-mounted Tramac Model TX62 hydraulic rock breaker until it passes through the grizzly and drops into a 100-t total capacity (50-t live capacity) primary crusher feed hopper. The ore is then drawn from the hopper through a chute onto a vibrating grizzly feeder that separates the undersize minus 100 mm ore from the coarse ore feed to the crusher. The undersize ore falls directly from the vibrating grizzly onto the secondary screen feed conveyor. The vibrating grizzly feeder is a Telsmith Model VGF SD which is equipped with a model 280HF vibration unit driven by a 37-kW electric motor and variable-frequency drive (VFD). The maximum capacity of the feeder is 430 dry metric tonnes per hour (dmt/h).

Material from the vibrating grizzly feeder drops into the primary crusher feed chute that feeds directly into the primary crusher. The ore is crushed to 80 percent minus 150 mm and then discharged onto the secondary screen feed conveyor via the jaw crusher discharge chute. The primary crusher is a Telsmith Model 3055 Standard single-toggle jaw crusher, 9.50 m x 1.25 m in size, which is driven a 160-kW electric motor with a flywheel diameter of 1,370 mm. The crusher feed opening is 762 mm by 1,397 mm. The design maximum operating capacity of the crusher is 430 dmt/h of oversize ore. The process design for the nominal feed rate to the primary crusher is 219 dmt/h.

Figure 17-2: Escobal Mine — Primary Crushing Plant

17.3.2Secondary and Tertiary Crushing

Secondary/tertiary crushing reduces the ore to minus 9 mm as feed to the ball mill. Ore coming from the primary crusher first passes a triple deck secondary screen with oversized material going to the secondary 1.73 m diameter, HP400 cone crusher. Undersized material from the screen and crushed material are fed to a triple deck screen, with oversized

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material going to the tertiary 1.47 m diameter HP300 cone crusher. Undersized material and crushed material exit the tertiary crushing area and is transported to one of two 2,000 t fine ore storage bins.

The secondary screen is a Telsmith Triple-Deck Vibro-King TL Model 6x20 Dual TL26 unit, powered by two 18.8-kW electric motors, that also serves as the feeder for the secondary crusher. Each deck is 1,829 mm by 6,096 mm, with each deck passing progressively finer material. The bottom deck passes minus 10 mm undersize ore, which then drops onto the fine ore transfer conveyor. The design nominal feed rate to the screen is 219 dmt/h of ore with an average particle size of 80 percent minus 150 mm. The bottom deck undersize to the fine ore transfer conveyor is approximately 46 dmt/h, while the oversize feed from all decks to the secondary crusher is about 235 dmt/h. The screen panel openings are selected from operating experience to optimize the throughput of the target size and tonnage to the fine ore bins.

The secondary crusher is a Telsmith Model 44SBS cone crusher driven with V-belts by a 225-kW electric motor. The design nominal feed averages 183 dmt/h. The crushed ore discharged from the secondary cone crusher is conveyed to the tertiary screens on the crusher discharge and tertiary screen feed conveyors at a nominal process design rate of 183 dmt/h of ore (of which 80 percent is minus 50 mm). The secondary crusher discharge is combined with 206 dmt/h of ore (of which 80 percent is minus 9 mm) from the two tertiary crushers, which then discharges onto the same conveyor to be returned to the tertiary screens.

The combined crushed ore is then delivered through tertiary screen feed bins that distribute the ore onto the two tertiary screen feeder belts. The feeder belts then discharge the ore onto each of the two tertiary screens. The crushed ore is weighed on the tertiary screen feed conveyor using a belt scale. The weight of crushed ore passing along the conveyor is measured to inform the operator of the actual crushing rate and then totalized for metallurgical accounting. The ore is fed from the bins directly onto two tertiary screen feeder belts. Each bin has a live capacity of 50 t and are equipped with slide gates to shut off the flow of ore to either or both tertiary screen belts. The ore then flows onto two tertiary screens which split the feed into two outlets to choke feed ore into the two tertiary crushers. The feeders each have a variable-speed drive with 22.4-kW motors. The two tertiary screens are triple-deck inclined vibrating screens, similar to the secondary screens described above. The size of the openings on the screen media on each deck is smaller than on the secondary screen, resulting in the bottom deck passing 80 percent minus 9 mm undersize ore. The undersize ore is the final crushing plant product. This product drops onto the fine ore transfer conveyor. The tertiary screens also serve as feeders for the tertiary crushers.

The process design calls for a nominal rate of 219 dmt/h of the undersize fine ore from both screens to fall directly onto the fine ore transfer conveyor. Accordingly, the oversize ore (80 percent minus 50 mm) feeds the two tertiary crushers at the rate of about 206 dmt/h.

The final step in the secondary and tertiary crushing system is the crushing of the oversize ore from the tertiary screens. Each tertiary screen discharges the oversize ore into its corresponding tertiary crusher. There are two Telsmith Model 44SBS tertiary cone crushers, each driven with V-belts by 225-kW motors. The process design is such that each tertiary cone crusher processes 103 dmt/h of feed, which is 80 percent minus 50 mm; the two crushers deliver a total of 206 dmt/h of 80 percent minus 9 mm ore to the crusher discharge conveyor. The maximum capacity of each crusher is 430 dmt/h.

The crushing plant product is stored in two fine ore bins, each with a capacity of 2,000 t live.

17.3.3Conveyance

There are six conveyors serving the secondary and tertiary crushing system—the crusher discharge conveyor, the tertiary screen feed conveyor, the fine ore transfer conveyor, the reversible conveyor, the fine ore bin feed conveyor, and the fine ore bin reversible conveyor. All of these conveyors are standard belt conveyors with the same basic components.

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·Crusher Discharge Conveyor — The crusher discharge conveyor is a 914 mm-wide belt conveyor that is 112 m long and inclined to rise 12.3 m to the head pulley. It discharges onto the tertiary screen feed conveyor. The nominal process design loading is 219 dmt/h, which is 60 percent of the maximum design load. The conveyor is driven through a gearbox for speed reduction by a 56-kW electric motor.

·Tertiary Screen Feed Conveyor — The tertiary screen feed conveyor is a 914 mm-wide belt conveyor that is 101 m long and inclined to rise 18.4 m to the head pulley. It discharges into the tertiary screen feed bins. The nominal process design loading is 219 dmt/h, which is 60 percent of the maximum design load. The conveyor is driven through a gearbox for speed reduction by a 56-kW electric motor.

·Fine Ore Transfer Conveyor — The fine ore transfer conveyor is a 914 mm-wide belt conveyor that is 84 m long and inclined to rise 11 m to the head pulley. It discharges onto the reversible conveyor. The nominal process design loading is 219 dmt/h, which is 60 percent of the maximum design load. The conveyor is driven through a gearbox for speed reduction by a 56-kW electric motor.

·Reversible Conveyor — The reversible conveyor is a horizontal 914 mm-wide belt conveyor that is 36 m long and discharges onto either the fine ore bin feed conveyor or the crushed waste stockpile. The nominal process design loading is 219 dmt/h, which is 34 percent of the maximum design load. The conveyor is driven through a gearbox for speed reduction by an 18.6-kW electric motor.

·Fine Ore Bin Feed Conveyor — The fine ore bin feed conveyor is a 914 mm-wide belt conveyor that is 83 m long and inclined to rise 20.7 m to the head pulley. It discharges onto the fine ore bin reversible conveyor. The nominal process design loading is 219 dmt/h, which is 34 percent of the maximum design load. The conveyor is driven through a gearbox for speed reduction by a 56-kW electric motor.

·Fine Ore Bin Reversible Conveyor — The fine ore bin reversible conveyor is a horizontal 914 mm-wide belt conveyor that is 14 m long and discharges into either of the fine ore bins. The nominal process design loading is 219 dmt/h, which is 34 percent of the maximum design load. The conveyor is driven through a gearbox for speed reduction by an 18.6-kW electric motor.

Figure 17-3: Escobal Mine — Secondary & Tertiary Crushing Plant

17.3.4Grinding

Crushed ore is fed to the grinding circuit from the two fine ore bins. Two variable speed belt feeders are located under each fine ore bin. Ore is withdrawn from the fine ore bins by these belt feeders and the ore is discharged onto the mill feed conveyor. The mill feed conveyor delivers the crushed ore to the ball mill for grinding. In addition to the fine ore, the ball mill receives grinding balls (generally 3- to 4.5-inch diameter) which grind the ore as the mill rotates. Process

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water is added to form a slurry in the mill. Zinc cyanide and zinc sulfate are added to the mill feed chute to prepare the ground ore for flotation.

The mill feed is introduced through a feed spout or chute that is connected to a feed hopper. The feed chute introduces ore, reagents, and process water into the mill. The process water provides the dilution water necessary to control slurry density in the mill. The design slurry density is 70 percent solids.

A ball feeder meters the grinding balls at a rate of up to 15 t/d. The ball storage hopper has a 10-t capacity that is filled twice per day, once per 12-hour shift. The Escobal ball mill operates with a maximum ball charge of 30 percent total volume.

The mill is driven by a single 3,357-kW fixed-speed synchronous motor having a nominal speed of 200 rpm. The ball mill design throughput rate is 203.8 t/h solids, which is sufficiently designed for the increased throughput of 4,500 tonnes per day (187.5 t/h). The grinding circuit was engineered and sized to reduce the ore to 80 percent passing 105 microns (P80105). The only modification to the existing ball mill components necessary to accommodate the expanded throughput is replacement of the existing pinion gear with a larger gear, which the company holds in its current inventory.

Slurry flows from the ball mill through the discharge trunnion into a trommel screen attached to the mill. The ball mill discharge slurry from the trommel undersize goes to the cyclone feed sump. The oversize is collected in a tote box and returned to the crushing plant for recycle. The trommel screen undersize is combined with process water in the cyclone feed sump. The cyclone feed pump delivers the slurry from the sump to the primary grinding cyclones.

The cyclones produce the final grinding circuit product. The cyclone overflow contains the fine solids fraction in the feed and it reports to lead flotation. The cyclone underflow contains the coarse solids fraction that returns for additional grinding in the ball mill. The target slurry densities are 33 percent solids for the cyclone overflow and 70 percent solids for the underflow. There are four cyclones installed (three in operation and one spare). The cyclones are Krebs model gMax26, each having a 660-mm diameter cylinder. The cyclone feed pumps are variable-speed horizontal centrifugal pumps that are equipped with a packed shaft seal with gland seal water. The pumps are each equipped with a 186-kW electric motor and a variable-frequency drive.

Figure 17-4: Escobal Mine — Ball Mill

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17.3.5Flotation

Differential, or selective, flotation is used to recover precious and base metals from the Escobal ore. The slurry exiting the ball mill is first routed through a lead flotation circuit which depresses the zinc and recovers the majority of the silver, gold, and lead. The lead tailing is then processed through a zinc flotation circuit where the zinc is reactivated for recovery with minor amounts of silver, gold, and lead.

17.3.5.1Lead Flotation

The lead flotation circuit consists of rougher flotation cells, cleaner flotation cells and cleaner scavenger cells. The lead flotation system produces a silver-rich lead concentrate which also contains the majority of the recovered gold.

The grinding cyclone overflow is delivered to the lead flotation circuit as slurry containing approximately 30 percent solids by weight. The slurry flows by gravity to the lead rougher flotation conditioning tank, where reagents are added and agitated, before flowing to the lead rougher flotation cells. Lead rougher flotation consists of one row of seven 40 m3 WEMCO self-aspirated cells with 75-kW agitator motors in series. The cells in each line are arranged so that they drop in elevation from one cell to the next. Lead rougher flotation takes place at a pH of approximately 8. Slurry enters the rougher flotation row through a feedbox and flows through the first cell, then passes into the next successive cell, with each cell producing concentrate. This process continues throughout the bank of cells until the slurry discharges by gravity into the tailings launder. The main process flow then passes to the zinc rougher conditioning tank where it becomes the primary feed for the zinc rougher flotation circuit. Overflow from the rougher flotation cells flows to the lead regrind sump.

Lead rougher flotation concentrate regrinding is performed in a lead regrind mill (Eirich ETM300 vertical mill) which operates in closed circuit with Krebs gMAX cyclones. The regrind mill feed size is 80 percent passing 100 microns and the product is 80 percent passing 37 microns. The lead regrind mill discharge is combined with the lead rougher flotation concentrate in the lead regrind sump and pumped by variable speed horizontal centrifugal slurry pumps, driven by 75-kW variable-frequency drive motors, at approximately 78 t/h of solids to the lead regrind cyclone with a slurry density of about 45 percent. Most of the coarser particles and some water in the feed stream are classified to the cyclone underflow stream which contains approximately 70 percent solids and flows to the lead regrind mill at a rate of about 80 t/h. The cyclone underflow feeds the lead regrind mill which discharges back to the lead regrind sump. The cyclone overflow stream (final regrind circuit product) flows by gravity to the lead first cleaner conditioning tank at about 93 tonnes per hour at approximately 24 percent solids, from which it is pumped to the cleaner circuits.

Overflow slurry from the lead regrind cyclone is the feed to the lead cleaner and cleaner scavenger flotation circuits. The design feed is 22.3 t/h of solids at a slurry density of 24 percent. The reground lead rougher flotation concentrate is agitated with reagents (zinc sulfate and zinc cyanide) in a 5.6 m3 conditioning tank then gravity-fed to the first cleaner flotation cells.

The first lead cleaner flotation consists of five Westpro flotation cells, each with a volume of 5 m3 and equipped with 15-kW rotor driven motors. Low-pressure flotation air supplied by three flotation air blowers is injected into the cells through the rotor shafts. Frother, xanthate, and promoter reagents are added to the first lead cleaner feedbox. Additional frother and zinc cyanide are added to the feedbox of the third flotation cell.

Tailing from the lead first cleaner flotation cells flows by gravity to the lead first cleaner scavenger flotation cells. The first cleaner scavenger circuit consists of two Westpro flotation cells identical to the first cleaner flotation cells. The design flow of lead first cleaner tailings to the lead first cleaner scavenger flotation cell is 30.4 t/h of solids at a slurry density of 22 percent. Xanthate, promoter, and frother reagents are added to the lead first cleaner scavenger flotation cell feedbox.

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Concentrate from the first cleaner scavenger flotation cells flows into the lead first cleaner scavenger flotation launder, which normally discharges into the lead regrind sump. The design production of the lead first cleaner scavenger concentrate is 7.6 t/h of solids at a slurry density of 25 percent. For metallurgical or maintenance requirements, a dart valve in the lead first cleaner scavenger flotation launder allows diverting the first cleaner scavenger concentrate to the lead rougher conditioning tank where the fist cleaner scavenger concentrate can be directed to the lead rougher conditioning tank, lead rougher flotation cells, or lead first cleaner conditioning tank.

Tailings from the lead first cleaner scavenger flotation circuit flow by gravity to the lead first cleaner scavenger tailing sump. There are two variable-frequency drive first cleaner scavenger tailings pumps (one operating and one on standby) equipped with gland seal water and are driven by 18.65-kW motors. Slurry from the sump is pumped to the lead first cleaner scavenger tailings sampler and through to the lead first cleaner scavenger tailings sampler splitter box where the tailings are directed to either the lead rougher conditioning tank or the zinc rougher tailings sump. The design production of lead first cleaner scavenger tailings is 22.6 t/h of solids at a slurry density of 21 percent.

Concentrate produced from the lead first cleaner flotation circuit is pumped by the lead second cleaner feed pumps to the lead second cleaner flotation circuit, which consists of two banks of two 5-m3 Westpro agitated flotation cells identical to the first cleaner and first scavenger flotation cells. Xanthate, promoter, and frother reagents are added to the lead second cleaner scavenger flotation cell feedbox. The flotation process in the second cleaner flotation cells is similar to that in the other cleaner flotation cells. The nominal lead second cleaner concentrate production is 12.4 tonnes per hour of solids at a slurry density of 33 percent.

Froth that overflows from the second cleaner flotation cells is collected in the lead first/second cleaner flotation launder. The launder for the first bank of second cleaner flotation cells has two dart valves that direct the concentrate to either the lead first/second cleaner flotation launder, where it becomes the feed for the lead third cleaner flotation cell, or to the lead third cleaner flotation launder, where it flows to the lead concentrate thickener feed sump. The launder for the second bank of cleaner flotation cells also has two dart valves that direct its concentrate in an identical manner.

Tailings from the second cleaner flotation flow by gravity back to the first cleaner conditioning tank for additional lead recovery. The design flow of lead second cleaner tailings is 21.1 t/h of solids at a slurry density of 24 percent.

Concentrate from the lead second cleaner flotation cells is pumped to the feedbox of the lead third cleaner flotation circuit to further raise the grade of the concentrate. Frother is also added at this location. The lead third cleaner flotation circuit is composed of three 1.5-m3 Westpro cells with 7.5-kW agitators. The lead third cleaner concentrate slurry flows to the lead third cleaner flotation launder and into the lead concentrate thickener feedbox. Concentrate slurry is pumped from the feedbox to the lead concentrate thickener bytwo vertical centrifugal pumps powered by 5.6-kW motors. The design lead third cleaner concentrate production is 6.7 t/h at a slurry density of 35 percent. The lead third cleaner concentrate is the final product of lead flotation.

Tailings from the second cleaner flotation flow by gravity back to the first cleaner conditioning tank for additional lead recovery. Likewise, tailings from the third cleaner flotation flow by gravity back to the second cleaner flotation circuit for additional lead recovery.

17.3.5.2Reagents — Lead Flotation

Reagents used in the Escobal lead flotation circuit include sulfide collector, promoter, frother, depressant, and activator reagents.

·Sulfide Collector — The sulfide collector used in the Escobal lead flotation circuit is potassium amyl xanthate (PAX). When added to the lead flotation circuit, it causes lead minerals to attach themselves to air bubbles allowing the lead to float to the top the flotation cells. Dry xanthate is mixed to a solution strength of 10 percent

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with the xanthate added to the process by means of flow control valves. The design consumption of xanthate in lead flotation is 0.010 kg/t of mill feed.

·Promoter — Flomin C-7931 is used as the flotation collector reagent to promote flotation of lead sulfide particles in the lead flotation feed. When added to the lead flotation circuit, it causes lead minerals to attach themselves to air bubbles, allowing the lead to float to the top the flotation cells. This ready-to-use reagent is pumped from a storage tank through a circulating loop where it is added to the process by means of metering pumps. The design consumption of Flomin C-7931 is 0.010 kg/t of mill feed.

·Frother — The frother used in the lead flotation circuits is Cytec AF-70 (methyl isobutyl carbinol). When added to the lead flotation cells, it creates froth to which lead minerals attach themselves. The bubbles within the flotation cells float to the top of the cell along with the attached lead mineral. The frother is received at the plant as a full-strength liquid and delivered to the flotation circuit without dilution. Individual meter pumps withdraw the frother from a supply header to deliver the reagent to various addition points in the flotation circuit. The design consumption of AF-70 in lead flotation is 0.10 kg/t of mill feed.

·Depressants — Iron sulfide that occurs in the lead flotation concentrate is depressed by the addition of zinc cyanide slurry, so that much of the pyrite reports to the tailings. Zinc cyanide is prepared by adding sodium cyanide to acid-neutralized zinc sulfate. The resulting zinc cyanide slurry is pumped to a storage tank from which it is distributed via a pipe loop to the various addition points throughout the plant. Actual addition of this reagent to the flotation circuit is accomplished by flow control valves. The design consumption of zinc cyanide in lead flotation is 0.025 kg/t of mill feed.

·Zinc sulfate is used as the zinc sulfide mineral depressant during lead flotation. Dry zinc sulfate is mixed to a solution strength of ten percent with the zinc sulfate added to the process by means of flow control valves. The design consumption of zinc sulfate in lead flotation is 0.075 kg/t of mill feed.

17.3.5.3Zinc Flotation

The zinc flotation circuit is similar to the lead flotation circuit, in that it consists of rougher flotation cells, cleaner flotation cells and cleaner scavenger cells. The zinc flotation system produces a zinc-rich concentrate with minor silver, gold, and lead that remained in the lead flotation tailings.

The zinc flotation conditioning tank receives the lead flotation tailings as a slurry of approximately 30 percent solids which are mixed with zinc activator, collector, and promoter reagents and agitated. The zinc rougher conditioning has an effective volume of 39 m3 and a design flow rate of approximately 471 m3/h. The zinc rougher conditioning tank includes a rubber-lined, dual-impeller agitator driven by an 18.65 kW motor. The zinc rougher conditioning tank discharges by gravity into the first zinc rougher flotation cell feedbox where frother is added to the slurry.

The zinc rougher flotation circuit consists of one row of seven 40 m3 WEMCO self-aspirated cells with 75-kW agitator motors in series. Each flotation cell has two dart valves that control the flow of tailings slurry into the next downstream flotation cell. The slurry from the first rougher flotation cell discharges through two dart valves into the next cell and into each successive cell. The nominal rougher concentrate production is 115 t/h of slurry, containing 25 percent solids.

Zinc rougher concentrate froth is directed into the zinc rougher concentrate sampler feedbox. From this feedbox, the concentrate passes through the zinc rougher concentrate sampler and continues to the zinc regrind sump, which is an open rectangular sump with an operating volumetric capacity of 7.7 m3. This sump serves as the pump tank for the zinc regrind cyclone feed pumps. In addition to the zinc rougher concentrate slurry, the zinc regrind sump also normally receives zinc first cleaner scavenger concentrate.

The final tailings from rougher flotation flow by gravity from the last rougher flotation cell into the tailings thickener feed sampler box then into the zinc rougher tailings sump, which is an open, rectangular sump with an operating volumetric capacity of 6.3 m3. Slurry from the zinc rougher tailings sump is pumped to the tailings thickener through two discharge lines which combine into a common tailings thickener feed line equipped with a flow meter and a density gauge. The

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tailings pumps (one operating and one spare) are horizontal, centrifugal pumps equipped with gland seal water and driven by 56-kW motors.

Slurry from the zinc regrind sump is pumped to Krebs gMAX regrind cyclones by two horizontal, centrifugal regrind cyclone feed pumps (one operating and one on standby) which are equipped with gland seal water and driven by 30-kW motors. Approximately 184 t/h of solids, at a slurry density of 31 percent, are pumped to the cyclones. Most of the coarser particles and some of the water in the feed stream are classified to the cyclone underflow stream. This stream, containing approximately 70 percent solids, flows to the zinc regrind mill at a rate of 190 t/h. The finer particles, and most of the water in the cyclone feed stream, are classified to the cyclone overflow stream which contains approximately 24 percent solids at about 219 t/h and flows to the zinc first cleaner conditioning tank, from which it flows by gravity to the zinc first cleaner circuits.

The regrind mill feed size is 80 percent passing 100 microns and the product is 80 percent passing 37 microns. The regrind mill is a vertical mill that uses steel balls as the grinding medium and is equipped with a coarse classifier fitted with an agitator and recirculating pump. The regrind mill feed enters top of the coarse classifier and partially separates into a coarse fraction and a fine fraction. The fine fraction overflows the coarse classifier and returns to the zinc regrind sump for circulation back through the zinc regrind cyclone cluster. The zinc regrind mill is driven by a 300-kW motor through a gear reducer.

