SilverCrest Mines (SVL) 6-K Silvercrest Mines Inc. 6-K

Exhibit 99.1

SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY

AND OPEN PIT RESERVE UPDATE

PREPARED FOR SILVERCREST MINES INC.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EFFECTIVE DATE: APRIL 30, 2013

RELEASED DATE: JULY 25, 2013

Contributors:

§   N. Eric Fier, C.P.G., P. Eng., President and Chief Operating Officer -  SilverCrest Mines Inc.

§   James Barr, P.Geo, Senior Geologist - EBA Engineering Consultants Ltd.

§   Mike Tansey, P. Eng., Senior Mining Engineer - EBA Engineering Consultants Ltd.

§   John Fox, P.Eng., Consulting Metallurgical Engineer - Laurion Consulting Inc.

§   Carlos Chaparro, P.Eng., Senior Geotechnical Engineer - EBA Engineering Consultants Ltd.

§   Nick Michael, P.Eng., Senior Mineral Economist - Tetra Tech.

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NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

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TABLE OF CONTENTS

1.0    EXECUTIVE SUMMARY	1
1.1   Introduction	1
1.2   Santa Elena Updated Mineral Resources and Mineral Reserves	1
1.2.1   Reserve Criteria	3
1.3   Mining Methods	4
1.4   Mining Schedule	4
1.4.1   Underground Production	5
1.5   Recovery Methods	6
1.6   Project Infrastructure	6
1.7   Capital and Operating Costs	7
1.8   Economic Analysis	9
1.9   Expansion Pre-Feasibility Recommendations	10
2.0   INTRODUCTION AND TERMS OF REFERENCE	11
2.1   Report Authors and EBA Quality Control	11
2.2   Site Visits	12
2.3   Effective Date of the Report	12
3.0   RELIANCE OF OTHER EXPERTS / DISCLAIMER	13
4.0   PROPERTY DESCRIPTION AND LOCATION	13
4.1   Mining Titles and Agreements	15
4.2   Royalties	17
4.3   Sandstorm Gold Stream	17
4.4   Environmental Liabilities and Permitting	17
5.0   ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCUTRE AND PHYSIOGRAPHY	18
5.1   Accessibility	18
5.2   Climate	18
5.3   Local Resources	18
5.3.1   Water Supply	18
5.3.2   Power	19
5.3.3   Community Services	19
5.4   Existing Infrastructure	20
5.5   Physiography	20
6.0   HISTORY	20
7.0   GEOLOGICAL SETTING	22
7.1   Regional Geological Setting	22
7.1.1   Local and Property Geology	23

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7.1.1.1   Structural Geology	25
8.0   DEPOSIT TYPE	26
8.1   Mineralization	26
9.0   EXPLORATION	27
10.0   DRILLING	27
10.1   Historical Drilling	27
10.2   Previous Drilling by SilverCrest	27
10.3   2012-2013 Drilling	28
10.3.1   Heap Leach Drilling	34
10.4   Core Recovery	35
10.5   Drill Surveying and Topography	35
11.0   SAMPLING METHODS AND APPROACH	36
11.1   Historical Sample Collection Methods	36
11.2   Previous SilverCrest Sample Collection Methods	36
11.3   2012-2013 Sample Collection Methods	36
11.4   2012-2013 Laboratory Analytical Methods	38
11.4.1   Nusantara Lab Preparation and Analysis	38
11.4.2   ALS-Chemex Lab Preparation and Analysis	38
11.5   Heap Leach Sampling, Preparation and Analysis	39
11.6   QP Statement	40
12.0   DATA VERIFICATION	40
12.1   Historical Data Verification	40
12.2   Previous SilverCrest Data Verification	40
12.3   2012-2013 Data Verification	41
12.3.1   Site Visits	41
12.3.2   SilverCrest Certified Reference Material Insertion	43
12.4   SilverCrest Blank Material Insertions	48
12.5   ALS-Chemex and Nusantara Laboratory Comparison	50
12.6   Heap Leach Data Verification	51
13.0   MINERAL PROCESSING AND METALLURGICAL TESTWORK	54
13.1   Previous Mineral Processing and Metallurgical test work	54
13.1.1   Previous Metallurgical Test Work Summary For Santa Elena	54
13.1.2   Previous Metallurgical Test Work Description	55
13.2   Recent Mineral Processing and Metallurgical test work	60
13.2.1   Leaching Cyanidation of Samples from Santa Elena Deposit	60
13.2.2   Filtering and Thickening Test work	71
14.0   MINERAL RESOURCE ESTIMATES	71
14.1   Introduction	71