The overflow slurry from the zinc regrind cyclone flows by gravity into the zinc first cleaner conditioning tank where zinc activator, pyrite depressant, collector, and promoter reagents are added and agitated. The zinc first cleaner conditioning tank has an effective volume of 11.9 m3 and a design slurry flow rate of approximately 219 m3/h. The conditioner discharges into the zinc first cleaner flotation circuit. The zinc first cleaner conditioning tank includes a rubber-lined, dual impeller agitator driven by a 7.5-kW motor.

The zinc first cleaner flotation circuit consists of a single line of seven Westpro flotation cells in series. Each cell has a volume of 5 m3 and is equipped with a rotor driven by a 15-kW motor. Low-pressure flotation air supplied by the three flotation air blowers is injected into the cells through the rotor shafts. Frother is added to the initial zinc first cleaner flotation cell, with additional frother and zinc cyanide added to the fourth cell. Froth that overflows from the zinc first cleaner flotation cells is collected in the zinc first cleaner flotation launder where it flows into the zinc second cleaner feed pump splitter box. Dart valves in the splitter box direct the concentrate slurry to one of two zinc second cleaner feed pumps. One of these 15-kW pumps is normally operating and the other is on standby. The design production for zinc first cleaner concentrate is 81.5 t/hour of slurry at 24.6 percent solids.

The zinc first cleaner concentrate is normally pumped by the operating zinc second cleaner feed pump to the first cell of the zinc second cleaner flotation circuit. However, provision is made to direct the concentrate from the first three first cleaner flotation cells to the third cleaner flotation circuit if is desired because of metallurgical or maintenance requirements. Provision is also made to direct the concentrate from the last four first cleaner flotation cells to the zinc regrinding sump.

The tailings from the last zinc first cleaner flotation cell flow by gravity to the zinc first cleaner scavenger flotation cells where the slurry is mixed with collector, promoter, and frother reagents. The design flow of zinc first cleaner scavenger tailings is 172.8 t/h of slurry at 23.6 percent solids.

The first cleaner scavenger circuit consists of three Westpro flotation cells, identical to the first cleaner flotation cells, arranged as a single bank. Concentrate from the first cleaner scavenger flotation cells flows into the zinc first cleaner scavenger flotation launder, which normally discharges into the zinc regrind sump. The design production of first cleaner scavenger concentrate is 16.3 t/h of solids at a slurry density of 25 percent. The zinc rougher conditioning tank feed pump discharges into a splitter box that has three dart valves which can direct the first cleaner scavenger concentrate to the zinc rougher conditioning tank, zinc rougher flotation cell, or zinc first cleaner conditioning tank.

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The tailings from the first cleaner scavenger flotation circuit flow by into the zinc first cleaner scavenger tailings sump. Slurry from the zinc first cleaner scavenger tailings sump is pumped to the zinc first cleaner scavenger tailings sampler. There are two zinc first cleaner scavenger tailings pumps (one operating and one on standby). These are variable-speed drive slurry pumps equipped with gland seal water driven by 18.65-kW motors. The slurry is then directed to either the zinc rougher conditioning tank or the zinc rougher tailings sump. The design production of first cleaner scavenger tailings is 172.8 t/h of solids at a slurry density of 22.8 percent.

As indicated previously, the concentrate from the zinc first cleaner flotation cells is pumped by the zinc second cleaner feed pumps to the zinc second cleaner flotation circuit to further raise the grade of the zinc concentrate. This circuit is composed of four flotation cells arranged in two banks of two cells each. The first bank of cells are 1.5 m3 Westpro cells identical to the first cleaner and first cleaner scavenger cells; the other two cells are Westpro cells with a capacity of 5 m3. The larger cells have 15-kW agitators and the smaller ones have 7.5-kW agitators. Concentrate slurry from zinc first cleaner flotation is pumped into the feedbox of the flotation cell, where frother is added. Tailings from the third cleaner circuit also flow into this feedbox. The cell agitation and the froth collection in the second cleaner flotation cells are similar to that in the other cleaner flotation cells.

The froth that overflows from the second cleaner flotation cells is collected in the zinc first/second cleaner flotation launder. The zinc second cleaner concentrate slurry flows from the launder into the zinc third cleaner feed pump splitter box, which directs the concentrate slurry to one of two zinc third cleaner feed pumps. One of these 11-kW froth pumps is normally operating and the other is on standby. The nominal zinc 2nd cleaner concentrate production is 14.3 t/h of solids at a slurry density of 25 percent.

Tailings from the second cleaner flotation flow by gravity back to the first cleaner flotation circuit for additional zinc recovery. The design flow of zinc second cleaner tailings is 32.9 t/h of slurry at 22.8 percent solids. The concentrate from the zinc second cleaner flotation cells is pumped by the zinc third cleaner feed pumps to the feedbox of the first flotation cell of the zinc third cleaner flotation circuit. Frother is also added at this location. The purpose of the third cleaner flotation circuit is to further raise the grade of the zinc concentrate. This circuit is composed of four 1.5 m3 Westpro cells with 7.5-kW agitators flotation cells arranged in two banks of two cells each.

The froth that overflows from the third cleaner flotation cells is collected in the zinc third cleaner flotation launder. The zinc third cleaner concentrate slurry flows from the launder into the zinc thickener feed sampler feedbox and then to the zinc concentrate thickener feed sump. Concentrate slurry is pumped from this sump by two vertical centrifugal concentrate thickener feed pumps powered by 11-kW motors. The design zinc third cleaner concentrate production is 51.5 t/h of slurry at 24.3 percent solids. Zinc third cleaner concentrate is the final product of zinc flotation.

Tailings from the third cleaner flotation flow by gravity back to the second cleaner flotation circuit for additional zinc recovery. The design flow of zinc third cleaner tailings is 8.5 t/h of slurry at 20.7 percent solids.

17.3.5.4Reagents — Zinc Flotation

Reagents used in the Escobal zinc flotation circuit include sulfide collector, promoter, frother, depressant, and activator reagents.

·Sulphide Collector — The sulphide collector used in the Escobal zinc flotation circuits is sodium isopropoyl xanthate (SIPX). This type of collector is generally referred to as xanthate. Dry xanthate is mixed to a solution strength of 10 percent and is added to the process through a circulating loop by means of flow control solenoid valves. The design consumption of xanthate in zinc flotation is 0.010 kg/t of mill feed.

·Promoter — FloMin C-4132 is used to promote flotation of zinc sulfide particles in the zinc flotation feed. This ready-to-use reagent is pumped from a storage tank through a circulating loop and is added to the process by metering pumps. The design consumption is 0.005 kg/t of mill feed.

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·Frother — The frother used in the zinc flotation circuits is Cytec AF-70 (methyl isobutyl carbinol). It is delivered to the flotation circuit without dilution. Individual meter pumps withdraw the frother from a supply header and deliver the reagent to various addition points in the flotation circuits. The design consumption is 0.100 kg/t of mill feed.

·Depressant — Pyrite that occurs in the zinc flotation concentrate is depressed by the addition of zinc cyanide slurry. Zinc cyanide is prepared by mixing sodium cyanide, zinc sulfate, and sodium hydroxide. The resulting zinc cyanide slurry is pumped to a storage tank from which it is distributed via a pipe loop to the various addition points throughout the plant, controlled by flow control solenoid valves. The design consumption of zinc cyanide is 0.025 kg/t of mill feed.

·Activator — Copper sulfate is used to activate zinc minerals during the zinc flotation process. At a sufficient dosage, the collection and subsequent flotation of zinc (sphalerite) is enabled. At higher than required dosages, the copper sulfate creates a brittle froth and gangue minerals float with the concentrate. The copper sulfate preparation and distribution system is identical to the zinc sulfate system and is added to the process by means of flow control solenoid valves. The design consumption of copper sulfate in zinc flotation is 0.030 kg/t of mill feed.

Figure 17-5: Escobal Mine — Flotation Plant

17.3.6Concentrate Thickening, Filtering and Packaging

Descriptions of concentrate thickening, filtering, and packaging are applicable to both lead and zinc concentrates, except where noted.

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17.3.6.1Concentrate Thickening

The concentrate thickening system increases the solids content of the concentrate slurry to approximately 65 percent solids by weight. Concentrate slurry flows from the third cleaner flotation area to the thickener feedbox, where it flows by gravity to the concentrate thickener prior to filtration. The overflow from the concentrate thickener flows by gravity to the concentrate thickener overflow standpipe and is then pumped to the tailings thickener feedbox in the tailings area for reuse. The concentrate thickener underflow is pumped to the concentrate splitter box where the flow can be recirculated back to the thickener or directed downstream to the concentrate tank, located in the concentrate filtration and packaging building.

The thickener feedbox collects the main stream of third cleaner flotation concentrate slurry after it passes through the concentrate thickener feed sampler and other plant process streams. The other streams collected by the feedbox are filtrate from the concentrate filtrate tank, flocculant from the flocculant supply circuit, concentrate splitter box (when the flow is recirculating), and concentrate thickener area sump pump. The combined stream flows by gravity to the concentrate thickener for processing. The material in the thickener feedbox discharges directly to the feedwell of the concentrate thickener.

There are 6-m-diameter concentrate thickeners for each of the lead and zinc concentrate slurries, each with rakes driven by an electric motor and gearbox combination. Thickening of the concentrate slurry involves separating the solid concentrate component of the slurry by gravity. Acceleration of the settling of the solids is achieved by the addition of flocculant, which is added to the slurry feed at the thickener feedbox. The flocculant addition rate is controlled manually by varying the flocculant distribution pump speed.

Thickened concentrate slurry is drawn from the bottom of the thickener by one of two concentrate thickener underflow electromechanical double diaphragm pumps. One pump is in operation while the other is on standby. The slurry density is measured on the discharge line of the thickener underflow pumps, where a density controller sends a remote set point signal to the thickener underflow flow controller, which utilizes a variable-frequency drive to control the speed of the pump. This, in turn, functions to control the density of the underflow. The slurry flow rate is measured on the discharge line of the thickener underflow pumps. The slurry density and slurry flow rate measurements are used to calculate the mass flow reading, which is displayed in the control room.

As the solids settle out of the slurry into the bottom of the thickener, clear liquid overflows the top of the thickener over a weir that encircles the top of the thickener tank. The clarified liquid flows over the weir and into a launder. The solution is collected at a low point in the launder and flows by gravity to the concentrate thickener overflow standpipe which receives overflow from both the lead concentrate thickener and the zinc concentrate thickener. The concentrate thickener overflow solution is withdrawn from the standpipe by one of two concentrate thickener overflow centrifugal pumps. One pump is in operation while the other is a standby spare. The concentrate thickener overflow pump delivers the clarified solution from the concentrate thickener overflow standpipe to the tailings thickener feedbox for reuse.

17.3.6.2Concentrate Filtration

The concentrate filtration systems remove sufficient amounts of moisture from the concentrate slurries in order to package and ship the concentrates as a relatively dry filter cake. Pressure filters remove water from the concentrate slurry to produce a filter cake containing approximately 8 percent moisture. The filter cake discharges from the pressure filter through a discharge chute to the appropriate stockpile.

The thickened concentrate slurry is directed to a filter feed screen from the concentrate thickener. The slurry passes through a screen to the concentrate tank where horizontal centrifugal slurry pumps draw the slurry from the tank and delivers it to the downstream concentrate pressure filter. One pump is in operation while the other is a standby spare.

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There are three automatic pressure filters installed in the Escobal plant, one each for filtering lead and zinc concentrates and a third standby unit available for either concentrate. The pressure filters are Diemme model GHT1200; the lead concentrate pressure filter is rated at 8.5 t/h feed rate and 5.2 t/h dry solids output and the zinc concentrate pressure filter is rated at 15.5 t/h feed rate and 9.7 t/h dry solids output. The standby unit is identical to the lead pressure filter.

Each pressure filter is equipped with filter plates and membrane plates that are configured to form chambers. Filter cloths, two for each chamber, are attached to each plate. The plates are suspended from a main structural member frame. The plates are moved laterally by a displacement device consisting of a set of special hooks assembled on chains that slide inside the beam of the filter press. This device allows smooth and uniform movement and precise spacing between the filter chambers when the press is opened.

The hydraulic system for the pressure filter consists of two independent hydraulic units. One unit is used to open and close the drip tray doors. The other hydraulic unit is used to open and close the filter by positioning the movable head and maintaining pressure on the plates after the initial closing of the pressure filter. The four hydraulic cylinders used to open and close the filter press move in parallel.

Dewatering/filtering the concentrate slurry in the pressure filters is performed as a batch process. When activated, a hydraulic unit moves a mobile head and plate pack toward the fixed header of the pressure filter. The pressure filter plate pack is closed and locked under high pressure by the hydraulic pump. Concentrate feed slurry is pumped into the filter chambers through feed ports. As slurry enters the chambers, solids are retained by filter cloths. Filtrate passes through the filter cloths and drains through the ports of each chamber. The concentrate feed pump continues pumping into the pressure filter until the appropriate feed pressure and flow rate are reached. When the filter cake is formed, it is stabilized by inflating a rubber membrane on one side of each cake, which compresses the filter cake. Process water flows into the feeding holes on each filter plate in the pressure filter which pushes the residual slurry that has accumulated during filtration through the feeding holes and to the lead filtrate tank. Compressed air pushes any residual washwater through the feeding holes and to the lead filtrate tank.

Compressed air is supplied to the filter cake, displacing the solution to the filtrate discharge on the opposite side of the cake. The filter cake is blown dry using compressed air. The hydraulic unit releases the pressure on the plate pack which allows the plate pack to move away from the fixed header. Filtrate leakages and residuals drop onto a drip tray on the bottom of the pressure filter. The drip tray hydraulic unit starts and opens the drip tray on the bottom of the pressure filter. After the drip tray has been opened, the hydraulic unit moves the mobile header back to its original position, opening the filter. As each chamber is opened, the dried filter cake falls down into a concentrate stockpile below the pressure filter deck.

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Figure 17-6: Escobal Mine — Concentrate Pressure Filter

17.3.6.3Concentrate Packaging

Stockpiled filter cake is reclaimed by a small front-end loader and delivered to a packaging and weigh station where it is placed in one- or two-tonne ‘super sacks’ and readied for shipment. The front-end loader reclaims concentrate from the stockpile and feeds it into a hopper where a screw conveyor draws the concentrate from the hopper and discharges it to a packaging and weigh system. The material discharging from the hopper is periodically intercepted by a sampler. Bagged concentrates are transported in 20-tonne sea containers by truck approximately 120 km to Puerto Quetzal, where the containers are loaded onto vessels for shipment to the Company’s smelter clients.

Figure 17-7: Escobal Mine — Concentrate Packaging

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Figure 17-8: Escobal Mine — Concentrate Loading

17.3.7Tailings Thickening, Filtration & Disposal

The tailings slurry from the zinc rougher flotation circuit is processed in a thickener to produce underflow slurry for the tailings filtration process. The tailings thickener feedbox collects zinc tailings, concentrate thickener overflow, reagents, and other process flows and delivers the combined flows to the tailings thickener, which reduces the water content of the feed slurry (approximately 25 percent solids) to produce slurry of approximately 60 percent solids to feed to the tailings filtration circuit. Settling of the thickener solids is accelerated by flocculant that is added at the tailings thickener feedbox. The tailings thickener is a 16-m-diamter WesTech high-rate thickener driven by an electric motor and gearbox combination.

Thickener overflow flows by gravity to a 42 m3 thickener overflow tank. The tank has a 14-inch discharge that flows by gravity to the process water pond, and a 6-inch discharge that reports to the tailings thickener overflow pump which pumps the solution to a water treatment system. Discharge lines from the underflow pumps merge and direct the underflow slurry to the tailings thickener splitter box. The tailings thickener box has two discharge ports, each with a dart valve that directs the tailings thickener underflow to recycle back to the tailings thickener feed box or to the filter feed tank in the tailings filtration system.

The thickened tailings slurry flows by gravity from the tailings thickener splitter box to a tailings filter feed tank. Three centrifugal pumps draw the slurry from discharge ports at the bottom of the filter feed tank and direct it to the tailings filters. Each pump is driven by a 336-kW variable-frequency motor capable of delivering 750 m3 of slurry per hour.

There are three Diemme GHT 2000 overhead beam pressure filters, each with a design throughput capacity of 85 wet t/h, that produce filter cake with a targeted moisture content of approximately 14 percent. Operation of the tailings pressure filters is similar to that described for the concentrate filtration system.

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Figure 17-9: Escobal Mine — Tailings Filtration

At the completion of each batch filter cycle, the dry filter cake drops from each tailings filter onto a dedicated discharge conveyor where it is transported to the tailings collecting conveyor. The tailings collector conveyor is a reversible conveyor driven by an 18.4-kW electric motor that directs the filter cake to either the paste plant conveyor or dry stack conveyor.

The dry stack conveyor delivers the filter cake to the dry stack radial conveyor, which discharges it onto a stockpile. The dry stack conveyor is a covered belt conveyor driven a 56-kW electric motor. The radial stacker is a McCord fixed-height stacker capable of handling approximately 440 t/h.

Figure 17-10: Escobal Mine — Tailings Conveyor to Dry Stack

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Figure 17-11: Escobal Mine — Dry Stack Radial Stacker

The paste plant conveyor system consists of two covered conveyors which transport the dry tailings from the tailings filtration plant to the paste backfill plant. The first conveyor, which receives dry tailings from the tailings collecting conveyor discharge, is one meter wide and 542 meters long, driven by a 56-kW motor. The dry tailings are then transferred to a second conveyor which is one meter wide and 542 meters long, driven by a 112-kW motor, and deposited into the feed hopper of the paste backfill plant.

Figure 17-12: Escobal Mine — Tailings Conveyor to Paste Plant

17.3.8Process Water

The Escobal mine uses mine discharge water, thickener overflow, and raw water to produce process water. Mine water reports to a 12-m-diameter thickener, where it is treated with flocculant. The mine water thickener overflow reports to the process water pond and the underflow is pumped to the tailings thickener feedbox. Concentrate and tailings thickener overflow also reports to the process water pond. The only water treatment required is the addition of

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hydrogen peroxide to the tailing thickener overflow to destroy cyanide. The process water does not exceed regulated contaminant levels, including those for cyanide and heavy metals. A water treatment plant is partially installed and can be brought into operation relatively quickly if it were to become necessary.

17.4CONCENTRATE PRODUCTION

Since the commencement of ore processing on September 30, 2014 through the end of June 2014, the Escobal mine produced 14,804 t of lead concentrates and 14,053 t of zinc concentrates, containing 12 million ounces of silver; 7,600 ounces of gold; 6,800 tonnes of lead; and 9,300 tonnes of zinc. Life of mine mill throughput and concentrate production through June 30, 2014 is summarized in Table 17-1.

Table 17-1: Mill Throughput and Concentrate Production (Life of Mine Through June 30, 2014)

Mill Production	Mill Feed	Pb Concentrate	Zn Concentrate
Tonnes	743,996	14,804	14,053
Ag Grade g/t	581	23,959	1,257
Au Grade g/t	0.47	14.70	1.27
Pb Grade %	1.0	44.7	1.4
Zn Grade %	1.4	13.5	51.9
Ag Ounces	13,888,635	11,403,376	568,014
Au Ounces	11,243	6,999	573
Pb Tonnes	7,528	6,611	202
Zn Tonnes	10,385	1,991	7,292

			Metal Recovery		Metal Recovery
Metal	Total Recovery		Pb Concentrate		Zn Concentrate
Ag	86.2	%	82.1	%	4.1	%
Au	67.4	%	62.2	%	5.1	%
Pb	90.5	%	87.8	%	2.7	%
Zn	89.4	%	19.2	%	70.2	%

The above table includes all process recovery data since the initiation of operations, including startup, commissioning and the ramp-up to full production. The mill has seen continuous improvements in metal recovery to concentrates since sustaining the design throughput rate of 3,500 t/d. For the second quarter of 2014, total recovery of silver, lead and zinc increased by about two percent, with the additional recovered metal reporting to the desired concentrate (i.e., silver and lead to the lead concentrate; zinc to the zinc concentrate), as summarized in Table 17-2. Work is ongoing to optimize metal recoveries.

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Table 17-2: Mill Throughput and Concentrate Production Apr 2014 - Jun 2014

			Metal Recovery		Metal Recovery
Metal	Total Recovery		Pb Concentrate		Zn Concentrate
Ag	88.1	%	85.0	%	3.1	%
Au	66.0	%	61.6	%	4.3	%
Pb	92.5	%	90.6	%	1.8	%
Zn	91.6	%	18.1	%	73.5	%

Process operations at the Escobal mine focus on maximizing the silver, gold and lead content and minimizing the zinc content of the lead concentrate; conversely, operations maximize the zinc content and minimize the silver, gold, and lead content of the zinc concentrate. Table 17-3 compares the payable metal recoveries predicted by the process design metallurgical testwork and the actual payable metal recoveries at the Escobal mine from the second quarter of 2014. Silver and lead recovery in the lead concentrate is 2.5% and 8% higher than indicated by the testwork, though a portion of the increase is predictable due to the differences in feed grade between the two sample sets. Gold recovery to the lead concentrate experienced in the plant is lower than predicted by the testwork. Zinc recovery to the zinc concentrate is 9% less than predicted by the testwork.

Table 17-3: Payable Metal Recovery Comparison — Design Parameters vs. Q2 2014 Actuals

Metal	Head Grade		Total Recovery		Metal Recovery Pb Concentrate		Metal Recovery Zn Concentrate
	Design	Actual	Design	Actual	Design	Actual	Design	Actual
Ag	415 g/t	657 g/t	86.7%	88.1%	82.5%	85.0%	4.2%	3.1%
Au	0.47 g/t	0.44 g/t	61.5%	66.0%	71.0%	61.6%	4.1%	4.3%
Pb	0.72%	1.22%	85.5%	90.6%	82.5%	90.6%	—	—
Zn	1.23%	1.65%	82.6%	73.5%	—	—	82.6%	73.5%

Grade-recovery curves for each metal and each concentrate were created from the current Escobal operational and metallurgical data to predict future metal recoveries in this study (Figure 17-12).

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Figure 17-13: Escobal Grade-Recovery Curves

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17.5PLANT EXPANSION

Additional capital requirements to expand the Escobal processing facilities from 3,500 t/d to 4,500 t/d include costs for upgrading the process plant infrastructure and ancillary equipment and the paste backfill plant. Estimated expansion capital costs for the process plant are summarized in Table 17-4.

Table 17-4: Plant Expansion Capital Expenditures

Area	Cost Estimate USD
Primary Crushing	$	649,900
Secondary/Tertiary Crushing	$	137,000
Grinding	$	473,300
Flotation	$	2,108,800
Concentrate Thickener	$	178,600
Tailings Filtration*	$	9,205,200
Tailings Dry Stack	$	475,000
Total	$	13,227,800

*$1,693,300 expended prior to July 1, 2014

The bulk of the expansion capital ($8.8M) is for the engineering, procurement, and installation of an additional tailings filtration unit to increase the capacity of the Escobal tailings filtration plant to handle the additional tailings generated from increased mill throughput. The filter selected is a Micronics LASTA pressure filter, which is capable of filtering 2,000 dry t/d at 90% availability. The Micronics LASTA filter was chosen for its throughput capacity and does not require significant expansion or modification to the existing tailings filtration building. Capital expenditures for the LASTA filter are summarized in Table 17-5.

Table 17-5: Tailings Pressure Filter Expansion Capital Expenditures

Description	Cost Estimate USD
Filter Plant Preparation	$	497,300
Equipment	$	4,500,500
Commissioning & Startup Spares	$	266,000
Construction & Shipping	$	578,300
EPCM	$	1,271,300
Commissioning	$	52,800
Contingency	$	1,648,200
Total	$	8,814,400	*

*$1,693,300 expended prior to July 1, 2014

The Company’s Board of Directors approved the expenditures for the additional tailings filtering capacity in May 2014, at which time engineering and equipment procurement was initiated. Installation and commissioning of the filter is expected to be completed at the end of the first quarter 2015.

The remaining expansion capital expenditures are for modifications and/or upgrades to ancillary process equipment, pumps, piping and motors necessary to handle the increased mill throughput of 4,500 t/d. The primary process equipment was originally sized for an anticipated increase in mill throughput; thus no replacement or upgrade of the major crushing, grinding, flotation, and concentrate handling equipment is necessary.

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Additional plant upgrades include:

·Primary Crushing — replacement or modification of motors to increase conveyor speed; modification of conveyor transfer chutes; increased dust suppression; and modification of feed discharge.

·Secondary & Tertiary Crushing — installation of air cannons or bin vibrators to better facilitate loading of conveyor to ball mill.

·Grinding — piping, pump and valve replacement; additional hydrocyclone distributor; pump replacement; upgrade particle size analyzer; and upgrade ball mill discharge box.

·Flotation — upgrade instrumentation; piping and valve replacement; installation of densimeter; installation of high-pressure water circuit; upgrade lead and zinc conditioning tanks; and relocation/installation of reagent dosifier, including pumps, tanks and lines.

·Concentrate Thickener — upgrade lead thickener pumps; upgrade recovery water pumps; additional zinc pressure filter pump; replace concentrate thickener sump pumps.

·Tailings Filtration — replacement of filter feeder pipes; upgrade sump pumps; upgrade hydraulic motors for conveyors; modification of conveyor chutes; and addition of process water filters.

·Tailings Dry Stack — additional Caterpillar 730 articulated haul truck.

Process plant upgrades are expected to be completed by mid-2015.

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18PROJECT INFRASTRUCTURE

This section summarizes the infrastructure and logistic requirements of the project. Figure 18-1 shows the site layout. Figure 18-2 shows an aerial view of the site layout with existing facilities.

Figure 18-1: General Site Layout

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Figure 18-2: Site Map With Existing Facilities

18.1SITE ACCESS

The project is approximately 2 kilometers from San Rafael Las Flores, a town of 3,500 people, and approximately 70 kilometers by paved highway from Guatemala City. All year access to the area is good via paved highway from Guatemala City.

Service and haul roads have been surfaced with gravel and bermed as necessary.

18.2TRANSPORTATION AND LOGISTICS

The major process and mining equipment will be procured overseas and shipped to Guatemala. No special handling requirements are foreseen, and normal shipping routes and ships can be utilized. Logistics to date has not proven to be problematic.

Guatemala has ports on both the Pacific and the Caribbean coasts. Access to the mine site from both ports is on paved highway.

Filtered concentrate is placed in 1,000 to 2,000 pound super-sacks, placed in sea-going containers, and carried on highway tractor trailer units along paved highway to either port for shipment to international smelters.

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18.3ELECTRICAL POWER

Electrical power for the Escobal mine is provided by on-site diesel generation capable of producing a maximum of 21.5 megawatts (MW). The estimated peak power load required for the operation is approximately 12 MW during startup of high-consumption equipment such as the ball mill. Normal operating requirement for mining, process and surface operations is 7.5 to 8.5 MW. Contractor supplied and maintained diesel generators with a capacity of 15.5 MW provide power for all operations; the Company owns four auxiliary diesel generators capable of producing 5.5 MW as backup power for critical equipment in the event several contracted generators were to fail.

The Company is actively investigating alternative lower-cost sources of power generation such as conversion from diesel to heavy fuel oil (HFO).

18.4WATER AVAILABILITY

Raw water and process water for mine operations is supplied from mine dewatering and from surface dewatering wells. Domestic water is supplied by dedicated domestic water wells. The use of groundwater to support operations is provided for in Article 71 of the Guatemalan Mining Law. Hydrological studies indicate sufficient water will be available to supply process and potable water requirements for the life of mine (Reidel, et.al. 2014).

18.5SITE DRAINAGE

The plant site facilities were designed and located such that the upstream primary natural watershed was not significantly diverted. Only the portion of the drainage near the operational facilities was realigned into a concrete-lined channel. The drainage returns to its original channel prior to exiting the mine site. The avoidance of diverting this major watershed reduced the overall area of disturbance and maintained the historic flow of water through the property.

18.6TAILINGS AND WASTE ROCK FACILITY

18.6.1Overview

The Escobal mine tailings facility is designed and operated as a dry-stack, in which dewatered tailings are spread, compacted and graded for erosion control and stability. Dry-stacking of tailings was selected and is implemented at the Escobal mine as it is an effective way to create a safe facility that will, upon closure, become a long-term stable geomorphic form in the landscape.

Dewatering equipment at Escobal is the plate and frame filter press, a proven technology used by many industries for nearly 100 years. The water content of the tailings is reduced from 40 percent to about 15 percent by weight through this process. After filtering, the dewatered tailings are transported to the tailings storage facility and compacted in relatively thin lifts. Mine development rock is placed and compacted as engineered fill within the dry stack area

Geochemical testing establishes that the tailings are non-acid generating in the long term. Specifically, testing establishes that the chemical composition of the tailings is acid neutralizing and that effluent chemical components in the tailings are in compliance with Guatemalan permissible discharge limits as required per Government Agreement 236-2006, Wastewater Discharge and Reuse and Sludge Disposal Rulings, as well as those set by EPA and the World Bank standards.

Before tailings or development rock placement, the area of the dry stack footprint is prepared by removal and stockpiling of organic topsoil for use in final reclamation. The foundation is further prepared by construction of foundation key, subdrains, and perimeter development rock starter buttress fills.

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The dry stack tailings facility will be constructed and operated over the mine life, which is estimated to be about eighteen to twenty years. Operation includes receiving the filter-pressed tailings from the plant, hauling and placement of tailings, compacting the tailings, and constructing additional surface water management facilities to accommodate the rising tailings.

Upgradient surface water control facilities divert and control upgradient, non-contact surface water runoff; this runoff is discharged from site through existing surface water control channels. Contact surface water is diverted to settling ponds prior to release. Shallow downslope-migrating groundwater and infiltrating surface water that contacts the dry stacked tailings is intercepted by a subsurface central spine and key trench drains. Chemical analyses demonstrate that the tailings contact water and effluent, if any, are suitable for discharge from site through existing surface water control channels. The tailings facility design includes the provision for capture, storage and pumping of collected infiltrating and migrating groundwater, tailings effluent, and contact water to treatment facilities, if necessary.

Ongoing reclamation of the dry stack is being undertaken during construction. As successive lifts of dewatered tailings are placed and compacted behind perimeter rock berms (buttresses), the lower front-face rock slopes are covered with stockpiled topsoil and revegetated. Thus at all stages of operation, the outer slope of the dry stack is, in essence, a vegetated slope that replicates natural slopes in the immediate vicinity of the mine.

The static and seismic stability of the dewatered tailings and the overall dry stack is enhanced by the facts that: (a) the tailings are dewatered to a point that they are not potentially liquefiable; (b) the tailings are compacted in place to increase density and strength; and (c) upper surface compaction limits infiltration of rain water into the tailings.

Drainage of shallow, perched downslope migrating water in the loose to medium dense (upper ten to 12 meters) ash foundation soils is achieved by an extensive system of deep subdrains. Thus the dry stack tailings and drained underlying dense subsurface soils are not be susceptible to earthquake-induced liquefaction.

Earthen materials such as the dry stack and the foundation soils are not perfectly rigid structures, and some deformation may be anticipated over the life of the facility, particularly in moderate to large (maximum credible) seismic shaking events. Whether due to the anticipated maximum credible earthquake, or lesser shaking events, as defined by probabilistic and deterministic site-specific ground motion studies, the dry stack facility, at full height of the stack and in the long-term post-closure period, is anticipated to remain in a stable configuration. The deformation of the dry stack due to postulated cyclic shaking scenarios is computed to be within tolerable limits and any slumping of outer rock face buttresses may be readily regraded, if necessary.

At mine closure, the dry stack will be essentially similar to the natural slopes at and in the vicinity of the mine: namely a natural landform of dry, dense soils that supports a natural stand of vegetation and cultivated crops, such as re-planted coffee trees. Thus a new geomorphological form is created that will, in the long term, respond to natural geomorphic forces as do the surrounding slopes and hillsides.

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Figure 18-3: Cross Section of the Tailing Dry Stack (End of Mining)

18.6.2Layout

As shown on Figure 18-4, the dry stack includes the following components:

·Conveyor: this is used to convey the dewatered tailings from the filter-press plant to the stacker pad.

·Stacker: this places the tailings in piles that are loaded into trucks that transport the tailings to the area of the stack where placement is ongoing.

·Front Face of the Dry Stack: the face of the dry stack rises at an inclination of three horizontal to one vertical (3:1) to a height sufficient to accommodate the tailings produced during operation of the mine.

·Front-Face Starter Buttresses: nominal 5-meters high front face starter buttresses constructed with mine development rock are placed ahead of tailings placement to form a stable perimeter control feature. Every third 5-meter high rock buttress is set back nominally 5 meters: these intermediate benches are sloped to drain north at nominal 1%.

·Access Road: this advances up the front face of the dry stack to provide passage for trucks carrying the tailings from the conveyor-stacker pad to the working deck of the facility. This haul road is also be used by trucks conveying mine development rock to build the outer berms and topsoil cover of the facility during concurrent reclamation and vegetation of the outer (front face) slopes.

·Perimeter Toe Buttress: these was constructed within the dry stack footprint at the toe areas, flanking the north, west and south facing toe of slopes of the facility.

·Contact Surface Water Controls: the upper deck of the active working area of the stack are sloped to drain north to northeast at a nominal 1 percent to hardened contact surface water control channels.

·Non-tailings Contact Surface Water Management Facilities: these have been and will continue to be constructed to intercept upgradient runoff and divert such water around, beneath, and away from the stack.

·Subsurface Water Management Facilities: in order to enhance the geotechnical stability and to collect tailings effluent for either direct discharge or routing to water treatment facilities, a series of low-fines, free-draining rock-filled trenches was constructed below grade at the south toe of the dry stack.

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Figure 18-4: Dry Stack Layout (End of Construction)

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Figure 18-5: Ongoing Construction During Tailing Placement

18.6.3Ongoing Monitoring and Testing

Every two to three years, or about 10 to 12 m height increase, a geotechnical performance review of the geotechnical properties of the tailings and the likely geotechnical response of the dry stack to both static and potential dynamic (seismic) loadings would be undertaken.

Concurrent reclamation of the dry stack is ongoing. As successive lifts of filter-pressed tailings are placed behind the perimeter rock berms, the lower slopes are revegetated. Thus at all stages of operation, the outer slope of the dry stack will be, in essence, a vegetated slope that replicates the natural slopes at and in the immediate vicinity of the mine. Every year a performance review of the contact and non-contact surface water facilities will be undertaken at the dry stack.

Figure 18-6: Example of Ongoing Reclamation Efforts

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19MARKET STUDIES AND CONTRACTS

Tahoe’s management team has extensive experience marketing lead and zinc concentrates containing precious metals to smelters throughout the world. The Company relied upon its collective experience to assess the marketability of the Escobal concentrates and execute smelter contracts with its customers.

Three years prior to commencement of operations at Escobal, smelters with precious metal refining capacity were identified as potential concentrate customers. Although silver is the most significant source of revenue in concentrates, the lead and zinc content required these concentrates be treated at lead and zinc smelters that demonstrate high levels of precious metal recovery and have integrated precious metal refining capacity. This requirement limited the number of smelters that could reasonably be considered as customers for the Escobal concentrates. Preliminary estimates of concentrate quality based on metallurgical tests on drill core were presented to these smelters for evaluation.

Samples of lead and zinc concentrates produced from the flotation pilot plant testwork (as described in Section 13.7.1) were distributed to potential smelter customers for evaluation. Upon completion of evaluation, several smelters indicated interest in purchasing substantial quantities of the concentrate, which exceeded planned production levels. Additional smelter companies have since processed test parcels and entered into long term contracts.

Concentrate production from the Escobal mine has been committed under long-term frame agreements with multiple lead and zinc smelter customers for which terms will be renewed annually to reflect market conditions for high precious metal lead and zinc concentrates. Average payable metal percentages (after minimum deductions) and refining charges per payable ounce of silver realized for contracts currently in place are summarized in Table 19-1.

Table 19-1: Average Payable Metal Percentages and Refining Charges

	Lead Concentrate				Zinc Concentrate
Metal	Payable %		Refining Charge		Payable %		Refining Charge
Ag	96	%	$	1.13	66	%	$	0
Au	92	%	$	10.00	56	%	$	0
Pb	94	%	n/a		0		n/a
Zn	0		n/a		84	%	n/a

Treatment charges are at or near annual benchmark terms. Based on contract terms, the Company estimates an average maximum penalty charge of less than $8.00/t of lead concentrate and less than $1.00/t of zinc concentrate for deleterious elements. Small premiums are granted to compensate some smelters for the handling of bags rather than bulk shipments. Refining charges are significantly below market levels for common silver-bearing concentrates reflecting the higher grades and relative value per tonne of Escobal concentrates.

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20ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

Based on the results of environmental baseline studies, ongoing site monitoring data, socioeconomic studies, and permitting and licensing requirements, no issues have been identified that could materially impact the Company’s ability to conduct operations at the Escobal mine as discussed in this Study.

20.1ENVIRONMENTAL STUDIES

20.1.1Baseline Environmental Study

Baseline environmental studies were conducted as part of the Environmental Impact Assessment (EIA) for the Escobal mine (further discussed below in Section 20.4.2) prior to any surface disturbance related to the operation.

Baseline environmental studies included:

·Ambient air quality;

·Ambient sound levels;

·Surface water flows and chemistry;

·Groundwater chemistry;

·Soil and subsoil geochemistry;

·Flora and fauna surveys; and

·Cultural and historical surveys.

The EIA identified no areas of study where the Escobal mine operations would have long-term negative impacts to the environment provided proper environmental prevention and mitigation measures are adhered to during construction, operations and closure.

20.1.2Geochemical Characterization

Laboratory tests conducted prior to, and concurrent with, the underground excavation and ore processing demonstrated that the waste rock and tailings generated by the Escobal mine are net neutralizing rather than acid generating due to the relatively low sulfide content and high carbonate content of the rock. Laboratory tests included Acid-Base Accounting (ABA), which determined the net neutralizing potential relative to the acid generating potential; Meteoric Water Mobility Procedures (MWMP), which determined the content of dissolved metals and other elements in waste rock and tailings effluent; and humidity cell tests (HCT) to determine the long-term chemistry of waste rock and tailing effluent.

The results of these investigations demonstrate the waste rock and tailings from the Escobal mine do not generate acid nor contain heavy metals or other deleterious elements above regulatory limits in their effluent. The waste rock and tailings present no adverse impacts to the environment.

Geochemical testing programs for development waste rock were conducted by Goldcorp/Entre Mares in 2009 and by Tahoe in 2010 and 2011, continuing through the present. All acid rock drainage (ARD) tests (ABA, MWMP, HCT) were performed by independent laboratories. McClelland Laboratories Inc. (Sparks, Nevada USA) conducted all ARD testing for Tahoe; Goldcorp’s tests were performed by SVL Analytical Inc. (Kellogg, Idaho USA).

20.1.2.1Acid-Base Accounting

Goldcorp/Entre Mares and Tahoe collected a total of 47 drill core samples for ABA testing from locations on the footwall and hanging wall of the Escobal Central Zone and East Zone at various intervals along strike and from elevations that

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are, or will be, excavated by the underground development. The samples represented all rock and alteration types expected to be encountered in the development.

In general, the degree in which sample material demonstrates acid generating potential (AGP) or acid neutralizing potential (ANP) is indicated by the Net Neutralizing Potential (NNP), where NNP= ANP— AGP. Of the 47 waste rock samples tested, only one sample had an NNP value less than zero. The average NNP value for all samples is 87.

Tahoe submitted nine samples of tailings obtained from the flotation and pilot plant metallurgical tests for ABA testing. All of these samples returned positive NNP values, with an average NNP of 87.9.

A summary of the ABA test results is shown in Table 20-1.

Table 20-1: Acid-Base Accounting Results

		Number of	Net Neutralizing Potential
Company	Sample Type	Samples	Mean	Minimum	Maximum
Goldcorp / Entre Mares	Waste Rock	27	87.4	-5.3	286.9
Tahoe / MSR	Waste Rock	20	86.5	23.1	239.0
	Tailings	9	87.9	78.9	106.12

20.1.2.2Meteoric Water Mobility Procedures

McClelland Laboratories also conducted Meteoric Water Mobility Procedure (MWMP) tests on each of the samples submitted by Tahoe for ABA tests. The purpose of the MWMP is to evaluate the potential for dissolution and mobility of metals and other constituents from waste rock and tailings by meteoric water.

The tests were conducted following ASTM E2242-02 procedures and methodology and consisted of 24-hour column leach of 5 kg of crushed mine waste samples (minus 2-inch) and final tailings material using an extraction fluid ratio of 1:1. The extraction fluid is Type II reagent grade water which simulates meteoric water in terms of composition and pH range. The pH of meteoric water influent for waste rock samples ranged from 5.38 to 5.75. The pH of meteoric water influent for tailing samples ranged from 5.26 to 5.5.

Results of the MWMP tests are summarized in Table 20-2. The results demonstrate the waste rock and tailings from the Escobal mine do not leach metals or other deleterious constituents and also show that both materials acted as a buffer as the pH of the effluent is greater than the pH of the meteoric water influent.

Table 20-2: Meteoric Water Mobility Procedure Results

		Number of	Effluent pH
Company	Sample Type	Samples	Mean	Minimum	Maximum
Tahoe / MSR	Waste Rock	20	7.12	6.34	7.80
	Tailings	9	7.93	7.84	8.05

20.1.2.3Humidity Cell Tests

Tahoe selected six waste rock samples with the lowest NNP value and all tailings samples used for the ABA tests and submitted them for long-term humidity cell tests (HCT). McClelland Laboratories conducted the HCTs following ASTM D-5744 procedures.

Humidity cell tests are kinetic tests designed to model the atmospheric and geological processes of weathering. The purpose of the test is to determine if acid generation will occur and, if so, at what rate. The test also quantifies the

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quality of leachate water over time. HCTs are performed to confirm or reduce the uncertainty in the results of static prediction tests, such as ABA tests. Humidity cell tests will determine if the material will become acidic, but not when the material will become acidic, as the operation of the humidity cell accelerates sulfide mineral oxidation.

HCTs on the waste rock and three tailings samples were run for 120 weeks; six tailings samples were run for 100 weeks. Effluent pH was sampled weekly. None of the samples became acidic over the duration of the tests.

Table 20-3: Humidity Cell Effluent pH

		Number of	Effluent pH
Company	Sample Type	Samples	Mean	Minimum	Maximum
Tahoe / MSR	Waste Rock	6	7.89	7.18	8.87
	Tailings	9	7.49	6.66	8.78

The HCTs accelerate the rate of oxidation and acid production. This results in an accelerated rate of oxidation products generated as dissolved metals and/or precipitated metal compounds. The metal concentrations in the HCT leachate are considered to be higher than those generated in the field; as such, the HCT results often indicate a “worse case” scenario. HCT effluent was sampled every four weeks of the test period and analyzed by ICP.

HCT effluent analyses of the waste rock samples are summarized in Table 20-4. The effluent chemistry results for all constituents in all of the samples are below regulatory permissible limits prescribed by Acuerdo Gubernativo No. 236-2006 (Government Agreement No. 236-2006), which stipulates water discharge standards.

Table 20-4: Humidity Cell Effluent Analysis — Waste Rock (mg/L)

Parameter	Mean	Minimum	Maximum	236-2006
Alkalinity, CaCO3	38	19	100	—
Antimony	0.010	<0.012	0.025	—
Arsenic	0.003	<0.0050	0.033	0.1
Cadmium	0.000	<0.0010	0.002	0.1
Calcium	17.4	8.6	85.0	—
Chloride	0.2	<1.0	14.0	—
Copper	<0.050	<0.050	<0.050	3.0
Iron	0.004	<0.010	0.290	—
Lead	0.001	<0.0025	0.019	0.4
Magnesium	1.45	<0.50	8.00	—
Manganese	0.057	<0.0050	0.540	—
Mercury	<0.00010	<0.00010	0.0001	0.01
Nitrate as N	0.01	<1.0	1.80	—
Nitrite as N	0.02	<0.025	3.00	—
pH	7.64	6.58	8.58	6.0 - 9.0
Potassium	3.1	<2.5	30.0	—
Selenium	<0.0050	<0.0050	<0.0050	—
Sodium	0.7	<0.50	22.0	—
Sulfate	17.0	<1.0	180.0	—
Total Dissolved Solids	78	15	500	—
Zinc	0.002	<0.010	0.04	10.0

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HCT effluent analyses of the tailings samples are summarized in Table 20-5. The analytical suite was the same as for the waste rock samples, but with WAD cyanide, free cyanide and total cyanide included in the analyses. The effluent chemistry results for all constituents in all of the tailings samples are below regulatory permissible limits.