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14.1.1   Previous Underground Resource Estimates	71
14.2   Basis of Current Resource Estimates	72
14.2.1   Gold Equivalent (AuEQ) Calculations and Assumptions	72
14.2.2   Specific Gravity used in the Resource Estimation	74
14.3   Geostatistics	76
14.3.1.1   Grade Compositing of Raw Samples	77
14.3.1.2   High Grade Capping	78
14.3.1.3   Cut-off grade calculation	79
14.4   Underground Mineral Resource Estimate	79
14.4.1   Drill Holes Used in Underground Mineral Resource	80
14.4.2   Grade Model Used in the Interpretation	82
14.4.3   Interpolation Method and Variography	85
14.4.4   Block Classification	86
14.4.5   Underground Mineral Resource Estimate	88
14.4.6   Underground Validation	91
14.4.7   EBA Statement on Mineral Resource Estimates	96
14.5   Heap Leach Mineral Resource Estimate	96
14.5.1   Heap Leach Resource Validation	96
14.6   Open Pit Mineral Resource Estimate	98
15.0   MINERAL RESERVES ESTIMATES	99
15.1   Introduction	99
15.2   Previous Reserve Estimates	99
15.3   Current Reserves Overview and Assumptions	100
15.4   Basis for Underground Mineral Reserve	101
15.4.1   Underground Mining Methodologies Consideration for Underground Reserves	106
15.5   Open Pit Reserves (SVL)	106
15.6   Heap leach Reserves (SVL)	107
16.0   MINING METHODS	108
16.1   Introduction	108
16.2   Geotechnical Analysis for Underground Mining	108
16.2.1   General
16.2.2   Geotechnical Data Collection	109
16.2.2.1   Drill Program	109
16.2.3   Rock Mass Characteristics	110
16.2.4   Design Parameters	110
16.2.4.1   Rock Mass parameters	110
16.2.4.2   Rock Strength	110
16.2.5   Design Approach	111
16.2.5.1   General	111
16.2.5.2   Design Assumptions	111
16.2.5.3   Tabulation of Results	111
16.2.5.4   Methodology of Stability Graph for Stope Design	112

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16.2.6   Span Length Design – Longitudinal and Transverse	113
16.2.7   Stope Back Width Design	113
16.2.8   Cut and Fill	113
16.2.9   Transverse Long Hole Stoping	114
16.2.10   Pillar Design	114
16.2.10.1   Sill Pillar Design	114
16.2.10.2   Crown Pillar Design	115
16.2.10.3   Rib Pillar Design	115
16.3   Proposed Underground Mining Methods	116
16.3.1   Longitudinal Long Hole Stoping	117
16.3.2   Transverse Long Hole Stoping	119
16.3.3   Mechanized Cut and Fill	121
16.3.4   Rib and Sill Pillar Recovery	123
16.3.5   Mining Equipment	123
16.4   Mine Design	123
16.4.1   Pillar Layout	123
16.4.2   Stope Layout	124
16.4.3   Development Layout	127
16.5   Backfill
16.5.1   Backfilling Cut and Fill Stopes	130
16.5.2   Backfilling Transverse Long Hole Stopes	130
16.5.3   Backfilling to Extract Pillars	130
16.6   Overall Mining Schedule	130
16.6.1   Underground Mine Production Schedule	131
16.6.2   Open Pit Mining Schedule	134
16.6.3   Heap Leach Mining Schedule	134
16.7   Ventilation	134
16.7.1   Ventilation Requirements	134
16.7.2   Ventilation Model	135
16.7.2.1   Conceptual ventilation design	135
16.7.2.2   Intake airway / Escape way Raise	135
16.7.2.3   Fresh Air Raise Head Without separate man way	137
16.7.2.4   Raised Bore for main exhaust ventilation airway	138
16.7.2.5   Vent Tubing Head For Tunnels and Ore drifts	138
16.7.2.6   Setup of ventilation model	139
16.7.3   Interpretation of Ventilation Conceptual Design	144
16.8   Open Pit Mining Method	144
16.9   Spent Ore Mining And Rehandling Methods	145
16.10   Tailings Disposal	145
17.0   RECOVERY METHODS	146
18.0   PROJECT INFRASTRUCTURE	150
18.1   Surface Infrastructure for the Expansion Plan	152

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18.1.1   Access Roads	152
18.1.2   Offices, Cafeteria, Warehouse and Change Houses	152
18.1.3   Fuel Storage Facility	152
18.1.4   Water storage facility	153
18.1.5   Electrical Distribution	153
18.1.6   Explosive Magazine	153
18.1.7   Maintenance Shop	153
18.1.8   On-site Laboratory	153
18.1.9   Cement Silo	153
18.2   Underground Infrastructure	153
18.2.1   Underground Maintenance Shop/Warehouse	154
18.2.2   Refuge Station	154
18.2.3   Dewatering	154
18.2.4   Underground Electrical Installations	156
18.2.5   Water Services	158
18.2.6   Powder Storage	158
18.2.7   Compressed Air	158
18.3   Processing Plant Infrastructure	158
18.4   Waste Rock and Tailings Storage Facilities	159
19.0   MARKET STUDIES AND CONTRACT	160
19.1   Sales of gold and silver doré	160
20.0   ENVIRONMANTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT	160
20.1   Land Use Change Permit	160
20.1.1   Summary of ARD/ML Work for Santa Elena	160
20.2   Closure Planning	161
21.0   CAPITAL AND OPERATING COSTS	162
21.1   Capital Costs	163
21.1.1   Mining Capital Costs	163
21.1.2   Underground Development	164
21.1.3   Underground Electrical Installations	165
21.1.4   Construction Consumables and Labour	166
21.1.5   Processing Capital Costs	167
21.1.6   Indirect Capital Costs	168
21.1.7   Sunk Costs	168
21.2   Operating Costs	168
21.2.1   Underground Mining Cost Estimate	169
21.2.2   Costs for Loading and Transporting Existing Heap Leach Material to the Mill for Reprocessing	171
21.2.3   Processing Plant Operating Costs	173
21.3   Tailings Handling Costs	174
21.4   General and Administrative Operating Costs	174

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22.0   ECONOMIC ANALYSIS	175
22.1   Introduction	175
22.3   Summary of Financial Results	176
22.4   Cash flows	177
22.5   Post - tax Results and Sensitivity Analysis	178
22.6   Purchase Agreement	182
22.7   Smelter Terms	182
23.0   ADJACENT PROPERTIES	182
24.0   OTHER RELEVANT DATA AND INFORMATION	182
24.1   Project Execution Plan – Progress Completed	182
24.2   Schedule to the Start of Underground Stoping	183
25.0   INTERPRETATION AND CONCLUSION	185
25.1   Geology and Underground Resource Estimate	185
25.2   Underground Reserves and Mining Methods	185
25.3   Costs and Project Economics	185
26.0   RECOMMENDATIONS	185
26.1   Geology and Underground Resource Estimate	185
26.1.1   Risks and Opportunities	185
26.2   Geotechnical Recommendation	186
26.2.1   Stope Size and Pillar Dimensioning	186
26.2.1.1   Hanging wall  and orebody NGI-Q values	186
26.2.1.2   Joint set orientation	186
26.2.1.3   In-situ stress regime	187
26.2.1.4   Hanging wall  and ore body strength and deformation characteristics	187
26.2.2   Geotechnical Considerations in Overall Mine Design	187
26.3   Recommendations on Mining Methods and Mining Costs	188
26.4   Environmental, Closure and Asset Retirement	188
REFERENCES	188

TABLES

Table 1.1: Update Mineral Reserve and Resource Estimates (April 30, 2013)	1
Table 1.2: Previous Reserve and Resource Estimates (January 2012) for comparative purposes	2
Table 1.3: Santa Elena Reserve Estimation Criteria	3
Table 1.4: Mining Method Selection Criteria	4
Table 1.5: Summary of Mining Schedule	5
Table 1.6: Capital Cost Summary	8
Table 1.7: Operating Costs	8
Table 1.8: Base Case Economic Analysis Results	9