Table 20-5: Humidity Cell Effluent Analysis — Tailings (mg/L)

Parameter	Mean	Min	Max	236-2006
Alkalinity, CaCO3	8	5	16	—
Antimony	0.0005	<0.0025	0.0080	—
Arsenic	0.0004	<0.0050	0.0037	0.1
Cadmium	0.0005	<0.0010	0.0030	0.1
Calcium	13.8	2.7	43.3	—
Chloride	0.4	<1.0	4.7	—
Copper	<0.050	<0.050	<0.050	3.0
WAD Cyanide	0.002	<0.010	0.029	—
Free Cyanide	0.009	<0.010	0.101	—
Total Cyanide	0.004	<0.010	0.033	1.0
Iron	0.01	<1.0	0.03	—
Lead	0.0002	<0.050	0.0022	0.4
Magnesium	1.58	0.07	6.48	—
Manganese	0.22	0.09	0.71	—
Mercury	<0.00010	<0.00010	0.00001	0.01
Nitrate as N	0.0003	<1.0	0.0111	—
Nitrite as N	0.001	<0.025	0.015	—
pH	6.83	6.55	7.27	6.0 - 9.0
Potassium	1.06	<2.5	6.93	—
Selenium	<0.0050	<0.0050	0.0014	—
Sodium	0.39	<0.50	3.61	—
Sulfate	33.0	4.2	120.9	—
Total Dissolved Solids	68	14	230	—
Zinc	0.010	0.001	0.046	10.0

20.2SITE MONITORING

The Company conducts continuous environmental monitoring programs for air quality (particulate matter, combustible gases), surface and subsurface water quality, stream sediment chemistry, blast vibration, sound pressure, rock geochemistry, and monitoring for occupational health and safety standards. Environmental compliance is regulated by the Ministry of Environment and Natural Resources (MARN), who follows guidelines set forth by the World Bank, International Finance Corporation, World Health Organization, the US Environmental Protection Agency, and US Occupational and Health Administration, in addition to Guatemalan compliance standards.

In addition, Minera San Rafael practices topsoil conservation and management, concurrent reclamation where practical, reforestation of previously disturbed areas, strict waste disposal procedures, and environmental training for employees.

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20.2.1Air Quality Monitoring

Nine air quality monitoring stations are employed within the industrial area and within nearby communities to continually measure concentrations of particulate matter (dust), seven of which also take continuous measurements of metal particulates, total settleable solids, and combustible gases (SO2, NOx). Three stations within the plant site also monitor for hydrogen sulfide (H2S) gas (occupational health & safety).

20.2.2Water Quality

Streams, rivers, springs, and wells within the project area and the surrounding regional watershed are routinely sampled for organic and inorganic constituents (dissolved metals, hydrocarbons), as well as tested for pH, electrical conductivity, dissolved oxygen, temperature, and dissolved solids. Water analyses are conducted by both US and Guatemalan laboratories. The Company uses certified analytical standards and blanks to ensure quality assurance and quality control at all laboratories. The Escobal mine employs water conservation practices such as recycling process water, the use of filtered dry stack tailings storage as opposed to a conventional tailing pond, and other water management systems to minimize water consumption and maintain water quality.

20.2.3Surface Water

There are eleven surface water sampling sites to monitor the water quality of streams and rivers within the project area, downstream of the project area, and elsewhere in the regional watershed. Five naturally occurring springs identified within the project area and surrounding area are regularly sampled and monitored for water quality. Rainfall and runoff that comes into contact with the industrial site is channeled to an engineered retaining pond system for use in the milling process or for sampling and treatment (if necessary) prior to discharge to the environment. Rainfall and runoff that does not come into contact with the industrial site is channeled into natural drainages.

20.2.4Ground Water

Twelve monitor and/or production wells located within the mine property are regularly sampled for water quality. Water that is pumped to the surface from the underground mine is recycled for use in the mine or process plant or sampled and treated (if necessary) prior to discharge to the environment.

20.2.5Process and Other Industrial Use Water

Process water is held in an engineered lined containment pond and recycled for use in the milling process. Water from the tailings filtering process is collected and returned to the process water pond. Precipitation and runoff upstream of the dry stack tailing facility is collected in channels and routed around the facility for discharge to the environment (non-contact water). Precipitation and runoff that comes in contact with the dry stack facility is collected in channels, sampled, and directed to either the process water pond or to the storm water contact ponds. Dry stack effluent is sampled prior to discharge to the environment. Sewage treatment plant effluent is regularly sampled to ensure water quality prior to discharge to the environment.

20.2.6Sediment Sampling

There are eleven sampling sites to monitor the quality of stream and river sediment within the project area, downstream of the project area, and elsewhere in the regional watershed.

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20.2.7Vibration Monitoring

There are three permanent surface vibration monitoring locations (seismograph installations); two are located in proximity to the neighboring communities of Los Planos and La Cuchilla, with the other located midway between the two underground portals. The seismographs continually record particle velocity and air pressure changes that result from underground blasting. All blasts are monitored. Seven additional locations within the property and surrounding communities are monitored intermittently.

20.2.8Noise (Sound Pressure) Monitoring

Nine sound monitoring stations are used to continually measure sound levels within the industrial site and within nearby communities.

20.2.9Rock Geochemistry (Acid Rock Drainage Monitoring)

Field paste pH tests to determine the acid generating or acid neutralizing potential of waste rock are continually conducted as the underground development advances. Likewise, paste pH tests are also conducted on the final tailings. Paste pH results are verified by independent laboratory ABA tests.

20.2.10Waste Disposal

All solid and liquid waste, as well as spent hydrocarbons (oil, grease, etc.) are disposed of by an approved waste handling contractor and deposited in a certified waste receiving facility.

20.2.11Reagent Storage

All chemicals and processing reagents are stored in facilities designed and constructed to provide containment in the event of a spill to prevent discharge to the environment. All employees who work with chemicals or processing reagents undergo routine safe handling and environmental training.

20.3WASTE ROCK AND TAILINGS DISPOSAL

Tailings from the Escobal mine are filtered and deposited in a dry stack tailings facility. Development waste rock is used to construct the structural buttress of the dry stack. Environmental monitoring of the dry stack includes paste pH and ABA tests on the rock buttress material and final tailings product. Monitoring also includes sampling of the dry stack underdrains. Installation of down-gradient alluvial and bedrock monitor wells is scheduled for 2015. A description of the dry stack tailings facility and water management controls is discussed in Section 18.6.1.

20.4PERMITS AND APPROVALS

Escobal mine operations and surface exploration activities within the Escobal exploitation concession are primarily permitted and regulated through two Guatemalan federal agencies, the Ministry of Environment and Natural Resources (MARN) and the Ministry of Energy and Mines (MEM). All permits for mine operations and surface exploration are in place.

20.4.1Exploration

Environmental requirements for surface exploration activities are specified in Resolution 4590-2008/ELER/CG, which was issued by MARN to Entre Mares in December 2008. These requirements and approvals were transferred from Entre Mares to the Company in September 2010 as specified in MARN Resolution 1918-2010/ECM/GB. Approval of surface exploration activities includes environmental commitments and environmental work plans.

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20.4.2Mine Operations

The Escobal mine operations are conducted under an Environmental Impact Assessment (EIA) approved by MARN and an Exploitation License issued by MEM.

The Company submitted an EIA for the construction, operation and closure/reclamation of the Escobal mine to MARN in August 2011. The EIA documented baseline environmental and socioeconomic environment of the Escobal property and surrounding area; provided an analyses of the project’s potential impacts to air quality, sound levels, surface and groundwater, soils, flora and fauna, cultural and historical resources, and socioeconomic conditions; and an Environmental Management Plan which included comprehensive prevention and mitigation measures. The EIA was prepared by Asesoría Manuel Basterrechea Asociados, S.A. (2011). Public disclosure and involvement was required and developed throughout each stage of the permitting process. Approval of the EIA was received from the ministry in October 2011 (Resolution 3061-2011). The approval of the EIA allowed the company to begin full construction of the mine, process plant and all other surface facilities. The Company files quarterly environmental monitoring reports with MARN as required by the Resolution.

Upon approval of the EIA, the Company posted an observance bond in the amount of Q8M (approximately US$1.05M) to MARN. This is not a reclamation bond per se, but rather, the bond is in place in the event of the Company’s failure to observe the mitigation measures stated in the EIA and required by the Resolution.

Mineral production in Guatemala is licensed through MEM. Application for the Escobal Exploitation License was submitted to MEM in November 2010 and approved by the ministry in April 2013 (License no. LEXT-015-11). The Company files annual reports with MEM as required by the license stipulations.

Land use changes and tree cutting are permitted through Guatemala’s National Institute of Forests (INAB). Archeological clearances were issued by the Ministerio de Cultura y Deportes (Ministry of Culture and Sports).

The export of concentrates from the Escobal mine is licensed through MEM, with annual renewal requirements. The Company’s export license (EXPORT-TI-17-2014) is current and valid.

20.5SOCIAL OR COMMUNITY IMPACTS

The Escobal mine is located about two kilometers from San Rafael Las Flores, a community of approximately 3,500 inhabitants. According to Guatemala’s National Institute of Statistics (Census 2002) San Rafael Las Flores’ population is 99.6% “Ladino”, i.e., of Hispanic origin and non-indigenous. The area surrounding the community, including several small villages, is generally used by local farmers to grow vegetables in the valleys and coffee at higher elevations. Other than the Company’s activities, there is no heavy industry in the immediate area. The Company recognizes the impacts to the community’s infrastructure due to increased industrial activities and the related influx of the growing workforce and is working directly with community leaders and community groups to minimize any potential negative impacts and maximize the numerous benefits related to the mine for the betterment of the community and surrounding areas. With few exceptions, community support for the mine is strong and the Company is committed to being an active and positive member of the local community.

Tahoe is generally aligned with the Equator Principles (EPs), Guiding Principles on Business and Human Rights (Ruggie Principles) and Voluntary Principles on Security and Human Rights. To strengthen the Company’s alignment with the EPs, Tahoe conducted a Social Impact Assessment (SIA) to identify the social impacts associated with the Escobal mine. To complement the SIA and avoid, minimize or mitigate real or perceived social impacts in the future, Tahoe created a Social Management Plan (SMP) outlining the Company’s strategies to meet its corporate social responsibility (CSR) commitments. In 2014 the Company strengthened its existing grievance mechanism to generally align with standards described in the EPs, Ruggie Principles and International Finance Corporation’s Performance Standards. The grievance procedure is intended to prevent and address human rights impacts linked to Tahoe’s

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business activities. It includes a stakeholder management system that facilitates timely responses to stakeholder concerns and ensuring community issues are addressed in an equitable, timely and consistent manner.

Approximately 850 full-time workers are currently employed at the Escobal mine. More than 95% of mine employees are Guatemalan, with more than half of these employees from surrounding communities. The Company’s Sustainable Development Department in San Rafael Las Flores works diligently with local and regional communities to promote development projects, build capacity for municipalities and landowners to responsibly use royalty payments and communicate meaningful information about the mine. The Sustainable Development Department has actively engaged with local and regional communities since 2010 when the Company started activity in the area, making more than 1,000 informative home visits, hundreds of dialogue meetings with key stakeholders, and providing tours of the Escobal mine to more than 4,500 interested citizens.

In 2013 the Sustainable Development Department partnered with local and regional communities to implement a series of CSR initiatives. Key projects included:

·Vocational Training Center: A 650 square meter training facility with three workshop areas and two equipped classrooms was inaugurated in 2013. Currently the center is providing sewing, entrepreneurship, computer skills and English as a second language lessons. Planned program offerings include training in other trades and technical careers such as welding, auto mechanics and electro mechanics.

·Regional Coffee Growers: The Company supported 1,550 coffee growers with a coffee rust control program including 600 coffee growers from San Carlos Alzatate and 950 coffee growers from six regional municipalities of Santa Rosa. The Company also donated fungicide to farmers and provided training sessions and technical visits.

·Reforestation Program: Over 100,000 small trees were planted in the reforestation program with the support of around 500 volunteers. Approximately 230 hectares of land were reforested and forest management plans were formulated to promote forest-growth incentives for local landowners.

·Education Projects: Over 300 first grade teachers in seven municipalities of Santa Rosa were trained in reading and writing teaching skills in a joint effort between the Company and the Ministry of Education. The Company and Del Valle University completed teacher training and provided classroom implementation monitoring and evaluation for natural science lessons under the national school curricula.

·Community Infrastructure Projects: The Company has initiated, funded, and supported multiple infrastructure projects in communities throughout the region, including El Copante (health clinic); Media Cuesta (classroom and restrooms); La Cuchilla (fencing, soccer field); Casco Urbano (multifunction classroom); Las Cortinas (water road drainages); Quequexque (stone road); Estanzuelas (multifunction classroom and school upgrades underway); Colonias Unidas (water system); El Chan (school upgrades and multifunction classroom underway); Casco Urbano (upgrades to middle school in collaboration with Glasswing International); Los Planes’ new elementary school; and El Copante school bathrooms and multifunction classroom.

·Health Projects: The Company supports the Technical Institute of Training and Productivity (INTECAP) in training 298 health employees in teamwork, patient attention, motivation at work, and food security.

In 2014, to support food and nutrition needs for MSR’s local and regional families, the Company donated $2.3 million to support training and provide educational materials to the national Mejores Familias program. The program teaches chronic malnutrition prevention strategies to over 92,000 women in the neediest Guatemalan departments. The donation will also help support a chronic childhood malnutrition prevention program and improvements to health infrastructures, such as the San Rafael Las Flores Health Center, Water Treatment Plant and the Santa Rosa

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Malnutrition Rehabilitation Center. The Company is looking for additional opportunities to help alleviate hunger throughout Guatemala.

A significant portion of the production royalties paid by the Company are directed to the local communities. In addition to the one percent NSR royalty mandated by the Guatemalan mining law, the Company entered into a voluntary royalty agreement which commits the Company to pay an additional four percent NSR royalty on the concentrates sold from the Escobal mine. Of the total royalty of five percent, two percent will benefit communities in the San Rafael municipality and one percent will benefit certain outlying municipalities in Santa Rosa and Jalapa departments. The remaining two percent is paid to the Guatemalan government. The Company also established a profit sharing program that provides a 0.5% NSR payment to an association of the former land owners of the Escobal mine property. Ten percent of this money is deposited in a special fund administrated by the association’s board of directors and used for improvements in the local communities.

20.6RECLAMATION

20.6.1Concurrent Reclamation

MSR conducts concurrent reclamation whenever possible by grading, contouring, and vegetating disturbed areas that are no longer necessary for use in the operation.

Topsoil removed during construction was stockpiled for use in the final reclamation of the site upon cessation of mining and processing activities. The topsoil stockpiles are contoured, seeded, and covered with soil-stabilizing biodegradable fabric to minimize erosion. Areas disturbed by the construction of the plant site have been graded, recontoured and revegetated to minimize erosion.

As the tailings dry stack facility increases in size, the front slope of the facility is continually reclaimed by placing topsoil covered by soil-stabilizing biodegradable mesh. Revegetation is accomplished by a combination of seeding and the planting of seedlings that are of the same variety found in the area.

The Company has also reclaimed and revegetated areas disturbed prior to the Company’s activities at the Escobal mine and maintains its own greenhouses to ensure the quantity, quality, and availability of domestic plants for revegetation.

20.6.2Final Reclamation and Closure

MARN Resolution 3061-2011 approving the Escobal EIA includes conditions and requirements for mine reclamation, closure and monitoring. Principal final reclamation and closure items include:

·Permanent closure of all access points to the underground workings;

·Removal of surface facilities or conversion to approved beneficial post-mining use;

·Distribution and grading of stockpiled topsoil; and

·Surface recontouring and revegetation.

Site conditions, including surface and groundwater quality, will be monitored post-closure to ensure successful permanent reclamation.

The Escobal mine was designed and engineered with closure in mind. To the greatest extent possible, the facilities were designed and are operated to minimize surface footprints and areas of disturbance and utilize advanced planning and reclamation techniques such as dry stack tailings, concurrent reclamation where possible, and landform grading. M3 (2011) authored a reclamation and closure plan for the Escobal mine which detailed reclamation and closure

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objectives and processes, and post-mining land use, surface water management, revegetation, and environmental monitoring.

Detailed cost estimates for M3’s reclamation and closure plan (2011) were calculated by Czarnowsky Inc. of Lakeside, Montana utilizing local labor and equipment rates.

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21CAPITAL AND OPERATING COSTS

21.1CAPITAL COSTS

21.1.1Expansion Capital

Capital costs to increase mine production and mill throughput to nominal 4,500 t/d rate total $23.4M to be expended through 2016. Expansion capital requirements for mine operations are for underground equipment fleet additions and paste backfill plant expansion. Expansion capital for processing operations is for the addition of a fourth tailings filter, upgrades to process plant infrastructure and ancillary equipment, and purchase of an additional haul truck for dry stack tailings operations. Table 21-1 summarizes the estimated expansion capital for the Escobal mine.

Table 21-1: Expansion Capital Costs

Area	Expansion Capital Item	Total Cost ($000s)
Mine	AC 282 Jumbo (2)	2,580
	AC Boltec (1)	875
	Simba Longhole Drill (1)	790
	Cubex Drill (1)	1,400
	CAT R2900 LHD (1)	1,400
	CAT R1700 LHD (1)	950
	CAT AD45 Truck (2)	1,960
	Paste Plant Expansion	2,824
	Total Mine Expansion Capital	12,779
Processing	Primary Crushing	650
	Secondary/Tertiary Crushing	137
	Grinding	473
	Flotation	2,109
	Concentrate Thickener	179
	Tailings Filtration	7,512
	Tailings Dry Stack Equipment	475
	Total Processing Expansion Capital	11,535
Total Expansion Capital		24,314

Underground equipment costs are based on recent quotes and/or purchases made by the Company. Actual costs for these pieces may vary depending on the vendor and available options but the differences are not material to this analysis.

21.2SUSTAINING CAPITAL

The sustaining capital costs for the Escobal mine are estimated to be $301.2M over the life of the mine. Table 21-2 shows a summary of the estimated sustaining capital costs. Sustaining capital requirements for the mine are generally limited to capitalized mine development and associated infrastructure (ventilation, electrical distribution, primary pumping stations, paste distribution lines, etc.), underground equipment purchases and equipment rebuilds and replacements. The majority (87%) of process plant sustaining capital is for bi-annual ball mill liner replacement. Surface operations sustaining capital is primarily for dry stack buttress construction and associated equipment.

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Table 21-2: Life of Mine Sustaining Capital Costs

Area	Capital Item	Total Cost ($000s)
Mine	Equipment Purchases	14,058
	Equipment Rebuilds	41,619
	Equipment Replacements	42,350
	Capitalized Equipment Operation	30,713
	Capitalized Mine Development	76,946
	Capitalized Mine Infrastructure	36,900
	Other	5,200
	Total Mine Sustaining Capital	247,786
Processing	Assay Lab	875
	Equipment Purchases	1,060
	Equipment Rebuilds	1,836
	Auxiliary Process Water Tank	160
	Zinc Bulk Loading System	300
	Ball Mill Liner Replacements	27,300
	Total Processing Sustaining Capital	31,531
Surface Operations*	Dry Stack Buttress Construction	9,361
	Dry Stack Equipment Purchases	840
	Dry Stack Equipment Rebuilds	3,152
	Dry Stack Equipment Replacement	1,159
	Water Truck	85
	Site Power Synchronization Equipment	60
	Other	295
	Total Surface Operations Sustaining Capital	14,952
Site G&A Capital	Infrastructure	3,215
	Warehouse Equipment Purchases	282
	Vehicle Purchases	881
	Other	2,532
	Total Site G&A Sustaining Capital	6,910
Total Life of Mine Sustaining Capital		301,179

*Sustaining capital for surface operations are incorporated into processing sustaining capital in the life of mine financial model.

The average sustaining cost per year is approximately US$15M over the life of the mine; however this fluctuates from year to year based on equipment rebuild and replacement schedules and the quantities of capitalized development necessary to sustain mine production.

Figure 21-1 illustrates the annual sustaining and expansion capital expenditures for the Escobal mine over the life of the mine.

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Figure 21-1: Life of Mine Capital Expenditures

21.3DEVELOPMENT COSTS

The development costs in the financial analysis are based upon actual costs encountered at the Escobal mine used in conjunction with the proposed development design. Development headings were grouped by type in the development schedule with costs for materials and consumables estimated using the rates in Table 21-3. In general, the development costs are all capitalized with the exception of the stope accesses where are expensed as an operating cost.

Table 21-3: Development Costs (Materials & Consumables Only)

Development Type	$/meter
Primary Ramp	791.96
Sublevel Access	751.74
Diamond Drill Station	751.74
Connection Drift	751.74
Muckbay	525.56
Service Cutout	525.56
Sump	645.84
Dewatering Cutout	751.74
Vent Raise Access	751.74
Ore Pass Access	751.74
Alimak Raises	12,000
Footwall Lateral	819.66
Stope Access (Operating Cost)	696.84

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21.4OPERATING COSTS

Table 21-4 summarizes the average life of mine operating costs estimated for the Escobal mine. The average mine operating costs is $75.13 per tonne of ore. Operating costs for the Escobal mine have been derived for the Company’s actual operating experience at Escobal and engineering first-principles.

Table 21-4: Escobal Mine Operating Costs

Area	$/tonne
Mine	37.23
Processing	20.36
Surface Operations*	2.47
General & Administrative	15.06
Total Mine Operating Costs	75.13

*Operating costs for surface operations are incorporated into processing operating costs in the life of mine financial model.

The annual operating cost per tonne is estimated to be $91.19 for the second half of 2014 and averaging $74.28 from 2016 through the end of the mine life, upon full implementation of the expansion to 4,500 t/d. The estimated average annual operating costs over the life of mine are shown in Figure 21-2.

Figure 21-2: Life of Mine Annual Operating Costs

21.4.1Mining Costs

Operating costs for mining, including capital development costs, were developed on a unit cost and quantity basis utilizing both first principals and similar operation comparisons. Actual operating costs, metallurgical performance, and smelter contracts from the Escobal mine and engineering first-principles were used to derive operating costs and revenue.

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The average mining operating cost over the life of the Escobal mine is approximately US$37.23 per tonne of ore delivered to the process plant. Table 21-5 summarizes the operating cost by work area. Approximately 87% of the operating costs are generated from two work areas, Mine Operations and Backfill Services. The remaining cost areas are primarily labor or G&A services in support of the underground operations.

Table 21-5: Mine Operating Costs by Work Area

Work Area	$/tonne
Mine Operations	23.29
Mine Services	0.43
Training	0.08
Mine Rescue	0.02
Backfill Services	9.25
Shotcrete	0.36
Mine & Surface Equipment Shop	1.45
Mine Electrical Services	0.36
Mine Engineering	0.38
Geology	0.65
Underground Drilling	0.96
Total Mine Operating Costs	37.23

The operating costs vary by year, depending on the production requirements and the allocation between expensed and capitalized development costs, as previously illustrated in Figure 21-1. The general trend is slightly downwards over time with some fluctuations primarily due to expensed development costs. Unit costs include the cost of labor and supplies, including ground support materials, explosives, installed piping, electrical, and communications lines and ventilation and pumping systems integral to the unit operation. Each cost item includes power allocation costs and labor costs derived from actual costs through June 30, 2014. Labor estimates are based the current employee roster and additions to labor necessary for the expanded production.

Mining costs are based upon actual costs incurred at the Escobal mine used in conjunction with the proposed stope designs and production schedule. Ore production was grouped by stoping method in the production schedule (as discussed in Section 16) with costs for materials and consumables estimated using the rates in Table 21-6.

Table 21-6: Production Costs (Materials & Consumables Only)

Stoping Method	$/tonne
Transverse Stopes	4.51
Longitudinal Stoping	11.48

Approximately 63% of the total mining operating costs come from the direct mining costs, as summarized in Table 21-7.

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Table 21-7: Mine Operations Costs

Cost Type	$/tonne
Labor	1.81
Materials & Consumables	6.61
Expensed Development	2.63
Electricity	5.22
Drilling Equipment	1.90
Loading Equipment	1.77
Haulage Equipment	1.90
Auxiliary Equipment	0.79
Light-duty Vehicles	0.04
G&A	0.62
Total Mine Operations Costs	23.29

As with the overall total mine operating costs, the costs in Table 21-7 vary over time to allow for portions of the labor, materials and equipment usage to be capitalized while developing new areas of the mine.

21.4.2Backfill Costs

The backfill costs include the operation and maintenance of the paste plant, pumping system, piping system, as well as cement as a binder. This cost includes personnel for the paste backfill plant and allocated power costs. The cement cost is estimated at $205/tonne.

Table 21-8: Backfill Operating Costs

Cost Type	$/tonne
Labor	0.47
Plant Operation	8.67
Light-duty Vehicles	0.01
G&A	0.10
Total Backfill Costs	9.25

21.4.3Processing Costs

Process plant operating costs are summarized by area in Table 21-9. Each cost item includes allocated power costs and labor costs derived from actual costs through June 30, 2014.

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Table 21-9: Process Operating Costs by Area

Area	$/tonne
Primary Crushing	0.45
Secondary & Tertiary Crushing	0.72
Grinding & Classification	2.98
Flotation & Regrind	1.46
Concentrate Filtration	0.31
Concentrator Operations	5.81
Concentrator Maintenance	3.42
Tailings Filtration	1.65
Metallurgical Laboratory	0.09
Assay Laboratory	1.04
Services & Water Treatment	0.42
Process Operations G&A	1.89
Ancillary	0.12
Total Process Operating Costs	20.36

21.4.4Surface Operation Costs

Surface operations includes tailings dry stack operations, operation and maintenance of dewatering and domestic wells, road maintenance, and other general surface operations. Each cost item includes power allocation costs and labor costs derived from actual costs through June 30, 2014.

Table 21-10: Surface Operations Operating Costs

Area	$/tonne
General Surface Operations	0.55
Tailing Disposal (Dry Stack)	1.70
Dewatering Well Operation & Maintenance	0.21
Total Surface Operations Costs*	2.47

Surface operations costs are incorporated into processing costs in the life of mine financial model.

21.4.5General and Administrative

The general and administrative (G&A) costs to support the Escobal mine operations are summarized in Table 21-11. General and administrative costs include employee salaries and benefits. A summary of the labor distribution in July 2014 and December 2016, after mine expansion is complete, is provided in Table 21-12)

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Table 21-11: General and Administrative Costs by Area

Area	$/tonne
General Administration	5.12
General Site Maintenance	0.22
Corporate Social Responsibility	0.89
Communications	0.18
Environmental	0.48
Medical Services	0.41
Warehouse	0.57
Purchasing	0.32
Safety	0.51
Security	2.05
Human Resources	1.56
Mine Camp	1.35
Offsite Housing	0.32
Information & Technology	0.84
Taxes & Finance	0.01
Accounting	0.24
Total G&ACosts	15.06

Table 21-12: Labor Distribution

Area	Jul 2014	Dec 2016
Mine Operations	443	496
Process Operations	168	194
Surface Operations	49	64
General & Administration	169	203
Total	829	957

21.5ELECTRICAL POWER

The current rate for the contractor-supplied diesel power generation averages $0.32 per kilowatt hour (kW-hr). Power costs are allocated to the main operations areas — mine (35%), process plant (62%), surface operations and G&A (3%) based on connected kW of each. The Company is actively investigating alternative lower cost power sources such as heavy fuel oil power generation and, based on contractor proposals received by the Company, M3 believes a 40% reduction in power costs is obtainable in the near term. This Study assumes a transition away from diesel generated power in mid-2015 and achieving the full estimated reduction in power costs by 2016.

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22ECONOMIC ANALYSIS

22.1INTRODUCTION

The financial evaluation presents the determination of the Net Present Value (NPV) and sensitivities for the project. Annual cash flow projections were estimated over the life of the mine based on the estimates of capital expenditures and production cost and sales revenue. The sales revenue is based on the production of lead and zinc concentrate also containing gold and silver. The estimates of capital expenditures and site production costs have been developed specifically for this project and have been presented in earlier sections of this report.

22.2MINE PRODUCTION STATISTICS

Mine production is reported as ore and waste from the mining operation. The annual production figures were obtained from the mine plan as reported earlier in this report.

The life of mine ore and waste quantities and ore grade are presented in Table 22-1.

Table 22-1: Life of Mine Ore, Waste and Metal Grades

	Tonnes
	(000’s)	Silver g/t	Gold g/t	Lead %		Zinc %
Ore Tonnes	31,433	346.8	0.3	0.7	%	1.2	%
Waste Tonnes	9,546

22.3PLANT PRODUCTION STATISTICS

Ore will be processed using crushing, grinding, and flotation technology to produce metals in a flotation concentrate. Two concentrate products will be produced, zinc concentrate and lead concentrate. Gold and silver will be recovered in both the zinc and lead concentrates.

The estimated metal recoveries in the lead and zinc concentrates are presented in Table 22-2.

Table 22-2: Metal Recovery Factors

	Silver %	Gold %	Lead %	Zinc %
Lead Concentrate	81.3	58.3	89.8	—
Zinc Concentrate	3.4	2.5	—	75.7

Estimated life of mine lead and zinc concentrate production is presented in Table 22-3 with the approximate metal contained.

Table 22-3: Life of Mine Concentrate Summary

	Tonnes (000’s)	Silver (kozs)	Gold (kozs)	Lead (klbs)	Zinc (klbs)
Lead Concentrate	495	285,265	196	459,459	—
Zinc Concentrate	551	11,739	9	—	636,449

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22.3.1Smelter Return Factors

Lead and zinc concentrates are shipped from the Escobal mine site to Puerto Quetzal, where they are loaded onto vessels for shipment to lead and zinc smelting and refining companies. Smelter and refining treatment charges are agreed to by long-term frame agreements between the Company and multiple lead and zinc smelter customers, the terms of which are renewed annually to reflect market conditions for high precious metal lead and zinc concentrates.

A smelter may impose a penalty either expressed in higher treatment charges or in metal deductions to treat concentrates that contain higher than specified quantities of certain elements. The Escobal concentrates are easily marketable as they are relatively clean concentrates that do not pose special restrictions on smelting and refining.

The smelting and refining charges calculated in the financial evaluation include charges for smelting lead and zinc concentrates and refining precious metal from both the lead and zinc concentrates. Transportation costs to deliver the concentrates from the mine site to the smelters are also included. The off-site charges that are incurred are presented in Table 22-4.

Table 22-4: Smelter Return Factors

Lead Concentrate
Payable lead in concentrate	94.0		%
Payable gold in concentrate	92.2		%
Payable silver in concentrate	95.9		%
Treatment charge ($/tonne)	$	270.00
Refining charge — Au ($/oz)	$	6.65
Refining charge — Ag ($/oz)	$	1.13
Penalty ($/tonne conc.)	$	7.27
Transportation Charges ($/wmt)	$	139.25
Moisture (%)	8.0		%
Zinc Concentrate
Payable zinc in concentrate	84.0		%
Payable gold in concentrate	56.0		%
Payable silver in concentrate	66.4		%
Treatment charge ($/tonne)	$	274.90
Refining charge — Au ($/oz)	$	0.00
Refining charge — Ag ($/oz)	$	0.00
Penalty ($/tonne conc.)	$	0.68
Transportation Charges ($/wmt)	$	277.10
Moisture (%)	8.0		%

22.4CAPITAL EXPENDITURE

22.4.1Expansion and Sustaining Capital

The financial indicators have been determined with cash flow from the Escobal mine financing of the sustaining and expansion capital. Any acquisition cost or expenditures prior to the July 2014 have been treated as “sunk” cost and are not included in the analysis.

The total capital carried in the financial model for expansion capital and sustaining capital is shown in Table 22-5.

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Table 22-5: Expansion and Sustaining Capital Summary

Period	Amount ($000)
Year 2014	$	19,991
Year 2015	$	44,977
Year 2016	$	22,376
Year 2017	$	19,117
Year 2018	$	20,194
Year 2019	$	23,258
Year 2020	$	12,360
Year 2021	$	12,151
Year 2022	$	17,054
Year 2023	$	18,301
Year 2024	$	25,480
Year 2025	$	15,476
Year 2026	$	15,640
Year 2027	$	12,851
Year 2028	$	11,334
Year 2029	$	13,828
Year 2030	$	9,419
Year 2031	$	5,205
Year 2032	$	3,666
Year 2033	$	2,814
Total	$	325,493

Expansion capital accounts for approximately 30% of total capital expenditures in the second half of 2014 and 2015.

22.4.2Working Capital

A 20-day delay of receipt of revenue from sales is used for accounts receivables and takes into account the provisional payment and final settlement arrangements provided in the existing smelter contracts. A delay of payment for accounts payable of 30 days is also incorporated into the financial model. All the working capital is recaptured at the end of the mine life and the final value of these accounts is $0.

22.4.3Salvage Value

No allowance for salvage value has been included in the cash flow analysis.

22.5REVENUE

Annual revenue is determined by applying estimated metal prices to the annual payable metal estimated for each operating year. Sales prices have been applied to all life of mine production without escalation or hedging. The revenue is the gross value of payable metals sold before treatment and transportation charges. Metal sales prices used in the evaluation are as follows:

Silver$18.00/troy ounce

Gold$1,300.00/troy ounce

Lead$0.95/pound

Zinc$0.90/pound

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22.6OPERATING COST

Life of mine Cash Operating Costs include mine operations, process plant operations, general administrative cost, smelting and refining charges and shipping charges. Table 22-6 shows the estimated operating cost by area per metric ton of ore processed.

Table 22-6: Operating Cost

Operating Cost	$/ore tonne
Mine	$	37.23
Process Plant	$	22.83
General Administration	$	15.06
Smelting/Refining Treatment	$	26.64
Total Operating Cost	$	101.77

22.6.1Total Cash Cost

The average Total Cash Cost over the life of the mine is estimated to be $111.74/t of ore processed. Total Cash Cost is the Total Cash Operating Cost plus royalties, property tax and tailing infrastructure and reclamation and closure costs.

22.6.1.1Royalty

Royalty payments are based on 5.5% of the NSR at market prices; the life of mine royalty payments are estimated to be $308.0 million.

22.6.1.2Depreciation

Depreciation is calculated using the straight line method starting with first year of production. The expansion capital was depreciated using a 10 year life and the sustaining capital was depreciated using an 8 year life. The last year of production is the catch-up year if the assets are not fully depreciated by that time.

22.6.2Reclamation & Closure

A non-cash allowance for the cost of final reclamation and closure of the property has been included in the net income projection at a rate of $0.19 per ore tonne mined, which is derived by dividing the estimated closure costs by the total tonnes processed over the life of mine.

22.7TAXATION

The Escobal project is evaluated with a 7% corporate income tax based on NSR. The tax is paid to the Superintendencia de Administración Tributaria (Superintendency of Tax Administration, SAT), the Guatemalan federal tax authority. All metal concentrates produced from the Escobal mine are shipped out of the country.

Corporate income taxes paid is estimated to be $379.6 million for the life of the mine.

22.8PROJECT FINANCING

For the purposes of this study it is assumed investment in the Escobal mine will be financed with cash flows generated by the mine. Therefore, no interest payments on debt are considered.

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22.9NET INCOME AFTER TAX

Net Income after Tax is approximately $1.2 billion for the life of the mine.

22.10NPV

The base case economic analysis indicates that the project has an NPV at 5% discount rate of $1.52 billion. Sensitivity analyses are presented in Table 22-7.

22.11IRR

The financial analysis does not present an internal rate of return (IRR) as it is not a meaningful metric for the Escobal mine. The calculated return on investment is magnified by the significant mine cash flows offset by the exclusion of the initial capital investment in the calculation and would exaggerate any IRR calculated. Net Present Value as presented above provides a more accurate and meaningful economic assessment of the mine.

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Table 22-7: Sensitivity Analysis After Taxes (in Thousands of $)

Change in Metal Prices			NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case			$	2,012,550	$	1,515,689	$	1,214,041
	20	%	$	3,100,197	$	2,288,349	$	1,805,971
	10	%	$	2,556,373	$	1,902,019	$	1,510,006
	0	%	$	2,012,550	$	1,515,689	$	1,214,041
	-10	%	$	1,526,811	$	1,160,056	$	935,374
	-20	%	$	1,029,418	$	800,057	$	654,775

Change in Silver Prices			NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case			$	2,012,550	$	1,515,689	$	1,214,041
	$	25.00	$	3,740,225	$	2,761,565	$	2,179,556
	$	22.00	$	2,999,793	$	2,227,618	$	1,765,764
	$	20.00	$	2,506,171	$	1,871,653	$	1,489,902
	$	18.00	$	2,012,550	$	1,515,689	$	1,214,041
	$	15.00	$	1,329,003	$	1,012,566	$	818,096

Change in Operating Cost			NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case			$	2,012,550	$	1,515,689	$	1,214,041
	20	%	$	1,605,357	$	1,228,547	$	995,688
	10	%	$	1,776,080	$	1,354,731	$	1,095,059
	0	%	$	2,012,550	$	1,515,689	$	1,214,041
	-10	%	$	2,249,020	$	1,676,647	$	1,333,023
	-20	%	$	2,485,489	$	1,837,605	$	1,452,005

Change in Initial Capital			NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case			$	2,012,550	$	1,515,689	$	1,214,041
	20	%	$	2,007,687	$	1,510,869	$	1,209,261
	10	%	$	2,010,119	$	1,513,279	$	1,211,651
	0	%	$	2,012,550	$	1,515,689	$	1,214,041
	-10	%	$	2,014,981	$	1,518,099	$	1,216,431
	-20	%	$	2,017,413	$	1,520,508	$	1,218,821

Change in Recovery			NPV @ 0%		NPV @ 5%		NPV @ 10%
Base Case			$	2,012,550	$	1,515,689	$	1,214,041
	2.0	%	$	2,106,005	$	1,582,562	$	1,265,556
	1.0	%	$	2,059,277	$	1,549,125	$	1,239,799
	0.0	%	$	2,012,550	$	1,515,689	$	1,214,041
	-1.0	%	$	1,965,823	$	1,482,253	$	1,188,284
	-2.0	%	$	1,919,095	$	1,448,816	$	1,162,526

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Table 22-8: Detail Financial Model

			H2 2014		2015		2016		2017		2018		2019		2020		2021		2022		2023
Base Case	Total		1		2		3		4		5		6		7		8		9		10
Mining Operations
Ore
Beginning Inventory (kt)	31,433		31,433		30,775		29,246		27,621		25,963		24,321		22,688		21,056		19,424		17,806
Mined (kt)	31,433		658		1,529		1,624		1,658		1,642		1,633		1,633		1,631		1,619		1,643
Ending Inventory (kt)	—		30,775		29,246		27,621		25,963		24,321		22,688		21,056		19,424		17,806		16,163
Gold Grade (g/t)	0.33		0.38		0.36		0.31		0.32		0.32		0.40		0.61		0.27		0.34		0.27
Silver Grade (g/t)	347		542		482		441		442		442		442		442		442		398		266
Lead Grade (%)	0.74	%	0.74	%	0.68	%	0.65	%	0.67	%	0.67	%	0.66	%	0.82	%	0.46	%	0.61	%	0.56	%
Zinc Grade (%)	1.21	%	1.27	%	1.19	%	1.13	%	1.12	%	1.13	%	1.11	%	1.39	%	0.78	%	1.01	%	1.03	%
Contained Gold (kozs)	336		8		18		16		17		17		21		32		14		18		15
Contained Silver (kozs)	350,484		11,466		23,707		23,007		23,555		23,337		23,200		23,202		23,184		20,706		14,056
Contained Lead (klbs)	510,876		10,744		23,033		23,377		24,668		24,399		23,816		29,553		16,436		21,894		20,431
Contained Zinc (klbs)	839,255		18,442		40,153		40,622		40,803		40,780		40,107		50,063		27,991		36,181		37,138
Waste
Beginning Inventory(kt)	9,546		9,546		9,336		8,817		8,060		7,336		6,531		5,806		5,136		4,604		3,972
Mined (kt)	9,546		210		519		757		724		805		725		670		532		632		620
Ending Inventory (kt)	—		9,336		8,817		8,060		7,336		6,531		5,806		5,136		4,604		3,972		3,352
Total Material Mined (kt)	40,979		868		2,048		2,381		2,382		2,447		2,358		2,303		2,164		2,251		2,262
Waste to Ore Ratio	0.30		0.32		0.34		0.47		0.44		0.49		0.44		0.41		0.33		0.39		0.38
Process Plant Operations
Concentrator
Beginning Ore Inventory (kt)	—		—		—		—		—		—		—		—		—		—		—
Mined Ore to Concentrator (kt)	31,476		684		1,489		1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,643
Mined Ore - Processed (kt)	31,476		684		1,489		1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,643
Ending Ore Inventory			—		—		—		—		—		—		—		—		—		—
Gold Grade (g/t)	0.33		0.37		0.36		0.31		0.32		0.32		0.40		0.60		0.27		0.34		0.27
Silver Grade (g/t)	346.75		527.22		482.71		441.99		441.93		441.95		441.97		441.98		442.05		398.50		266.18
Lead Grade (%)	0.74	%	0.76	%	0.68	%	0.65	%	0.67	%	0.67	%	0.66	%	0.82	%	0.46	%	0.61	%	0.56	%
Zinc Grade (%)	1.21	%	1.27	%	1.19	%	1.14	%	1.12	%	1.13	%	1.11	%	1.38	%	0.79	%	1.01	%	1.03	%
Contained Gold (kozs)	336		8		17		17		17		17		21		32		14		18		15
Contained Silver (kozs)	350,902		11,593		23,104		23,340		23,337		23,338		23,340		23,340		23,343		21,044		14,056
Contained Lead (klbs)	511,814		11,399		22,450		23,677		24,422		24,400		23,973		29,575		16,823		22,137		20,430
Contained Zinc (klbs)	840,398		19,111		39,119		41,148		40,442		40,770		40,362		50,094		28,644		36,596		37,138
Zinc Concentrate
Recovery Zinc (%)	75.73	%	75.95	%	75.54	%	75.22	%	75.11	%	75.16	%	75.10	%	76.52	%	72.93	%	74.45	%	74.55	%
Recovery Gold (%)	2.54	%	2.65	%	2.68	%	2.76	%	2.76	%	2.76	%	2.57	%	1.25	%	2.78	%	2.73	%	2.78	%
Recovery Silver (%)	3.35	%	2.30	%	2.61	%	2.89	%	2.89	%	2.89	%	2.89	%	2.89	%	2.89	%	3.18	%	4.02	%
Zinc Concentrate (kt)	551		12		26		27		26.647		27		27		33		19		24		25
Zinc Concentrate - Grade (Zn %)	52.41	%	52.76	%	52.26	%	51.85	%	51.71	%	51.78	%	51.69	%	53.44	%	48.75	%	50.84	%	50.97	%
Recovered Zinc (klbs)	636,449		14,514		29,550		30,953		30,376		30,644		30,310		38,330		20,889		27,247		27,686
Recovered Gold (kozs)	9		0		0		0		0		0		1		0		0		0		0
Recovered Silver (kozs)	11,739		266		603		674		674		674		674		674		674		669		565
Lead Concentrate
Recovery Lead (%)	89.77	%	89.86	%	88.99	%	88.60	%	88.87	%	88.86	%	88.71	%	90.52	%	85.78	%	88.02	%	87.35	%
Recovery Gold (%)	58.34	%	59.66	%	59.30	%	57.31	%	57.48	%	57.42	%	60.49	%	63.59	%	55.34	%	58.22	%	55.47	%
Recovery Silver (%)	81.29	%	85.15	%	84.28	%	83.37	%	83.37	%	83.37	%	83.37	%	83.37	%	83.37	%	82.29	%	78.31	%
Lead Concentrate (kt)	495		11		22		23		24		24		24		28		18		22		21
Lead Concentrate - Grade (Pb%)	42.09	%	42.64	%	41.28	%	40.62	%	41.07	%	41.06	%	40.81	%	43.57	%	35.44	%	39.62	%	38.40	%
Recovered Lead (klbs)	459,459		10,243		19,979		20,978		21,704		21,682		21,266		26,770		14,431		19,486		17,845
Recovered Gold (kozs)	196		5		10		10		10		10		13		20		8		10		8