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Table 1.9: Comparison between the Base Case and the Spot Price for the Pre-Tax Financial Results	10
Table 2.1: Independent QP’s in this study and date of their visit to the Santa Elena operations.	12
Table 4.1: List of Mining Titles	15
Table 4.2: Permits for Santa Elena	18
Table 10.1: Summary of 2006-2009 Drill Holes	28
Table 10.2: Summary of all 2012-2013 Santa Elena Drilling.	29
Table 10.3: Significant Drill Hole Intercepts from 2012-2013 Drilling	31
Table 10.4: 2012 Trench and Bulk Composite Samples	34
Table 11.1: Table showing RC Cutting Sample Length (From ‘SERC’- series drill holes) and frequency from the hanging wall	36
Table 11.2: Number and Preparation Method of Samples sent to ALS-Chemex from Santa Elena Mine	38
Table 12.1: Drill Core Grade Verification Sampling (Oct. 2012)	42
Table 12.2: Certified Reference Material Reporting Values used in the Santa Elena 2012-2013 Drilling by SilverCrest	44
Table 12.3: Key Statistic Comparison of Au and Ag for Nusantara and ALS-Chemex.	51
Table 12.4: Heap Leach Blank Samples Results	52
Table 12.5: Tabulated CRM Insertion Results for the Heap Leach Drilling	54
Table 13.1: Previous Metallurgical Reports and Comments	55
Table 13.2: Santa Elena Bottle Roll Results	56
Table 13.3: Santa Elena Bottle Roll Results for Ground Composites	56
Table 13.4: Santa Elena Bottle Roll Results for Coarse Composites	57
Table 13.5: Santa Elena Column Test Results for Particle Size Composites	58
Table 13.6: Santa Elena Grinding/Leaching Test Work	59
Table 13.7: Crushing Work Indices	59
Table 13.8: Composites Work Indices, Specific Gravities and Bulk Densities	60
Table 13.9: Baseline Cyanidation Performance	61
Table 13.10: Leach Optimization Test Summary	64
Table 13.11: Confirmatory Test Comp. 4 +/- 0 Nivel	66
Table 13.12: Average Recovery Rates for Underground Bulk Samples	68
Table 13.13: Results of Additional Testing on Samples 4, 6 and 8.	68
Table 13.14: Bulk Leach Summary	70
Table 14.1: Previous Santa Elena Underground Resources (EBA, May 2011)	72
Table 14.2: Metal Prices and Gold to Silver Ratio Estimates	74
Table 14.3: SG Tally Rock Type, Rock Codes used in SG Analysis	74
Table 14.4: Santa Elena SG Analysis and Average SG per Rockcode	75
Table 14.5: Descriptive Statistics for Raw Samples within Vein	76
Table 14.6: Descriptive Statistics for 1 metre Composite Samples within Vein	77
Table 14.7: Block Model Origin and Dimensions of the Santa Elena Model	80
Table 14.8: Drill Holes (all years) used for 2013 Resource Calculations.	80
Table 14.9: Description of Mineralized Domains for Santa Elena	83
Table 14.10: Ellipse and Variogram Parameters based on Experimental Variogram Analysis for the Santa Elena Model	86

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Table 14.11: Updated Underground Mineral Resource Estimate including Reserves for Santa Elena (Effective April 30th, 2013)	89
Table 14.12: Updated Underground Mineral Resource Estimate Excluding Reserves for Santa Elena (Effective April 30th, 2013)	90
Table 14.13: Validation of Heap Leach Resource Estimate	96
Table 14.14: Leach Pad Block Model Origin and Dimensions	98
Table 14.15: Leach Pad Material Mineral Resource Estimate (Effective April 30th, 2013)	98
Table 15.1: Previous Santa Elena Open Pit Reserves (January 2011)	99
Table 15.2: 2013 Santa Elena Mineral Reserve Estimates	100
Table 15.3: Breakdown Summary of Santa Elena Underground Reserves	103
Table 15.4: Santa Elena indicated Stope Tonnage & Grade	104
Table 15.5: Mining Methods % Utilization throughout LOM	106
Table 15.6: Santa Elena Open Pit Mineral Reserves (Effective December 31, 2012)	107
Table 15.7: Heap Leach Material Mineral Reserve Estimate (Effective April 30th, 2013)	107
Table 16.1: 2012 Borehole Summary	109
Table 16.2: RMR’ and NGI-Q’ Values for Design – at 50%	110
Table 16.3: NGI-Q' Values for Stope Design	112
Table 16.4: Transverse Stope Design	114
Table 16.5: Sill Pillar Stability	115
Table 16.6: Rib Pillar Stability	116
Table 16.7: Mining Method Selection Criteria	116
Table 16.8: Mining Equipment List	123
Table 16.9: Linear Meters of development over LOM	127
Table 16.10: Summary of Overall Production Schedule	131
Table 16.11: Scheduling Parameters	131
Table 16.12: Underground Mining Schedule	132
Table 16.13: Equipment Power and Utilization	134
Table 16.14: Design Parameters for Escape Way Raise	137
Table 16.15: Design Parameters for Fresh Air Raise	137
Table 16.16: Design Parameters for Exhaust Raise	138
Table 16.17: Design Parameters for Vent Tubing	139
Table 16.18: Dimensions and Friction Factors for Underground Workings	140
Table 18.1: Diesel Consumption (Litres) per day	152
Table 18.2: Major Equipment Components as part of the Processing Plant	158
Table 18.3: Waste Rock Volumes	159
Table 21.1: Capital Cost Summary	163
Table 21.2: Equipment Capital Expenditure	164
Table 21.3: Development Expenditure	165
Table 21.4: Underground Electrical and Service Cost	166
Table 21.5: Construction Consumables and Labour	167
Table 21.6: Processing Capital Expenditure	168
Table 21.7: Operating Costs	169
Table 21.8: Long Hole Mining Cost Summary	170