	2024		2025		2026		2027		2028		2029		2030		2031		2032		2033		2034
Base Case	11		12		13		14		15		16		17		18		19		20		21
Mining Operations
Ore
Beginning Inventory (kt)	16,163		14,513		12,878		11,231		9,592		7,951		6,296		4,659		3,021		1,362		(0	)
Mined (kt)	1,650		1,635		1,647		1,639		1,642		1,654		1,638		1,637		1,660		1,362		—
Ending Inventory (kt)	14,513		12,878		11,231		9,592		7,951		6,296		4,659		3,021		1,362		(0	)	(0	)
Gold Grade (g/t)	0.34		0.33		0.41		0.44		0.37		0.32		0.28		0.17		0.17		0.25		—
Silver Grade (g/t)	278		283		332		312		272		266		243		251		246		224		—
Lead Grade (%)	0.63	%	0.63	%	0.75	%	0.81	%	0.86	%	0.75	%	0.87	%	0.80	%	1.07	%	1.08	%	0.00	%
Zinc Grade (%)	1.04	%	1.11	%	1.31	%	1.50	%	1.47	%	1.27	%	1.41	%	1.02	%	1.27	%	1.80	%	0.00	%
Contained Gold (kozs)	18		17		22		23		20		17		15		9		9		11		—
Contained Silver (kozs)	14,719		14,904		17,596		16,416		14,378		14,130		12,817		13,188		13,118		9,797		—
Contained Lead (klbs)	22,995		22,588		27,259		29,372		31,056		27,429		31,291		28,850		39,233		32,453		—
Contained Zinc (klbs)	37,728		39,922		47,583		54,239		53,368		46,209		50,823		36,675		46,472		53,956		—
Waste
Beginning Inventory(kt)	3,352		2,797		2,158		1,634		1,159		832		529		306		152		75		0
Mined (kt)	555		639		524		475		327		302		223		153		78		75		—
Ending Inventory (kt)	2,797		2,158		1,634		1,159		832		529		306		152		75		0		0
Total Material Mined (kt)	2,205		2,275		2,171		2,114		1,969		1,956		1,861		1,791		1,738		1,436		—
Waste to Ore Ratio	0.34		0.39		0.32		0.29		0.20		0.18		0.14		0.09		0.05		0.05		—
Process Plant Operations
Concentrator
Beginning Ore Inventory (kt)	—		—		—		—		—		—		—		—		—		—		—
Mined Ore to Concentrator (kt)	1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,381		—
Mined Ore - Processed (kt)	1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,643		1,381		—
Ending Ore Inventory	—		—		—		—		—		—		—		—		—		—		—
Gold Grade (g/t)	0.34		0.33		0.41		0.44		0.37		0.32		0.28		0.17		0.17		0.25		—
Silver Grade (g/t)	277.55		283.43		332.25		311.67		272.44		265.69		243.61		250.51		245.81		224.11		—
Lead Grade (%)	0.63	%	0.63	%	0.75	%	0.81	%	0.86	%	0.75	%	0.87	%	0.80	%	1.07	%	1.08	%	0.00	%
Zinc Grade (%)	1.04	%	1.11	%	1.31	%	1.50	%	1.47	%	1.27	%	1.41	%	1.02	%	1.27	%	1.79	%	0.00	%
Contained Gold (kozs)	18		17		22		23		20		17		15		9		9		11		—
Contained Silver (kozs)	14,657		14,967		17,545		16,458		14,387		14,030		12,865		13,229		12,981		9,947		—
Contained Lead (klbs)	22,896		22,686		27,180		29,436		31,073		27,234		31,356		28,953		38,812		32,902		—
Contained Zinc (klbs)	37,567		40,082		47,444		54,349		53,398		45,881		50,940		36,852		45,975		54,487		—
Zinc Concentrate
Recovery Zinc (%)	74.62	%	75.05	%	76.16	%	77.03	%	76.92	%	75.94	%	76.62	%	74.50	%	75.96	%	78.02	%	0.00	%
Recovery Gold (%)	2.73	%	2.74	%	2.52	%	2.40	%	2.66	%	2.75	%	2.78	%	2.60	%	2.60	%	2.77	%	0.00	%
Recovery Silver (%)	3.95	%	3.91	%	3.61	%	3.74	%	3.98	%	4.02	%	4.15	%	4.11	%	4.14	%	4.27	%	0.00	%
Zinc Concentrate (kt)	25		26		31		35		35		30		33		24		30		35		—
Zinc Concentrate - Grade (Zn %)	51.07	%	51.63	%	53.02	%	54.00	%	53.89	%	52.76	%	53.56	%	50.90	%	52.77	%	54.88	%	0.00	%
Recovered Zinc (klbs)	28,034		30,081		36,135		41,865		41,074		34,844		39,033		27,454		34,922		42,509		—
Recovered Gold (kozs)	0		0		1		1		1		0		0		0		0		0		—
Recovered Silver (kozs)	579		586		633		615		573		564		534		544		537		425		—
Lead Concentrate
Recovery Lead (%)	88.31	%	88.23	%	89.80	%	90.48	%	90.93	%	89.81	%	91.00	%	90.34	%	92.56	%	92.61	%	0.00	%
Recovery Gold (%)	58.19	%	58.01	%	60.91	%	61.61	%	59.54	%	57.65	%	55.95	%	49.13	%	49.13	%	54.22	%	0.00	%
Recovery Silver (%)	78.70	%	78.89	%	80.43	%	79.80	%	78.53	%	78.30	%	77.53	%	77.77	%	77.61	%	76.83	%	0.00	%
Lead Concentrate (kt)	23		23		26		28		29		26		29		27		36		30		—
Lead Concentrate - Grade (Pb%)	40.13	%	39.99	%	42.54	%	43.51	%	44.09	%	42.57	%	44.18	%	43.32	%	45.35	%	45.35	%	0.00	%
Recovered Lead (klbs)	20,221		20,017		24,406		26,632		28,253		24,460		28,534		26,155		35,924		30,471		—
Recovered Gold (kozs)	10		10		13		14		12		10		8		4		4		6		—

209

ESCOBAL MINE GUATEMALA

FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

			H2 2014		2015		2016		2017		2018		2019		2020		2021		2022		2023
Base Case	Total		1		2		3		4		5		6		7		8		9		10
Recovered Silver (kozs)	285,265		9,872		19,471		19,459		19,456		19,457		19,458		19,458		19,461		17,317		11,008
Payable Metals
Zinc Concentrate
Payable Zinc (klbs)	534,617		12,192		24,822		26,000		25,515		25,741		25,460		32,198		17,547		22,887		23,256
Payable Gold (kozs)	5		0		0		0		0		0		0		0		0		0		0
Payable Silver (kozs)	7,791		177		400		448		448		448		448		448		448		444		375
Lead Concentrate
Payable Lead (klbs)	431,891		9,629		18,780		19,719		20,401		20,381		19,990		25,164		13,565		18,317		16,774
Payable Gold (kozs)	181		5		10		9		9		9		12		19		7		10		7
Payable Silver (kozs)	273,669		9,470		18,679		18,668		18,665		18,666		18,667		18,667		18,670		16,613		10,560
Income Statement ($000)
Zinc ($/lb.)	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90
Lead ($/lb)	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95
Gold ($/oz)	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00
Silver ($/oz)	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00
Revenues
Zinc Concentrate - Zn	$	481,155	$	10,972	$	22,340	$	23,400	$	22,964	$	23,167	$	22,914	$	28,978	$	15,792	$	20,598	$	20,930
Zinc Concentrate - Au	$	6,214	$	159	$	339	$	334	$	337	$	336	$	395	$	290	$	291	$	353	$	294
Zinc Concentrate - Ag	$	140,243	$	3,183	$	7,205	$	8,057	$	8,057	$	8,057	$	8,057	$	8,057	$	8,057	$	7,994	$	6,746
Lead Concentrate - Pb	$	410,296	$	9,147	$	17,841	$	18,733	$	19,381	$	19,362	$	18,991	$	23,906	$	12,887	$	17,401	$	15,936
Lead Concentrate - Au	$	234,866	$	5,886	$	12,386	$	11,398	$	11,578	$	11,515	$	15,298	$	24,304	$	9,540	$	12,378	$	9,653
Lead Concentrate - Ag	$	4,926,041	$	170,468	$	336,230	$	336,016	$	335,967	$	335,983	$	336,003	$	336,009	$	336,066	$	299,041	$	190,087
Total Revenues	$	6,198,816	$	199,816	$	396,342	$	397,938	$	398,284	$	398,420	$	401,658	$	421,543	$	382,633	$	357,767	$	243,647
Operating Cost
Mining	$	1,171,965	$	26,949	$	59,357	$	62,845	$	64,646	$	65,782	$	66,538	$	66,593	$	59,487	$	57,523	$	58,789
Process Plant	$	718,600	$	20,998	$	40,471	$	38,350	$	38,108	$	38,108	$	38,106	$	37,685	$	37,307	$	36,966	$	36,660
General Administration	$	474,131	$	14,425	$	26,042	$	24,885	$	24,951	$	24,946	$	24,942	$	24,687	$	24,483	$	24,319	$	24,189
Treatment & Refining Charges
Zinc Concentrates
Treatment Charges	$	151,418	$	3,430	$	7,051	$	7,443	$	7,325	$	7,380	$	7,312	$	8,944	$	5,343	$	6,682	$	6,773
Gold Refining Charges	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—
Silver Refining Charges	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—
Transportation	$	164,842	$	3,734	$	7,676	$	8,103	$	7,975	$	8,034	$	7,960	$	9,737	$	5,817	$	7,275	$	7,373
Penalties	$	372	$	8	$	17	$	18	$	18	$	18	$	18	$	22	$	13	$	16	$	17
Lead Concentrates
Treatment Charges	$	133,684	$	2,942	$	5,928	$	6,324	$	6,471	$	6,467	$	6,383	$	7,525	$	4,988	$	6,023	$	5,691
Gold Refining Charges	$	1,201	$	30	$	63	$	58	$	59	$	59	$	78	$	124	$	49	$	63	$	49
Silver Refining Charges	$	308,972	$	10,692	$	21,089	$	21,076	$	21,073	$	21,074	$	21,075	$	21,075	$	21,079	$	18,757	$	11,923
Transportation	$	74,460	$	1,639	$	3,302	$	3,522	$	3,604	$	3,602	$	3,555	$	4,191	$	2,778	$	3,355	$	3,170
Penalties	$	3,600	$	79	$	160	$	170	$	174	$	174	$	172	$	203	$	134	$	162	$	153
Total Operating Cost	$	3,203,245	84,927		171,157		172,795		174,404		175,644		176,138		180,787		161,477		161,141		154,786
Royalty	$	307,976	$	10,045	$	19,912	$	19,957	$	19,974	$	19,979	$	20,164	$	21,101	$	19,306	$	17,933	$	12,047
Property Tax	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—
Salvage Value	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—
Reclamation & Closure	$	5,980	$	130	$	283	$	312	$	312	$	312	$	312	$	312	$	312	$	312	$	312
Total Production Cost	$	3,517,202	$	95,102	$	191,352	$	193,064	$	194,690	$	195,935	$	196,614	$	202,200	$	181,096	$	179,387	$	167,145
Operating Income	$	2,681,614	$	104,714	$	204,990	$	204,874	$	203,593	$	202,485	$	205,043	$	219,343	$	201,537	$	178,380	$	76,501
Existing Asset Depreciation	$	828,900	$	18,011	$	39,205	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255
Expansion Capital Depreciation	$	24,314	$	2,431	$	2,431	$	2,431	$	2,431	$	2,431	$	2,431	$	2,431	$	2,431	$	2,431	$	2,431
Sustaining Capital Depreciation	$	294,269	$	2,390	$	4,914	$	7,821	$	9,366	$	10,885	$	13,017	$	15,304	$	18,489	$	18,034	$	17,465
Total Depreciation	$	1,147,483	$	22,832	$	46,550	$	53,507	$	55,052	$	56,571	$	58,703	$	60,990	$	64,175	$	63,720	$	63,151
Net Income After Depreciation	$	1,534,132	81,881		158,440		151,367		148,541		$	145,914	$	146,341	$	158,353	$	137,362	$	114,660	$	13,350
Tax Loss Carry Forward Applied	$	—							—		—		—		—		—		—		—

	2024		2025		2026		2027		2028		2029		2030		2031		2032		2033			2034
Base Case	11		12		13		14		15		16		17		18		19		20			21
Recovered Silver (kozs)	11,534		11,808		14,112		13,134		11,297		10,985		9,974		10,289		10,074		7,642			—
Payable Metals
Zinc Concentrate
Payable Zinc (klbs)	23,549		25,268		30,354		35,167		34,503		29,269		32,787		23,061		29,334		35,707			—
Payable Gold (kozs)	0		0		0		0		0		0		0		0		0		0			—
Payable Silver (kozs)	384		389		420		408		380		374		355		361		357		282			—
Lead Concentrate
Payable Lead (klbs)	19,007		18,816		22,942		25,034		26,558		22,992		26,822		24,586		33,769		28,643			—
Payable Gold (kozs)	9		9		12		13		11		9		8		4		4		6			—
Payable Silver (kozs)	11,065		11,328		13,538		12,600		10,838		10,539		9,568		9,870		9,664		7,332			—
Income Statement ($000)
Zinc ($/lb.)	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90	$	0.90		$	0.90
Lead ($/lb)	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95	$	0.95		$	0.95
Gold ($/oz)	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00	$	1,300.00		$	1,300.00
Silver ($/oz)	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00	$	18.00		$	18.00
Revenues
Zinc Concentrate - Zn	$	21,194	$	22,741	$	27,318	$	31,650	$	31,052	$	26,342	$	29,509	$	20,755	$	26,401	$	32,136		$	—
Zinc Concentrate - Au	$	352	$	348	$	400	$	406	$	379	$	341	$	304	$	166	$	166	$	224		$	—
Zinc Concentrate - Ag	$	6,913	$	6,995	$	7,563	$	7,349	$	6,840	$	6,739	$	6,383	$	6,499	$	6,420	$	5,073		$	—
Lead Concentrate - Pb	$	18,057	$	17,875	$	21,795	$	23,783	$	25,230	$	21,843	$	25,481	$	23,357	$	32,080	$	27,211		$	—
Lead Concentrate - Au	$	12,345	$	12,138	$	15,945	$	17,152	$	13,967	$	11,754	$	10,078	$	5,157	$	5,159	$	7,235		$	—
Lead Concentrate - Ag	$	199,176	$	203,901	$	243,684	$	226,797	$	195,085	$	189,696	$	172,233	$	177,666	$	173,960	$	131,972		$	—
Total Revenues	$	258,037	$	263,999	$	316,704	$	307,137	$	272,553	$	256,714	$	243,987	$	233,599	$	244,187	$	203,851		$	—
Operating Cost
Mining	$	60,879	$	58,912	$	59,576	$	62,119	$	58,512	$	56,417	$	57,293	$	57,525	$	59,004	$	53,221		$	—
Process Plant	$	36,384	$	36,136	$	36,136	$	36,136	$	36,136	$	36,136	$	36,136	$	36,136	$	36,136	$	30,372		$	—
General Administration	$	24,084	$	24,001	$	24,001	$	24,001	$	24,001	$	24,001	$	24,001	$	24,001	$	24,001	$	20,173		$	—
Treatment & Refining Charges
Zinc Concentrates
Treatment Charges	$	6,845	$	7,265	$	8,498	$	9,666	$	9,504	$	8,235	$	9,087	$	6,725	$	8,251	$	9,658		$	—
Gold Refining Charges	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—		$	—
Silver Refining Charges	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—		$	—
Transportation	$	7,452	$	7,909	$	9,251	$	10,523	$	10,347	$	8,965	$	9,893	$	7,321	$	8,983	$	10,514		$	—
Penalties	$	17	$	18	$	21	$	24	$	23	$	20	$	22	$	17	$	20	$	24		$	—
Lead Concentrates
Treatment Charges	$	6,171	$	6,130	$	7,026	$	7,496	$	7,848	$	7,037	$	7,910	$	7,394	$	9,702	$	8,229		$	—
Gold Refining Charges	$	63	$	62	$	82	$	88	$	71	$	60	$	52	$	26	$	26	$	37		$	—
Silver Refining Charges	$	12,493	$	12,789	$	15,284	$	14,225	$	12,236	$	11,898	$	10,803	$	11,144	$	10,911	$	8,278		$	—
Transportation	$	3,437	$	3,414	$	3,913	$	4,175	$	4,371	$	3,920	$	4,406	$	4,118	$	5,404	$	4,584		$	—
Penalties	$	166	$	165	$	189	$	202	$	211	$	189	$	213	$	199	$	261	$	222		$	—
Total Operating Cost	157,990		156,800		163,976		168,654		163,261		156,878		159,814		154,605		162,699		145,311			—
Royalty	$	12,776	$	13,066	$	15,708	$	15,149	$	13,346	$	12,610	$	11,875	$	11,445	$	11,826	$	9,757		$	—
Property Tax	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—		$	—
Salvage Value	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—		$	—
Reclamation & Closure	$	312	$	312	$	312	$	312	$	312	$	312	$	312	$	312	$	312	$	262		$	—
Total Production Cost	$	171,078	$	170,179	$	179,996	$	184,115	$	176,919	$	169,800	$	172,001	$	166,362	$	174,837	$	155,330		$	—
Operating Income	$	86,959	$	93,821	$	136,708	$	123,022	$	95,634	$	86,914	$	71,987	$	67,236	$	69,350	$	48,521		$	—
Existing Asset Depreciation	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	43,255	$	36,356		$	—
Expansion Capital Depreciation	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	—	$	0		$	—
Sustaining Capital Depreciation	$	16,164	$	16,036	$	16,246	$	15,291	$	13,654	$	10,927	$	9,345	$	7,390	$	5,783	$	65,747		$	—
Total Depreciation	$	59,419	$	59,290	$	59,500	$	58,546	$	56,909	$	54,182	$	52,599	$	50,644	$	49,038	$	102,103		$	—
Net Income After Depreciation	$	27,540	$	34,530	$	77,208	$	64,476	$	38,725	$	32,732	$	19,388	$	16,592	$	20,312	$	(53,582	)	$	—
Tax Loss Carry Forward Applied	—		—		—		—		—		—		—		—		—		—			—

210

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

			H2 2014			2015			2016			2017			2018			2019			2020			2021			2022			2023
Base Case	Total		1			2			3			4			5			6			7			8			9			10
Taxable Income	$	1,534,132	$	81,881		$	158,440		$	151,367		$	148,541		$	145,914		$	146,341		$	158,353		$	137,362		$	114,660		$	13,350
Income Taxes - ( % of Gross Revenue)	$	379,552	12,784			25,342			25,399			25,421			25,427			25,663			26,856			24,572			22,824			15,333
Net Income After Taxes	$	1,154,580	69,097			133,098			125,968			123,120			120,487			120,677			131,497			112,790			91,835			(1,983		)
Cash Flow
Operating Income	$	2,681,614	$	104,714		$	204,990		$	204,874		$	203,593		$	202,485		$	205,043		$	219,343		$	201,537		$	178,380		$	76,501
Working Capital
Account Recievable (60 days)	$	—	$	(5,004	)	(4,915		)	(9,963		)	(17		)	(5		)	(185		)	(933		)	1,788			1,368			5,864
Accounts Payable (30 days)	$	—	$	6,980		$	7,087		$	135		$	132		$	102		$	41		$	382		$	(1,587	)	$	(28	)	$	(522	)
Inventory - Parts, Supplies	$	—	$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—
Total Working Capital	$	—	1,977			2,172			(9,829		)	115			97			(144		)	(551		)	200			1,340			5,342
Capital Expenditures
Expansion Capital
Mine	$	12,779	$	2,624		$	5,610		$	4,545		$	—		$	—		$	—		$	—		$	—		$	—		$	—
Process Plant	$	11,535	$	2,925		$	8,610		$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—
Owners Cost	$	—	$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—		$	—
Sustaining Capital
Mining	$	247,786	$	8,811		$	23,938		$	14,658		$	16,543		$	16,165		$	21,297		$	10,474		$	9,958		$	14,555		$	15,935
Process Plant	$	46,483	$	1,811		$	4,614		$	2,363		$	2,548		$	4,004		$	1,936		$	1,886		$	2,192		$	2,498		$	2,366
G&A	$	6,910	$	3,820		$	2,205		$	810		$	25		$	25		$	25		$	—		$	—		$	—		$	—
Total Capital Expenditures	$	325,493	$	19,991		$	44,977		$	22,376		$	19,117		$	20,194		$	23,258		$	12,360		$	12,151		$	17,054		$	18,301
Current Cash Net of Debt			$	30,000
Cash Flow before Taxes	$	2,392,101	$	116,829		$	162,469		$	172,982		$	184,904		$	182,700		$	181,953		$	206,744		$	189,899		$	162,979		$	63,854
Cummulative Cash Flow before Taxes			$	116,829		$	279,298		$	452,279		$	637,183		$	819,883		$	1,001,836		$	1,208,581		$	1,398,480		$	1,561,458		$	1,625,312
												—			—			—			—			—			—			—
Taxes
Income Taxes	$	379,552	$	12,784		$	25,342		$	25,399		$	25,421		$	25,427		$	25,663		$	26,856		$	24,572		$	22,824		$	15,333
Cash Flow after Taxes	$	2,012,550	$	104,045		$	137,126		$	147,582		$	159,482		$	157,273		$	156,290		$	179,889		$	165,327		$	140,154		$	48,521
Cummulative Cash Flow after Taxes			$	104,045		$	241,171		$	388,753		$	548,236		$	705,509		$	861,798		$	1,041,687		$	1,207,014		$	1,347,168		$	1,395,689
																					—			—			—			—

	2024			2025			2026			2027		2028			2029			2030		2031			2032			2033			2034
Base Case	11			12			13			14		15			16			17		18			19			20			21
Taxable Income	$	27,540		$	34,530		$	77,208		$	64,476	$	38,725		$	32,732		$	19,388	$	16,592		$	20,312		$	(53,582	)	$	—
Income Taxes - ( % of Gross Revenue)	16,260			16,630			19,992			19,281		16,986			16,049			15,113		14,567			15,051			—			—
Net Income After Taxes	11,280			17,900			57,216			45,196		21,739			16,683			4,275		2,025			5,261			(53,582		)	—
Cash Flow
Operating Income	$	86,959		$	93,821		$	136,708		$	123,022	$	95,634		$	86,914		$	71,987	$	67,236		$	69,350		$	48,521		$	—
Working Capital
Account Recievable (60 days)	(726		)	(290		)	(2,632		)	557		1,796			733			733		428			(379		)	2,061			9,721
Accounts Payable (30 days)	$	263		$	(98	)	$	590		$	385	$	(443	)	$	(525	)	$	241	$	(428	)	$	665		$	(1,429	)	$	(11,943	)
Inventory - Parts, Supplies	$	—		$	—		$	—		$	—	$	—		$	—		$	—	$	—		$	—		$	—		$	—
Total Working Capital	(462		)	(388		)	(2,042		)	942		1,353			209			974		(0		)	286			632			(2,223		)
Capital Expenditures
Expansion Capital
Mine	$	—		$	—		$	—		$	—	$	—		$	—		$	—	$	—		$	—		$	—		$	—
Process Plant	$	—		$	—		$	—		$	—	$	—		$	—		$	—	$	—		$	—		$	—		$	—
Owners Cost	$	—		$	—		$	—		$	—	$	—		$	—		$	—	$	—		$	—		$	—		$	—
Sustaining Capital
Mining	$	23,594		$	13,590		$	13,448		$	10,352	$	8,967		$	11,942		$	7,533	$	3,318		$	1,780		$	928		$	—
Process Plant	$	1,886		$	1,886		$	2,192		$	2,498	$	2,366		$	1,886		$	1,886	$	1,886		$	1,886		$	1,886		$	—
G&A	$	—		$	—		$	—		$	—	$	—		$	—		$	—	$	—		$	—		$	—		$	—
Total Capital Expenditures	$	25,480		$	15,476		$	15,640		$	12,851	$	11,334		$	13,828		$	9,419	$	5,205		$	3,666		$	2,814		$	—
Current Cash Net of Debt
Cash Flow before Taxes	$	61,328		$	78,269		$	119,338		$	111,425	$	85,965		$	73,606		$	63,854	$	62,343		$	66,282		$	46,601		$	(2,223	)
Cummulative Cash Flow before Taxes	$	1,686,641		$	1,764,910		$	1,884,247		$	1,995,673	$	2,081,638		$	2,155,244		$	2,219,098	$	2,281,441		$	2,347,723		$	2,394,324		$	2,392,101
	—			—
Taxes
Income Taxes	$	16,260		$	16,630		$	19,992		$	19,281	$	16,986		$	16,049		$	15,113	$	14,567		$	15,051		$	—		$	—
Cash Flow after Taxes	$	45,069		$	61,639		$	99,345		$	92,145	$	68,979		$	57,557		$	48,741	$	47,777		$	51,231		$	46,601		$	(2,223	)
Cummulative Cash Flow after Taxes	$	1,440,758		$	1,502,397		$	1,601,742		$	1,693,887	$	1,762,866		$	1,820,423		$	1,869,164	$	1,916,941		$	1,968,172		$	2,014,773		$	2,012,550
	—			—			—			—		—			—			—		—			—			—			—

Economic Indicators before Taxes
NPV @ 0%	0	%	$	2,379,317
NPV @ 5%	5	%	$	1,778,074
NPV @ 10%	10	%	$	1,414,897
IRR			N/A
Payback			N/A
Economic Indicators after Taxes
NPV @ 0%	0	%	$	2,012,550
NPV @ 5%	5	%	$	1,515,689
NPV @ 10%	10	%	$	1,214,041
IRR			N/A
Payback	Years		N/A

211

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

23ADJACENT PROPERTIES

There are no properties adjacent to the Project with identified resources or reserves.

The Company has identified prospective areas for exploration within their concession holdings surrounding the Escobal exploitation concession. Five core holes were drilled in 2009 on the San Juan Bosco prospect six km west of Escobal; three core holes were drilled on the Morales prospect seven kilometers north of Escobal in 2010 and 2011; and 26 core holes were drilled at the Varejones project 20 km east of Escobal in 2012. Ongoing reconnaissance prospecting and exploration continues on a regional level throughout Tahoe’s other concessions.

Regional exploration has identified a number of additional targets surrounding the Escobal Project on exploration or reconnaissance concessions which are under application or have been approved. These targets are illustrated in Figure 23-1 and Figure 23-2.

The Neque target parallels the Escobal vein and is exposed in an 800 m diameter window of post-mineral tuff where grades of up to 375 g Ag/t and 16.0 g Au/t were produced through rockchip sampling. Previous work on this property by Entre Mares includes prospecting in 2006, rock sampling in 2007, soil geochemistry lines in 2007 and geological mapping in 2007. No drilling has been conducted on this property.

The San Nicholas area contains a high-sulphidation alteration target with surface gold values reportedly up to 30 g/t. No drilling has been conducted in the area. The area has been explored by Entre Mares by prospecting in 2000, rock chip sampling in 2007 and 2008, soil sample lines in 2007 and 2008, and geological mapping in 2008. Tahoe carried out some field work, community relations and drill site preparation in 2010 and 2011. Environmental baseline studies commenced in 2011 and a drill plan has been approved by MARN.

The Las Flores target is comprised of a series of east-west trending quartz veins exposed intermittently over a 5km strike length where grades of up to 150 g Ag/t and 7.0 g Au/t are reported. Previous work on this property by Entre Mares included prospecting in 2006 and 2008, rock sampling from 2006 to 2009, soil geochemistry lines in 2008 covering a portion (approximately 5%) of the west area, and preliminary surface mapping in 2008. This property has not been drill tested.

212

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Figure 23-1: Regional Exploration Targets

The San Juan Bosco vein located 6 km west of and parallel to the Escobal vein is a 5 m wide vein that can be traced along a 1 km strike length. A reconnaissance surface sampling program generated results grading up to 15 g Ag/t and 1 g Au/t. Previous work by Entre Mares on this property included prospecting in 2008, rock sampling in 2008, soil geochemistry lines in 2008 and geological mapping in early 2009. Five drill holes were drilled in 2009 testing a portion of the vein where permitted on the Oasis exploration license; JB09-01 contained an 86.6 m interval grading 0.24 g Au/t, which terminated in mineralization. Drill hole JB09-04 contained a single 3m interval grading 2.37 g Au/t. Additional drilling is warranted to test the higher-grade portion of the vein.

213

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

Figure 23-2: Escobal District Exploration Targets

The Varejones target is comprised of numerous red-bed hosted low-sulphidation veins exposed over a 3 km northeast trend where surface results generated values of up to 200 g Ag/t and 4.9 g Au/t. Previous work by Entre Mares on this property includes prospecting in 2001 and 2006, rock sampling in 2007, and soil geochemistry lines in 2007.

In 2012, a total of 26 drill holes for 8,085 m were completed on four target areas along the three-kilometer-long Varejones vein trend. Drilling identified vein mineralization that was not deemed economically viable and the drilling was suspended.

214

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

24OTHER RELEVANT DATA AND INFORMATION

There is no additional information or explanation necessary to make the technical report understandable and not misleading.

215

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FORM 43-101F1 TECHNICAL REPORT — FEASIBILITY STUDY

25INTERPRETATION AND CONCLUSIONS

The Escobal mine reached commercial production January 1, 2014 and production at design levels in the second quarter of 2014. Through June 2014, approximately 17,000 meters of mine development and 230 meters of vertical development had been completed. As of June 30, 2014, the mine produced 786,551 tonnes of ore grading 566 g/t silver, 0.46 g/t gold, 1.01% lead and 1.39% zinc. Saleable lead and zinc concentrates containing nearly 12 million ounces of silver have been produced as of the effective date of this Study.

The results of this Study conclude:

·The Escobal mine Feasibility Study demonstrates the economic viability of the Escobal mine from July 1, 2014 through the end of the estimated mine life and supports the declaration of Proven and Probable Mineral Reserves.

·The Escobal Proven and Probable reserves total 31.4 million tonnes at average grades of 347 g/t silver, 0.33 g/t gold, 0.74% lead and 1.21% zinc containing 350.5 million ounces of silver, 335,600 ounces of gold, 232,100 tonnes of lead and 381,600 tonnes of zinc.

·Operating results to date validate the development and mining methods and process design in use at the Escobal mine.

·Expanding mine and mill production to meet the 4,500 t/d mill throughput rate requires the purchase of additional underground equipment to increase development and production capabilities, increasing the capacity of the paste backfill plant, addition of a fourth tailings filter and infrastructure upgrades and minor ancillary equipment additions in the process plant. The Study updates the cost and confirms that the expansion of the mill throughput rate earlier than previously estimated is achievable.

·The Escobal Mineral Resource Estimate is supported by the geologic model and is based on sufficient drill sample analytical and density measurements, detailed drill-hole lithology and alteration data, and metallurgical testing.

·An independent verification program including a complete audit of the drill hole assay database, drill location and survey data, sample verification, sample handling and logging procedures, and QA/QC analysis support the estimation of the Escobal resource and the assignment of Measured and Indicated classification to much of the stated resource.

·The recent infill drill program and underground development has increased the confidence in the resource resulting in the transition of a significant portion of the previously Inferred sulfide resources to Indicated and a portion of Indicated sulfide resource to Measured. The Measured resources are spatially associated with the closely-spaced underground drilling.

·There are no Measured resources within the oxide portions of the deposit, nor any Indicated resources within the gold-dominant oxide portions of the East zone, due to some uncertainty in the density data, the limited amount of metallurgical testing of these material types, and spatial uncertainty of the high-grade gold. None of these issues detract from the overall confidence in the global project resource estimate, but they do preclude the classification of Measured or Indicated in these specific areas. Approximately 2% of the Indicated resource tonnes and 9% of the Inferred resource tonnes are oxide.

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26RECOMMENDATIONS

The recommendations for this report are as follows:

·Tahoe holds exploration and exploitation licenses near the existing facilities where the regional geology appears to be a favorable environment for the formation of high grade mineral deposits. M3 recommends Tahoe continue to aggressively explore for extensions or offsets of the Escobal vein and accelerate district exploration with the goal of discovering additional high grade feed for the Escobal flotation plant. Exploration should begin to investigate vein systems ancillary to the Escobal vein under Quaternary cover.

·Concurrent with the prior recommendation, M3 recommends that Tahoe initiate studies to investigate increased mining and plant throughput rates to increase the annual silver production particularly in the second half of the mine life.

·Generating power with diesel generators is an acceptable but higher cost source of power for the Escobal mine and processing plant. M3 believes there are several lower cost alternatives that could potentially reduce operating cost by as much as 10%. M3 recommends Tahoe investigate these alternatives fully and implement the best solution to replace diesel power generation.

·Escobal has been in commercial production for six months as of the effective date of this report and is operating at or above design parameters. M3 recommends Tahoe systematically evaluate mining, development, processing and surface operations to optimize processes/procedures and reduce operating and capital costs.

·Metal recoveries to date are at design parameters for silver and lead but below design for gold and zinc. M3 recommends that Tahoe continue in-house metallurgical studies to improve and optimize recoveries of all four metals. Priority should continue to be given to silver but incremental recovery in all three byproducts can enhance mine economics.

·M3 and Blattman believe there are opportunities to reduce mining dilution and ore loss compared to this feasibility study. M3 recommends Tahoe measure in a timely manner and critically evaluate mining dilution and ore loss. Tahoe should act to reduce excess dilution which will in turn improve utilization of mining and milling capacity and mine economics.

·M3 recommends Tahoe perform annual geotechnical and water management performance reviews of the dry stack tailings facility to ensure stability and reclamation success.

·As additional drill data becomes available, the resource model should be refined and resource estimates updated. Additional infill drilling will lend further confidence to the resource model.

·Drilling from underground platforms should be utilized to test for deeper extensions of “ore shoots” in the Central and East zones. MDA recommends that future drilling from underground include drill orientations subparallel to the Escobal vein to test for mineralized cross-structures as appear to be indicated by structural trends observed in the underground workings.

·Tahoe has followed MDA’s recommendation to sample the full length of underground drill holes collared in close proximity to the ore body as unsampled intervals create modeling and estimation concerns particularly when the sample intervals end in mineralization. Tahoe should continue this practice.

·Tahoe should continue the use of suitable standards for each of the important metals. Exploration and development drilling would benefit from more types of quality control sampling and analyses, including the

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collection and analysis of half- or quarter-core duplicates that are processed and analyzed at the primary assay lab, and the use of coarse reject or preparation duplicates analyzed at the primary assay lab.

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27                   REFERENCES

AMEC Americas Ltd., 2010, Escobal Project Guatemala NI 43-101 Technical Report, prepared under the guidance of Mr. Greg Kulla, P. Geo., a Qualified Person. The effective date of the report and resource estimate is April 30, 2010.

Barton, N.R., Lien, R. and Lunde, J. 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mech. 6(4), 189-239.

Basterrechea, A.M., 2011, Estudio de Evaluación de Impacto Ambiental Proyecto Minero Escobal. Asesoría Manuel Basterrechea Asociados, S.A.

Bieniawski, Z.T. 1976, Rock Mass Classification in Rock Engineering, in Exploration for Rock Engineering, proc. of the symp., (ed.Z.T. Bieniawski) 1, 97-106. Cape Town: Balkema.

Caceras, C., 2013, Numerical Modeling, Testing and Basic Calibration of Input Parameters, Rock Mechanics Consulting Services.

Corbett, G.J., 2002, Epithermal Gold for Explorationists, AIG News No 67, 8p.

Czarnowsky, Inc., 2011, Reclamation and Closure Cost Estimate, Escobal Project, San Rafael, Guatemala.

Davies, Michael P., and Rice, Stephen, “An Alternative to Conventional Tailings Management — “Dry-Stack” Filtered Tailings”, AMEC Earth and Environmental, 2004.

Golder Associates, Inc., 2014, Draft Geotechnical Assessment: EL. 1190 to 1265 m, Escobal mine Guatemala.

Hazen Research, Inc., 2014, Flotation Pilot Plant for Minera San Rafael — Escobal.

ISO (the International Organization for Standardization), 2014, www.iso.org.

Kappes, Cassiday & Associates, “2,500 Tonne Per Day Flotation Plant and Cyanidation Plant Cost Comparisons”, 22 July, 2009.

Langston, R., 2013-2014, various geotechnical investigations and ground support recommendations, internal memorandums to Tahoe Resources Inc., Langston and Associates.

M3 Engineering & Technology Corporation, 2010, Escobal Guatemala Project, NI 43-101 Technical Report Preliminary Economic Assessment, prepared under the guidance of Mr. Conrad Huss, P.E., a Qualified Person. The effective date of the report is November 29, 2010; the effective date of the resource estimate is August 24, 2010.

M3 Engineering & Technology Corporation, 2011, Escobal Reclamation and Closure Plan.

M3 Engineering & Technology Corporation, 2012, Escobal Guatemala Project, NI 43-101 Technical Report Preliminary Economic Assessment, prepared under the guidance of Mr. Conrad Huss, P.E., a Qualified Person. The effective date of the report is May 7, 2012; the effective date of the resource estimate is January 23, 2012.

M3 Engineering & Technology Corporation, 2013, Revision 1 Escobal Guatemala Project, NI 43-101 Technical Report Preliminary Economic Assessment, prepared under the guidance of Mr. Conrad Huss, P.E., a Qualified Person. The effective date of the report is May 7, 2012; the effective date of the resource estimate is January 23, 2012.

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McClelland Laboratories, Inc., “Report on Scoping Metallurgical Testing — Escobal Drill Core Composites, MLI Job No. 3324”, 20 May, 2009.

Pakalnis, R., 2010, Preliminary Geotechnical Assessment—Escobal Project, Pakalnis & Associates, report no. Tahoe 1/10.

Pakalnis. R. 2011-2012, various geotechnical and geomechanical site investigations, internal memorandums to Tahoe Resources Inc., Pakalnis & Associates.

Performance Associates International, Inc., 2013, Computer-Based Introductory Training Modules, prepared for Mineral San Rafael, S.A. (English & Spanish language editions).

Performance Associates International, Inc., 2013, Plant-Specific Hard-Copy Training Manuals, prepared for Mineral San Rafael, S.A. (English & Spanish language editions).

Reidel J., Cluff, T., Huang, Y., 2014, Escobal Project Hydrogeological Characterization and Groundwater Modeling (draft), Schlumberger Water Services USA, Inc.

Sheorey, P.R. 1994, A Theory for In Situ Stresses in Isotropic and Transversely Isotropicrock, in Int. J. Rock Mech. Min. Sci. & Geomech, Abstr. 31(1), 23-34.

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APPENDIX A: FEASIBILITY STUDY CONTRIBUTORS AND PROFESSIONAL QUALIFICATIONS

221

CERTIFICATE of QUALIFIED PERSON

Conrad E. Huss

I, Conrad E. Huss, P.E., Ph.D., do hereby certify that:

1.             I am Senior Vice President and Chairman of the Board of:

M3 Engineering & Technology Corporation

2051 W. Sunset Rd., Suite 101

Tucson, Arizona 85704

U.S.A.

2.I graduated with a Bachelor’s of Science in Mathematics and a Bachelor’s of Art in English from the University of Illinois in 1963. I graduated with a Master’s of Science in Engineering Mechanics from the University of Arizona in 1968. In addition, I earned a Doctor of Philosophy in Engineering Mechanics from the University of Arizona in 1970.

3.I am a Professional Engineer in good standing in the State of Arizona in the areas of Civil (No. 9648) and Structural (No. 9733) engineering. I am also registered as a professional engineer in the States of California, Illinois, Maine, Minnesota, Missouri, Montana, New Mexico, Oklahoma, Texas, Utah, and Wyoming.

4.I have worked as an engineer for a total of forty-four years. My experience as an engineer includes over 36 years designing and managing mine development and expansion projects including material handling, reclamation, water treatment, base metal and precious metal process plants, industrial minerals, smelters, special structures, and audits.

5.I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

6.I am the principal author for the preparation of the technical report titled “Escobal Mine Guatemala NI 43-101 Feasibility Study” (the “Technical Report”), dated November 5, 2014, prepared for Tahoe Resources Inc.; and am responsible for Sections 1 and 22. I have visited the project site on 1 December 2010.

7.I have prior involvement with the property that is the subject of the Technical Report. I was a contributing author of a previous Technical Report on the subject property entitled “Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment” dated July 24, 2013. M3 Engineering & Technology Corporation was the EPCM contractor for the Escobal Project.

8.As of the date of this certificate, 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.

9.I am independent of the issuer applying all of the tests in Section 1.5 of National Instrument 43-101.

10.I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.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 accessible by the public, of the Technical Report.

Signed and dated this 5th day of November, 2013.

/s/ Conrad Huss

Signature of Qualified Person

Conrad Huss

Print Name of Qualified Person

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

Resume

Project Manager and Project Sponsor Founded M3 in 1986 Education Ph.D., Engineering Mechanics, University of Arizona Master of Science, Engineering Mechanics, University of Arizona Bachelor of Arts, English, University of Illinois Bachelor of Science, Mathematics, University of Illinois Registrations Civil Engineer — Arizona, California Structural Engineer — Arizona, Utah Professional Engineer — Illinois, Maine, Minnesota, Missouri, Montana, New Mexico, Oklahoma, Texas, Wyoming

Forty-three years of design in industrial, municipal, and commercial projects, including material handling, reclamation, water treatment, base metal and precious metal process plants, industrial minerals, smelters, chemical plants, special structures and audits. Career highlights include thirty-five years of design/construct experience, plant startups in South America and Mexico, oceanography/surveying in Alaska and Hawaii, and six years of university teaching. RELEVANT PROJECT EXPERIENCE M3 ENGINEERING & TECHNOLOGY CORPORATION PROJECT MANAGER AND PROJECT SPONSOR (26 YEARS) ·Goldcorp, Cerro Negro - Argentina ·Tahoe Resources Polymetallic — Guatemala ·Minera Peñasquito NPSC — Mexico ·Casino Copper, 110,000 MTPD — Yukon, Canada ·Carmacks Copper SX-EW — Yukon, Canada ·Resolution Copper Mining Stage 2 (Pre-Feasibility) with Mill Trade-off Studies - Arizona ·Goldcorp, Minera Peñasquito 130,000 MTPD - Mexico ·Global Alumina, Republic of Guinea 43-101 Study - Africa ·Pan American, Alamo Dorado 43-101 Study - Mexico ·Pan American, Manantial Espejo Silver/Gold - Argentina ·Phelps Dodge Bagdad, Primary Crusher Relocation - Arizona ·Kinross, Refugio Reopening - Chile ·Climax Molybdenum, Moly Metal Plant - Arizona ·Phelps Dodge Safford, CASC Design - Arizona ·Phelps Dodge Safford, Copper Leach - Arizona ·CEMEX Victorville, Clinker Hall - California ·Western Silver, Peñasquito Scoping, NI 43-101 Study and Feasibility Studies - Zacatecas, Mexico ·Frontera Copper, Piedras Verdes Copper Leach NI 43-101 Study - Mexico ·AVESTOR, Lithium Vanadium Polymer Battery Plant with Laboratories - Nevada ·Kennecott Utah, Lime Plant Feasibility Study - Utah ·Phelps Dodge, Arizona Closure/Closeout Plans at 7 properties - Arizona ·Alamos Gold, Mulatos Prefeasibility - Mexico ·Teck Cominco, Glamis Gold Feasibility - Mexico ·Kennecott Rawhide, Conceptual Closeout Plan for Leach Pile - Nevada

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

·Phelps Dodge El Abra, Structural and Material Handling Audit — Chile

·Kennecott Utah, Bid Call for Restructure of Maintenance

·Workforce - Utah

·Phelps Dodge El Abra, SX-EW ER Tank Replacement - Chile

·Fischer-Watt, Copper SX-EW Prefeasibility - Mexico

·Kerr McGee, 1200 MTPY BLVO Plant Apex - Nevada

·Billiton/BHP Worsley ,Alumina Plant Audit - Australia

·Phelps Dodge Tyrone ,Closure/Closeout Plans with Water Treatment Plant - New Mexico

·Mitsubishi Cement, Lucerne Valley Plant Upgrades - California

·Phelps Dodge Chino, Closure/Closeout Plans with Water Treatment Plant - New Mexico

·Mitsubishi Cement Longbeach, Ocean Port - California

·Phelps Dodge Cobre, Closure/Closeout Plans - New Mexico

·Phelps Dodge El Abra, Material Handling and Structural Audit - Chile

·Billiton/ALCOA, Alumar Alumina Refinery Plant Material Handling/Structural Audit - Brazil

·Peñoles F.I. Madero, 8,000 TMPD Greensfield Silver/Lead/Zinc - Mexico

·Kennecott Greens Creek, Flotation Expansion, Silver/Lead/Zinc - Alaska

·Phelps Dodge Henderson, Material Handling and Structural Audit - Colorado

·Kennecott Greens Creek, Pyrite Circuit for Reclamation - Alaska

·Phelps Dodge Morenci, Coronado Leach - Arizona

·Cyprus Cerro Verde, Crush/Convey - Peru

·California Portland Cement, RIMOD 3 Expansion - Arizona

·Phelps Dodge Candelaria, Material Handling and Structural Audit - Chile

·Minera Alumbrera, Startup and Performance Test for Copper/Gold Plant - Argentina

·Echo Bay Gold, Aquarius Feasibility Study - Canada

·Arizona Portland Cement, Expansion - Arizona

·Minera Alumbrera, SAG Mill Run In — 3 month field assignment - Argentina

·Phelps Dodge Ajo, Open Air Copper Mill - Arizona

·Echo Bay Paredones Gold, Amarillos EPCM Basic Engineering - Mexico

·Cyprus Sierrita, Inpit Crush/Convey - Arizona

·Kennecott, Smelter Upgrade following Audit - Utah

·Battle Mountain Crown Jewel, Gold and Silver Detail Engineering - Washington

·Kennecott Greens Creek, Reopening and Reclamation, Lead/Zinc/Silver - Alaska

Page 2 of 8 12-12

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

·Cyprus Bagdad, Material Handling and Structural Audit — Arizona

·Hecla Rosebud, Precious Metal Detail Engineering and Reclamation - Nevada

·Phelps Dodge Morenci, Ball Mills A-7 and B-32, Copper Concentrator — Arizona

·Phelps Dodge Hidalgo, Smelter Upgrade Simulation - Arizona

·Kerr-McGee West Chicago, Physical Separation Reclamation Facility with Water Treatment - Arizona

·Lluvia Del Oro Gold, Plant Detail Engineering - Mexico

·Cyprus Miami, Smelter Modifications for Ancillaries - Arizona

·Phelps Dodge Chino, Smelter Upgrade including Uptake Shaft - Arizona

·Cyprus Miami, Smelter Casting Furnace Upgrade - Arizona

·Phelps Dodge Morenci, Material Handling and Structural Audit - Arizona

·Geomaque Gold, Detail Engineering - Sonora, Mexico

·Phelps Dodge Morenci, Smelter Equipment Relocation - Arizona

·Phelps Dodge Hidalgo, Smelter Fugitive Gas Collection System - Arizona

·Magma San Manuel, Smelter Anode Press Plant - Arizona

·Phelps Dodge Ajo, Smelter Demolition - Arizona

·Penmont, La Herradura Gold Plant - Mexico

·Phelps Dodge Hidalgo, Smelter Upgrade including Reaction Shaft - New Mexico

·Cyprus Casa Grande, Roaster Upgrades, Copper - Arizona

·Zinc Corp, Roaster Upgrade - Oklahoma

·Cyprus Bagdad, Water Flush Crusher, Copper - Arizona

·ASARCO Hayden, Smelter Dust System - Arizona

·Placer Dome, Mulatos Gold Plant Basic Engineering - Mexico

·Cyprus Bagdad, 156,000 TPD Feasibility Study - Arizona

·Cyprus Bagdad ,Feasibility Study for Inpit Crushing and Mill Expansion - Arizona