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Table 21.9: Cut and Fill Cost Summary	171
Table 21.10: Reprocess Material Rehandling Cost	172
Table 21.11: Grasshopper Conveyor Costs	173
Table 21.12: Material Handling Cost from Heap Leach to Mill	173
Table 21.13: Summary of Processing Costs	174
Table 22.1: Technical-Economic Modeling Assumptions	175
Table 22.2: Base Case Economic Analysis Results	176
Table 22.3: Comparison of Base Case Post Tax Results used in the PFS with Spot Price Post Tax Results at mid July 2013	177
Table 22.4: Discounted Post Tax Cash Flow Model	180
Table 24.1: Process Plant Equipment Purchase Progress	183

FIGURES

Figure 1.1: Summary of annual tonnes and grade for the life of mine	6
Figure 4.1: Santa Elena Property Location	14
Figure 4.2: Santa Elena Property Map	16
Figure 7.1: Santa Elena Property Geology Map	24
Figure 10.1: Map showing 2012-2013 Santa Elena Drilling	30
Figure 10.2: Drilling setup on the Heap Leach Pad	34
Figure 10.3: Core Recovery and RQD of Santa Elena 2012-2013 Drilling	35
Figure 11.1: Santa Elena Sample Length Frequency used in 2013 Resource Estimate	37
Figure 11.2: Heap Leach Sampling	40
Figure 12.1: Grade Verification Sampling, Silver	42
Figure 12.2: Grade Verification Sampling, Gold	43
Figure 12.3: Classification of Santa Elena CRM’s Material	44
Figure 12.4: CRM: CDN-CM-17 - Gold	45
Figure 12.5: CRM: CDN-CM-17 – Silver	45
Figure 12.6: CRM: CDN-ME-19 – Gold	46
Figure 12.7: CRM: CDN-ME-19 - Silver	46
Figure 12.8: CRM: CDN-ME-5 – Gold	47
Figure 12.9: CRM: CDN-ME-5 - Silver	47
Figure 12.10: CRM: OxD87 - Gold	48
Figure 12.11: SilverCrest 2012-2013 Reported Blank Material Au (gpt)	49
Figure 12.12: SilverCrest 2012-2013 Reported Blank Material Ag (gpt)	49
Figure 12.13: Au (gpt) Comparison between Santa Elena and ALS-Chemex	50
Figure 12.14: Ag (gpt) Comparison between Santa Elena and ALS-Chemex	51
Figure 12.15: Distribution of CRM Au Grades from Leach Pad Drilling Insertions	53
Figure 12.16: Distribution of CRM Ag Grades from Leach Pad Drilling Insertions	53
Figure 13.1: Gold Leach Kinetics on Milled NIVEL 4 Comp.	62
Figure 13.2: Silver Leach Kinetics on Milled NIVEL 4 Comp.	63
Figure 13.3: Optimization Gold Extraction Kinetics	64
Figure 13.4: Optimization Silver Extraction Kinetics	65

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	66
	67
Figure 14.1: 3 Year Trailing Average Monthly Gold Price with 3 and 5 Year Composite Averages	73
Figure 14.2: 3 Year Trailing Average Monthly Silver Price with 3 and 5 Year Composite Averages	73
Figure 14.3: Santa Elena Specific Gravity Mean Value and ±1SD Ranges	75
Figure 14.4: Log-histogram of Raw Gold Grades within Santa Elena Vein	76
Figure 14.5: Log-histogram of Raw Silver Grades within Santa Elena Vein	77
Figure 14.6: Cumulative Log-probability Graph for Composited Gold Grades	78
Figure 14.7: Cumulative Log-probability Graph for Composited Silver Grades	79
Figure 14.8: Location of Drill Holes used in 2013 Mineral Resource Update	81
Figure 14.9: Geological Model showing Vein Domains, looking north	84
Figure 14.10: Geological Model showing Vein Domains, looking ESE	85
Figure 14.11: Factor applied to Total Sill of the Major Experimental Variogram Range	88
Figure 14.12: Oblique Perspective of Santa Elena Block Model, Indicated Resource (AuEq >1.4 gpt)	91
Figure 14.13: Vertical Section Comparison of Model Runs and input Au grades	92
Figure 14.14: Vertical Section Comparison of Model Runs and input Ag Grades	93
Figure 14.15: Plan Section Comparison of Model Runs and input Au Grades	94
Figure 14.16: Vertical Section Comparison of Model Runs and input Ag Grades	95
Figure 14.17: Leach Pad Verification Drill Hole Location Map	97
Figure 15.1: Stope Shapes defined for use in Reserve Definition	102
Figure 16.1: General Sublevel Stoping Operation	118
Figure 16.2: General Transverse Long Hole Stoping Method	120
Figure 16.3: General Mechanized Cut and Fill Stoping Method	122
Figure 16.4: Longitudinal View of Stopes, Looking North	125
Figure 16.5: 3D View of Stope Layout, Looking Northwest	126
Figure 16.6: Development and Stope Layout, Looking North	128
Figure 16.7: 3D View of Stope and Development Layout, Looking Northwest	129
Figure 16.8: Summary of Annual Tonnes by Underground Mining Methods and Grade for the Life of Mine	133
Figure 16.9: Escape Way Raise	136
Figure 16.10: Phase 1 Ventilation Air Flow Diagram	141
Figure 16.11: Phase 2 Ventilation Air Flow Diagram	142
Figure 16.12: Phase 3 Ventilation Air Flow Diagram	143
Figure 17.1: Santa Elena Expansion 3000 tpd Process Flowsheet	148
Figure 17.2: Santa Elena Expansion Plan Plant Mechanical General Arrangement	149
Figure 18.1: Site General Arrangement Plan	151
Figure 18.2: Conceptual Water Balance	155
Figure 18.3: Conceptual U/G Electrical Distribution Diagram	157
Figure 22.1: Pre-tax and Post-tax Cash Flows	178
Figure 22.2: NPV Sensitivity Analysis	181
Figure 22.3: RR Sensitivity Analysis	181
Figure 24.1: Santa Elena Expansion Project Development Schedule	184

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APPENDICES

Appendix A	Qualified Person Certificates
Appendix B	Certified Reference Material Data Sheets
Appendix C	Metallurgical Investigation, Summary Reports
Appendix D	2012-2013 Drill Hole Collar Locations
Appendix E	Geotechnical Data
Appendix F	Economic Analysis

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ACRONYMS & ABBREVIATIONS

AA	Atomic Absorption
Au	Gold
Au EQ	Gold equivalent
BC	British Columbia
CIM	Canadian Institute for Mining, Metallurgy and Petroleum
CNCF	Cumulative Net Cash Flow
CONAGUA	Comision Nacional Del Agua
CRM	Certified Reference Material
DCF	Discounted Cash Flow
DDH	Diamond Drill Hole
DGGFS	Dirección General de Gestión Forestal y de Suelos
DTM	Digital terrain model
EBA	EBA Engineering Consultants Ltd. operating as EBA, A Tetra Tech Company
FA	Fire Assay
GRA	Gravimetric Determination
ICP-AES	Inductively coupled plasma-atomic emission spectrometry
IDW	Inverse Distance Weighted
INAH	Nation Institute of Archaeology and History
INEGI	Instituto Nacional de Geografía y Estadística de México
LOM	Life of Mine
NCF	Net Cash Flow
NI 43-101	National Instrument 43-101 Standards of Disclosure for Mineral Projects
NGI	Norwegian Geotechnical Institute
NN	Nearest Neighbor
NSR	Net Smelter Returns
NPV	Net present Value
OK	Ordinary Kriging
P.Eng	Professional Engineer
PFS	Pre-Feasibility Study
P.Geo	Professional Geologist
QA-QC	Quality Assurance and Quality Control
QP	Qualified Person
RC	Reverse Circulation
RMR	Rock Mass Rating
RQD	Rock Quality Designation
S.A. de C.V.	Sociedad Anónima de Capital Variable
SCdM	SilverCrest de Mexico S.A. de C.V.
SD	Standard Deviation
SEDENA	Mexican Secretariat of National Defense
SEMARNAT	Secretaría de Medio Ambiente y Recursos Naturales
SWRPA	Scott Wilson Roscoe Postle Associates

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UBC	University of British Columbia
UG	Underground
UTM	Universal Transverse Mercator
WGS	World Geodetic System
XRD	X-ray diffraction analysis

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UNITS OF MEASUREMENT AND CONVERSIONS

$USD	United States Dollar
%	Percent
°C	Degree Celsius
Bhp	Brake Horse Power
cu.m/s	Cubic Metre per Second
deg	Degrees
gal	Gallon
gpt	Grams per tonne
ha	Hectare (10,000 square meters)
HP	Horsepower
hrs	Hours
kg	Kilogram
km	Kilometer
kV	Kilovolt
kW	Kilowatt
kWhr	Kilowatt Hour
Lps	Liters per second
M	Million
M	Metre
m/s	Metre per second
m3/s	Metre cubed per second
Ma	Million Years
mm	Milimeter
MxnPeso$	Mexican Pesos
MW	MegaWattPascal
Pa	Pascal
oz	Ounce (troy)
ppb	Part per billion
ppm	Parts per million
Sec	Second
V	Volt
AuEQ (gpt)
% Difference
Contained oz
Contained lbs

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NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL 30, 2013 | RELEASED DATE: JULY 25, 2013

1.0   EXECUTIVE SUMMARY

1.1   Introduction

EBA Engineering Consultants Ltd. operating as EBA, A Tetra Tech Company (EBA) has prepared this Pre-Feasibility Study (PFS) for SilverCrest Mines Inc. (SilverCrest, SVL) of Vancouver, British Columbia, Canada for the Santa Elena Mine, located in central Sonora, Mexico.  Nusantara de Mexico S.A. de C.V., a 100% owned Mexican subsidiary and a legal operating entity of SilverCrest, holds the rights to the Santa Elena Mine.  The effective date of this report is April 30, 2013.