·Hecla, La Choya Gold Plant - Sonora, Mexico

·Chemstar Lime Plants - Western United States

·Majdanpek, Crush/Convey for Copper Mine - Yugoslavia

·Phelps Dodge Chino, SX-EW Expansion - New Mexico

·Magma, McCabe Gold Plant Expansion - Arizona

·Phelps Dodge Morenci, Flotation Expansion, Copper - Arizona

·Phelps Dodge Chino, Waterflush Crusher, Copper - New Mexico

·Granite Sand & Gravel Plant - Arizona

·Kerr-McGee, Manganese Dioxide Chemical Plant - Nevada

·Cyprus Sierrita, Acid Plant (Rhenium Recovery) - Arizona

Page 3 of 8
12-12

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

·Phelps Dodge Chino, Conveyor System Rebuild - New Mexico

·Molycorp, Mountain Pass Crush/Convey System for Rare Earths - California

·Cyprus Esperanza/Twin Buttes, Cross Country Conveyor Upgrade - Arizona

·Mt. Graham, Utilities and Tankage - Arizona

·Old Tucson, Utility Inventory and Upgrade - Arizona

·ASDM, Utility Inventory and Upgrade - Arizona

·Cyprus Twin Buttes, Fuel Stations Demolition and Upgrade - Arizona

·Cyprus Sierrita, Fuel Station Demolition and Upgrade - Arizona

·Cyprus Sierrita ADM Chemical Plant - Arizona

·ASARCO Mission, Mill Feed Upgrade, Copper - Arizona

·Cyprus Miami, Road and Bridge - Arizona

·ASARCO Mission, Dust Collection, 96,000 CFM - Arizona

·Magma San Manuel, No. 4 Head Frame Upgrade - Arizona

·Cyprus Sierrita, Ferro Moly Dust Collection - Arizona

·St. Cloud, Flotation Upgrade Lead/Zinc/Silver - New Mexico

·University of Arizona Optical Mirror Laboratory - Arizona

·Mt. Graham SMT Telescope Facility - Arizona

·Mt. Graham Observatory, Site Programming, Utilities and Maintenance Building - Arizona

·Phelps Dodge Chino, Inpit Crush/Convey Study - New Mexico

·Philippines Crush/Convey Study

·Cyprus Bagdad, Tankhouse Expansion - Arizona

·Phelps Dodge Morenci, Inpit Crush/Convey Checking - Arizona

·Cyprus Sierrita, Inpit Crush/Convey - Arizona

·AZANG, Maintenance Hangar and Hush House - Arizona

·Cyprus Sierrita, Column Cell Expansion I & II, Copper/Moly - Arizona

·Cyprus Sierrita, Moly Roaster Feed Systems I & II - Arizona

·Magma Pinto Valley, #4 Tailing Dam Slurry Pump Station - Arizona

·Ft. Huachuca, General Instruction Building - Arizona

·Mt. Bell Communication Centers - Arizona

·Cyprus Sierrita, Moly Packaging System Upgrade — Arizona

RGA Engineering Corporation, Structural Engineer, V. President, Engineering Director (4 Years)

·Coronado Post Office for USPS - Arizona

·Amphitheater Elementary School and University of Arizona Science Building - Arizona

·Tanque Verde and Campbell Avenue Street Lighting - Arizona

Page 4 of 8 12-12

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

·Reid Park Band Shell and Master Plan - Arizona

·AZANG Engine Shop, General Purpose Shop, Hush House -Arizona

·Northern Arizona University Information Center - Arizona

·Davis-Monthan Combat Support Center - Arizona

·Ft. Huachuca Communications Facilities - Arizona

·University of Texas Submillimeter Telescope - Texas

·Arizona-Sonora Desert Museum Mountain Habitat - Arizona

·Ina Road Bridge, Tanque Verde Bridge, Clifton Bridge, I-17 AC/DC Bridge, Orange Grove Bridge - Arizona

·University of Arizona Submillimeter Telescope - Arizona

·University Heights Shopping and Parking Complex - Arizona

·La Paloma Resort Hotel and Office Complex - Arizona

·Design of warehouses and greenhouses - Worldwide

·Design of banks, apartments, office buildings - Arizona

·Design of elephant enclosure - Arizona

·Design of conveyor head frames and maintenance shops -Arizona

·Design of schools, libraries, churches - Arizona

·Project Engineer for conversion of U.S.P.S. power system in - Phoenix, AZ with 2-350 ton chillers

·Analysis for parking garage and pedestrian bridge - Arizona

·Finite element earthquake analysis of ten-story office building - Arizona

·Design of eight-story reinforced concrete hotel - Mexico

·Converter Blower Electrification, Project Manager - Arizona

Mountain States Engineers, Vice President and Manager of Engineering (4 Years)

·52,000-65,000 TPD Mill Expansion, SPCC - Cuajone, Peru

·Pennsylvania Fuels Group, Coal Gasification Plant Study

·Gold Mill Expansion, Newmont - Nevada

·Shuichang Bethlehem International Crush/Convey, 10,000 TPH - China

·Gold Heap Leach Study, Newmont - Australia

·Fly Ash Disposal System, Tucson Electric - Arizona

·Concentrate Loadout, Cyprus Bagdad - Arizona

·Tailings System, Cyprus Bagdad - Arizona

·No. 19 Dump Leach System — Arizona

·8000 TPH Crushing/Conveying of Waste, Kennecott Copper Corp. - Arizona

·150,000 TPD Crush/Convey, Kennecott - Arizona

·50,000 TPD Crushing/Conveying System, Island Copper - BC, Canada

·10,000 TPD Limestone Loadout, Grupo Cementos - Mexico

·40,000-54,000 TPD Mill Expansion, Cyprus Bagdad - Arizona

·Smelter Coal Conversion, Phelps Dodge - Hidalgo, New Mexico

Page 5 of 8 12-12

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

·Modification of 3000 TPH Wash Plant, Carbon Coal - New Mexico

·Moly By-Product Plant, Phelps Dodge - Ajo, Arizona

·50 TPD Zinc Skimmings Plant, National Zinc - Oklahoma

·4000 TPH Portable Crusher, Duval Corp. - Sierrita, Arizona

·Sulfur Unloading Facility, Duval Corp. - Galveston, Texas

Mountain States Engineers, Piping Department Head (1 Year)

·Gulf & Western Sonora Gold - California

·1000 TPD Moly By-Product Plant, ASARCO, Mission - Arizona

·40,000 TPY Flotation Retrofit, ASARCO, Mission - Arizona

·Tailings System, Plateau Resources - Utah

·2600 TPH Crush/Convey, Climax Molybdenum Co. - Colorado

·2000 TPD Moly By-Product Plant, La Caridad - Mexico

·6600 TPH Crushing/Conveying - Majdanpek, Yugoslavia

·Coal Loadout Facility, CF&I, Maxwell Mine - Colorado

·12,000 TPD Uranium/Vanadium Plant, Cotter Corp. - Colorado

·Lined Tailings Pond, Cotter Corp. - Colorado

·Spent Catalyst Plant, Cotter Corp. - Colorado

·750 TPD Uranium Mill, Plateau Resources - Utah

·Potash Plant Modifications, Duval Corp. - Carlsbad, New Mexico

·Feasibility Study for Urangesellschaft, Site Determination and Environmental

·Delamar Silver Plant CCD — Idaho

Mountain States Engineers, Structural Engineer and Department Head (5 Years)

·Round Mountain, Nevada, Heap Leach Gold including recovery pad Lime Plant, Nafinsa, Job Engineer - Santa Rita, Arizona

·250,000 TPD Crushing/Conveying, Job Engineer, Duval Corporation - Sierrita, Arizona

·Ferro-Moly Plant, Job Engineer, Duval Corp. - Sierrita, Arizona

·Special Investigations of Towers, Thickener and Frames

Hughes Aircraft Company, Structural Engineer (½ Year)

·Finite element analysis of missiles and vibration isolation of missile components

·Strain gauge layout and destructive testing

Page 6 of 8 12-12

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

LTJG U.S. Coast and Geodetic Survey (3 Years)

·3rd in Command, USC & GSS Hydrographer, Caribbean

·Hydrographical surveys in Hawaii, Alaska and Florida

·Reconnaissance Alaskan Good Friday Earthquake

·Photogrammetry and land surveying in Alaska

Teaching

University of Arizona (2 Years)

Adjunct Lecturer

Northern Arizona University (1 Year)

Assistant Professor of Engineering

University of Arizona (4 Years)

Graduate Fellow in Engineering Mechanics

West Tampa Junior High School (½ Year)

Science and Math Teacher

Courses

·Cold Regions Engineering Short Course

·Completed MSHA Training

Publications

HUSS, Conrad, and Dan Neff; “Horizontally Stiffened Angular Hoppers Analyzed by Beam Action Versus Finite Element”, Bulk Solids Handling, June 1984.

HUSS, Conrad, and Nikita G. Reisler; “A Comparison of Handling Systems for Overburden of Coal Seams”, Bulk Solids Handling, March 1984.

HUSS, Conrad, and Dan Neff; “Horizontally Stiffened Membrane Hoppers Analyzed by Virtual Work Versus Finite Element”, Bulk Solids Handling, November 1983.

HUSS, Conrad, Nikita G. Reisler, and R. Mead Almond; “Practical and Economic Aspects of In-Pit Crushing Conveyor Systems”, SME/ AIME, October 1983.

HUSS, Conrad, and Dan Neff; “Finite Element Structural Analysis of Movable Crusher Supports”, Bulk Solids Handling, March 1983.

HUSS, Conrad; “Cost Considerations for In-Pit Crushing/Conveying Systems”, Bulk Solids Handling,

ALMOND, R. Mead, and Conrad Huss; “Open-Pit Crushing and Conveying Systems”, Engineering

Page 7 of 8 12-12

Conrad E. Huss, P.E.

Project Manager and Project Sponsor

M3 Tucson

MUNSELL, Stephen, R. Mead Almond, and Conrad Huss; “The Trend Toward Belt Conveying of Ore and Waste in Arizona Open Pit Mines”, SME/AIME, September 1978.

SANAN, Bal, and Conrad Huss; “Foundation Design for Rod and Ball Mills”, ACI Conference 1977, Presentation Only.

HUSS, Conrad, and Ralph Richard; “Dynamic Earthquake Analysis of Tucson Federal Office Building”, GSA Contract 70-6-02-0058, May 1972.

HUSS, Conrad; “Axisymmetric Shells Under Arbitrary Loading”, The University of Arizona, 1970. Doctoral Thesis.

HUSS, Conrad; “Airy’s Function by a Modified Trefftz’s Procedure” The University of Arizona, 1968. Master’s Thesis.

Page 8 of 8 12-12

CERTIFICATE of QUALIFIED PERSON

I, Thomas L. Drielick, P.E., do hereby certify that:

1.I am currently employed as Sr. Vice President by:

M3 Engineering & Technology Corporation

2051 W. Sunset Road, Ste. 101

Tucson, Arizona 85704

U.S.A.

2.I am a graduate of Michigan Technological University and received a Bachelor of Science degree in Metallurgical Engineering in 1970. I am also a graduate of Southern Illinois University and received an M.B.A. degree in 1973.

3.I am a:

·Registered Professional Engineer in the State of Arizona (No. 22958)

·Registered Professional Engineer in the State of Michigan (No. 6201055633)

·Member in good standing of the Society for Mining, Metallurgy and Exploration, Inc. (No. 850920)

4.I have practiced metallurgical and mineral processing engineering and project management for 44 years. I have worked for mining and exploration companies for 18 years and for M3 Engineering and Technology, Corporation for 26 years.

5.I have read the definition of “qualified person” set out in National instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.I am responsible for the preparation of Sections 13 and 17, as well as corresponding items of Sections 1, 25 and 26 of the technical report titled “Escobal Mine Guatemala NI 43-101 Feasibility Study,” dated 5 November 2014 (the “Technical Report”).

7.I have prior involvement with the property that is the subject of the Technical Report. I was responsible for Sections 13 and 17 of a previous Technical Report on the subject property entitled “Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment,” dated 24 July 2013.

8.As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information required to be disclosed to make the report not misleading.

9.I am independent of the issuer applying all of the tests in section 1.5 of National Instrument 43-101.

10.I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

11.I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

Dated this 5 November 2014.

/s/ Thomas L. Drielick

Signature of Qualified Person

Thomas L. Drielick

Print name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

I, Daniel Roth, P.E., do hereby certify that:

1.I am currently employed as a project manager and civil engineer at M3 Engineering & Technology Corporation located at 2051 West Sunset Road, Suite 101, Tucson, AZ, 85704.

2.I graduated with a Bachelor’s of Science degree in Civil Engineering from the University of Manitoba in 1990.

3.I am a registered professional engineer in good standing in the following jurisdictions:

·British Columbia, Canada (No. 38037)

·Alberta, Canada (No. 62310)

·Ontario, Canada (No. 100156213)

·Yukon, Canada (No. 1998)

·New Mexico, USA (No. 17342)

·Arizona, USA (No. 37319)

I am also a member in good standing with the Society of Mining, Metallurgy and Exploration.

4.I have practiced engineering and project management for 22 years. I joined M3 Engineering in November 2003.

5.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.I am responsible for Sections 2, 3, 4, 5, 6, 15, 16, 18, 19, 20, 21, 23, 24, 25, 26 and corresponding items of Section 1 of this technical report titled Escobal Mine Guatemala NI 43-101 Feasibility Study for Tahoe Resources Inc. dated November 5, 2014 (“Technical Report”).

7.I have prior involvement with the property that is the subject of the Technical Report. I was a contributing author of a previous Technical Report on the subject property entitled “Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment” dated 24 July 2013. M3 Engineering & Technology Corporation was the EPCM contractor for the Escobal Project. I visited the Escobal project site multiple times from 2010 through 2014.

8.As of the date of this certificate, 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.

9.I am independent of the Tahoe Resources Inc. and all their subsidiaries as defined in Section 1.5 of NI 43-101.

10.I have read National Instrument 43-101 and Form 43-101F1. The sections of the Technical Report that I am responsible for have been prepared in compliance with that instrument and form.

11.The Technical Report contains information relating to mineral titles, and legal agreements, and I do not offer a professional opinion regarding these issues.

12.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 accessible by the public, of the Technical Report.

Dated November 5, 2014.

/s/ Daniel Roth

Signature of Qualified Person

Daniel Roth

Print Name of Qualified Person

Daniel Roth, P.E.

Project Manager/Civil Engineer

M3 Tucson

Resume

Lead Civil Engineer Joined M3 in 2003 Education Bachelor of Science, Civil Engineering, University of Manitoba Registrations Civil Engineer — Arizona, New Mexico  Professional Engineer — Alberta, Canada; British Columbia; Ontario, Canada; Yukon

Twenty-three years of experience in mining, civil and environmental engineering work in Canada, United States, Mexico, Guatemala and Barbados. Extensive design and construction knowledge in hard rock mine site reclamation and water management. Responsibilities have included project management, design, and construction administration. PROJECT EXPERIENCE M3 ENGINEERING & TECHNOLOGY CORPORATION PROJECT MANAGER /CIVIL ENGINEER (10 YEARS) ·Tahoe Resources, Escobal Mine, 4,500 tpd Gold/Silver/Lead/Zinc Differential Flotation Circuit. PEA, Environmental Permitting Assistance, and EPCM services (Project Manager / Civil Engineer) — Guatemala ·Victoria Gold, Eagle Gold Mine, 29,000 tpd Heap Leach (Civil Engineer) — Yukon ·Allied Nevada, Hycroft Mine, 120,000 tpd Gold/Silver Flotation with Concentrate Oxidation. Pre-feasibility Study (Project Manager) - Nevada ·Coeur, Palmarejo Mine, Cyanide Detruct Water Treatment Plant, 600 Cubic Meters Per Hour (Project Manager) — Chihuahua, Mexico ·Freeport- McMoRan, Tyrone Mine Tailing Impoundment and Stockpile Reclamation (Project Manager / Chief Engineer) - New Mexico ·Rosemont Copper, Rosemont Mine Reclamation and Closure Plan (Civil Engineer) — Arizona ·Rio Tinto, Kennecott Eagle Mine Impacted Water Controls (Civil Engineer) — Michigan BROWN AND CALDWELL (1 YEAR) ·Pima County Wastewater Management Department, Northwest Outfall Sewer Rehabilitation Main Bypass - Arizona STANTEC CONSULTING (5 YEARS) ·Pima County Wastewater Management Department, Continental Ranch Sanitary Pump Station - Arizona ·Department of Emergency and Military Affairs, Fire Pump Station Rehabilitation - Arizona

Daniel Roth, P.E.

Project Manager/Civil Engineer

M3 Tucson

·Pima County Wastewater Management Department, Mobile Sludge Thickening Equipment /,Sludge management study and evaluation of mobile dewatering equipment - Arizona

·Yuma County Department of Development Services, Sanitary Sewer Lift Station and Force Main - Arizona

·Pima County Wastewater Management Department, Sanitary Sewer Capacity Evaluation - Arizona

·Pima County Wastewater Management Department, Avra Valley Wastewater Treatment Facility - Arizona

·Barbados West Coast Sewerage Project — West Indies

·Gila Indian Community, Pima-Maricopa Irrigation Project - Arizona

·Water Booster Pump Stations — Arizona

·Elbow Valley Sanitary Sewer System, Project management and resident engineer for seven contracts — Alberta, Canada

COCHRANE ENGINEERING, FORMERLY KNOWN AS POETKER MACLAREN LAVALIN (7 YEARS)

·City of Portage La Prairie, Wastewater Treatment Plant Upgrade and expansion (11MGD) — Manitoba, Canada

·Stony Mountain, Village of Stony Mountain Water/Sewer System — Manitoba, Canada

·RM of Rockwood & Winchester, Wastewater Stabilization Ponds — Manitoba, Canada

·Taillieu Construction, Water/Sewer Pipe Layer

Courses

·Completed MSHA Training

·MSHA Refresher Course, 2012

Page 2 of 2
04-14

CERTIFICATE OF QUALIFIED PERSON

I, Paul Tietz,C.P.G., do hereby certify that I am currently employed as Senior Geologist for Mine Development Associates, Inc. located at 210 South Rock Blvd., Reno, Nevada 89502 and:

1.I graduated with a Bachelor of Science degree in Biology/Geology from the University of Rochester in 1977, a Master of Science degree in Geology from the University of North Carolina, Chapel Hill in 1981, and a Master of Science degree in Geological Engineering from the University of Nevada, Reno in 2004.

2.I am a Certified Professional Geologist (#11004) with the American Institute of Professional Geologists.

3.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. I am independent of Tahoe Resources, Inc., applying all of the tests in section 1.5 of National Instrument 43-101.

4.I have practiced geology, project management, and resource estimation for 33 years. I joined Mine Development Associates in May 2006.

5.I am responsible for Sections 7, 8, 9, 10, 11, 12, and 14 and corresponding items of Sections 1, 24, 25 and 26 of this technical report titled Escobal Mine Guatemala NI 43-101 Feasibility Study for Tahoe Resources Inc. dated November 5, 2014 (“Technical Report”).

6.I was a co-author of a two previous Technical Reports on this property; the first entitled “Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment” and dated 29 November 2010, and the second entitled “Escobal Guatemala Project NI 43-101 Preliminary Economic Assessment” and dated 7 May 2012. I visited the Escobal project site on September 7th through the 10th, 2010, on February 6th through the 9th, 2012, and on January 28th through January 31st, 2014.

7.As of the date of this certificate, 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.

8.I have read National Instrument 43-101 and Form 43-101F1. The sections of the Technical Report that I am responsible for have been prepared in compliance with that instrument and form.

9.The Technical Report contains information relating to mineral titles, and legal agreements, and I do not offer a professional opinion regarding these issues.

10.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 accessible by the public, of the Technical Report.

Dated November 5, 2014.

/s/ Paul Tietz

Signature of Qualified Person

Paul Tietz

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

I, Matthew Blattman, P.E., do hereby certify that:

1.I am currently employed as Principal Engineer at Blattman Brothers Consulting, LLC located at 15201 Mason Road, Ste. 1000 #141, Cypress, TX 77433.

2.I graduated with a Bachelor’s of Science degree in Mining Engineering from the University of Nevada, Reno in 1996.

3.I am a registered professional engineer in good standing in the following jurisdictions:

·Nevada, USA (No. 015254)

I am also a Registered Member in good standing with the Society of Mining, Metallurgy and Exploration (No. RM4059667).

4.I have worked as a mining engineer continuously for 17 years since my graduation from university. I founded Blattman Brothers Consulting LLC in November 2010.

5.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.I am responsible for Sections 15 and 16 and corresponding items of Sections 1, 21, 22, 24, 25 and 26 of this technical report titled Escobal Mine Guatemala NI 43-101 Feasibility Study for Tahoe Resources Inc. dated November 5, 2014 (“Technical Report”).

7.I have prior involvement with the property that is the subject of the Technical Report. I visited the Escobal project site multiple times from 2012 through 2014.

8.As of the date of this certificate, 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.

9.I am independent of the Tahoe Resources Inc. and all their subsidiaries as defined in Section 1.5 of NI 43-101.

10.I have read National Instrument 43-101 and Form 43-101F1. The sections of the Technical Report that I am responsible for have been prepared in compliance with that instrument and form.

11.The Technical Report contains information relating to mineral titles, and legal agreements, and I do not offer a professional opinion regarding these issues.

12.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 accessible by the public, of the Technical Report.

Dated November 5, 2014.

/s/ Matthew Blattman

Signature of Qualified Person

Matthew Blattman

Print Name of Qualified Person

CERTIFICATE OF QUALIFIED PERSON

I, Jack Caldwell, P.E., do hereby certify that:

1.I am currently employed as a consultant (part time) at Robertson GeoConsultants located at 900-580 Hornby Street, Vancouver, B.C. Canada, V6C 3J6.

2.I graduated with a Master of Science (Engineering) degree in Civil Engineering from the University of the Witwatersrand, Johannesburg, South Africa in 1972.

3.I am a registered professional engineer in good standing in California, USA (C58841)

4.I have practiced engineering and project management for 42 years. I joined Robertson GeoConsultants in January of 2005.

5.I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.I am responsible for Section 18.6 and corresponding items of Sections 1, 25 and 26 of this technical report titled Escobal Mine Guatemala NI 43-101 Feasibility Study for Tahoe Resources Inc. dated November 5, 2014 (“Technical Report”).

7.I have prior involvement with the property that is the subject of the Technical Report. I visited the site in 2011 and subsequently and have been the lead of a team of engineers on the design and start of construction of the tailings facility. I continue to provide specialist consulting services on the operation of the tailings facility. I visited the Escobal project site on numerous occassions between 2011 and 2013.

8.As of the date of this certificate, 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.

9.I am independent of the Tahoe Resources Inc. and all their subsidiaries as defined in Section 1.5 of NI 43-101.

10.I have read National Instrument 43-101 and Form 43-101F1. The sections of the Technical Report that I am responsible for have been prepared in compliance with that instrument and form.

11.The Technical Report contains information relating to mineral titles, and legal agreements, and I do not offer a professional opinion regarding these issues.

12.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 accessible by the public, of the Technical Report.

Dated November 5, 2014.

/s/ Jack Caldwell

Signature of Qualified Person

Jack Caldwell

Print Name of Qualified Person