The Santa Elena Mine is currently producing gold and silver from a 2,500 tonne per day open pit and heap leaching operation, which declared commercial production in July of 2011.  The Santa Elena Expansion Project involves combined processing of ore from the remaining reserves in the open pit, new reserves from underground development and reprocessing of spent ore from the existing heap leach pad in a new milling and processing plant.  Construction of a 3,000 tonne per day mill and plant facility is currently underway on the property. An underground exploration decline is being developed and has intersected the main mineralized zone (MMZ) at an elevation of 625 metres above sea level.

The purpose of this report is to document a PFS completed for the Santa Elena Expansion Project and support recent Mineral Resource and Mineral Reserve Estimates announced on May 29, 2013.

This report conforms to National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101 and Form 43-101F1), and incorporates the Canadian Institute for Mining, Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves (CIM Definition Standards).

1.2   Santa Elena Updated Mineral Resources and Mineral Reserves

The updated Probable Reserves (underground, open pit and leach pad) are estimated at 8.2 million tonnes grading 74.9 gpt Ag and 1.24 gpt Au, containing 19.7 million ounces of silver and 327,430 ounces of gold. This represents a 103% increase in contained silver and 50% increase in contained gold over previous Probable Reserves (Table 1.2). Updated Indicated Resources (exclusive of Probable Reserves) are estimated at 2.1 million tonnes grading 114.9 gpt Ag and 1.69 gpt Au, containing 7.9 million ounce of silver and 116,000 ounces of gold. This represents a 127% increase in contained silver and 99% increase in contained gold over previous Indicated Resources.

Table 1.1: Update Mineral Reserve and Resource Estimates (April 30, 2013)

Classification1	Tonnes	Au Gpt	Ag Gpt	Au Oz	Ag Oz
Santa Elena Underground Diluted and Recoverable Reserves2
Probable	3,920,510	1.57	108.1	198,170	13,624,640
Santa Elena Open Pit Reserves3
Probable	1,426,710	1.52	66.8	69,830	3,062,200
Santa Elena Leach Pad Reserves4
Probable	2,844,530	0.65	33.3	59,420	3,048,200
Total Reserves	8,191,760	1.24	74.9	327,430	19,735,050
Santa Elena Underground Resources5
Indicated	2,142,820	1.69	114.9	116,000	7,919,000
Inferred	1,489,750	1.50	155.6	72,000	7,453,000

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Note: All numbers are rounded. Underground and Leach Pad Reserves and Resources are based on 3 year historic metal price trends of US$28/oz silver, US$1450/oz gold and metallurgical recoveries of 92% Au and 67.5% Ag with a metal ratio of Ag:Au at 70:1 used for grade cut-off determination . All Mineral Resources and Reserves conform to NI 43-101 and CIM definitions for Resources and Reserves. Inferred Resources have been estimated from geological evidence and limited sampling and must be treated with a lower level of confidence than Indicated Resources.

1 Open Pit and Leach Pad Probable Reserves were classified by SilverCrest.  Underground Reserves and Resources were classified by EBA, a Tetra Tech Company.

2 Underground Probable Reserve is based on a cut-off grade of 1.47 gpt AuEq with an average 10% dilution and 90% mine recovery.  Average true thickness of the designed stopes is 13.4 metres.

3 Open Pit Reserve is based on a cut-off grade of 0.20 gpt AuEq in a constrained pit shell with applied capping of 8 gpt Au and 300 gpt Ag.

4 Leach Pad Reserve based on production and drill hole data for volumetrics and grade model using a cut-off grade of 0.5 gpt AuEq. No capping was applied.

5 Underground Resources are exclusive of Probable Reserves and based on 1 gpt AuEq grade shell, a cut-off grade of 1.4 gpt AuEq, and applied capping of 12 gpt Au and 600 gpt Ag.

Table 1.2: Previous Reserve and Resource Estimates (January 2012) for comparative purposes

Classification	Tonnes	Au gpt	Ag gpt	Au oz	Ag oz
Santa Elena Open Pit Reserves1
Probable	3,459,200	1.96	87.3	217,800	9,713,600
Santa Elena Underground Resources2
Indicated	991,100	1.83	109.1	58,330	3,476,960
Inferred	1,879,000	1.53	86.9	92,470	5,250,190

Note: All numbers are rounded.

1 based on $1,000/oz of gold and $18/oz of silver, cut-off grade of 0.38 gpt gold equivalent with applied metallurgical recoveries

2 based on $1,000/oz of gold and $18/oz of silver, cut-off grade is 1.77 gpt gold equivalent with applied metallurgical recoveries.

The updated underground Reserves and Resources have been independently estimated by EBA utilizing 2012 and 2013 drilling results along with previous independently-validated data (see SEDAR for previous NI 43-101 compliant Technical Reports). Drilling in 2012 and early 2013 was focused on conversion and expansion of underground Resources to Reserves with an average drill hole spacing of approximately 35 metres. Drilling data from 2012 and 2013 used for underground Resource estimate included 135 holes (48,980 metres) and 96 previous holes (14,865 metres). A total of 2,676 composite samples were used as the basis for the block model.

The updated open pit Reserves and Resources have been estimated by SilverCrest utilizing extensive production data (blastholes), exploration drilling data and computer modelling. All Indicated Resources in the currently operating open pit have been converted to Reserves as of April 30, 2013.

The Leach Pad Reserves (spent ore) have been estimated by SilverCrest utilizing crusher composite samples, drill data from verification holes on the pad, bottle roll tests to process design specifications and SilverCrest production leach curves. In early May 2013, the Company completed 25 large diameter drill holes (355 metres) on the pad to obtain additional information for the purposes of validation of production data. All spent ore material on the leach pad will be reprocessed through the new processing facility once operational. Only the material currently on the pad and leached (300 day cycle completed or estimated full cycle) has been declared as Mineral Reserves. Approximately 750,000 tonnes of open pit material is planned to be placed on the pad during the remainder of 2013 and will undergo only partial leaching before being reprocessed through the new facility. Once this material is placed on the pad and leaching has been discontinued, it will be declared as leach pad Reserves for processing and metal recovery in late 2014 and beyond.

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Before production commenced, the Santa Elena open pit defined a mine life of approximately 6.5 years (including ramp up) having a Reserve of 5.1 million tonnes grading 1.72 gpt Au and 73.4 gpt Ag (refer to Technical Report dated April 1, 2011). In April 2011, the open pit Reserve was depleted by approximately 6 months of production and optimized with the remaining 4.79 million tonnes grading 1.81 gpt Au and 75.9 gpt Ag (reported in Technical Report dated April 1, 2011).  By January 2012, the open pit was depleted from 2011 production and re-optimized with a resultant Reserve of 3.45 million tonnes grading 1.96 gpt Au and 87.3 gpt Ag showing lesser tonnes and higher grade. Re-optimization of the open pit between April 2011 and April 2013 has reclassified approximately 880,000 tonnes of open pit Reserve to underground Resources which was converted mostly (excluding stope pillars) to the new underground Reserves. The main purpose for re-optimization was to avoid a higher open pit strip ratio at higher cost in future years in the mine life but still be able to mine the displaced Reserves from underground.

Extensive metallurgical test work including ongoing operations data show that all declared Reserves are amenable to conventional leaching either by heap leach technology or standard CCD milling with a Merrill Crowe recovery system for doré bar production. Estimated recovery grades are stated in Table 1.3 below, and in Section 1.5.

1.2.1   Reserve Criteria

The updated Santa Elena Mineral Reserves as stated above were completed using GEMS resource models for defining open pit, underground and leach pad Reserves along with criteria as presented in the following table.

Table 1.3: Santa Elena Reserve Estimation Criteria

Base Case Metal Prices1	All US$
Gold	$1,450.00
Silver	$28.00
Mining Method	Open Pit	Open Pit	Underground	Leach Pad
Process Method	Heap Leach	CCD Mill	CCD Mill	CCD Mill
Mining Cost/T ore	$2.42	$2.50	$21 to 482	$0.25 to 0.75
Mining Cost/T waste	$1.52	$2.00	NA	NA
Processing Cost/T3	$6.65	$20.50	$20.50	$18.50
General & Administration4	$3.30	$3.00	$3.00	$3.00
Overall Metal Recoveries (Life of Pad)5
Gold	63%	92%	92%	92%
Silver	40%	67.5%	67.5%	67.5%
Pit walls – North Wall overall angle	55.0	55.0	NA	NA
Pit walls – South Wall overall angle	45.0	45.0	NA	NA
Ramp Width (metres)	10 to 15	10 to 15	4.5	NA
Average Dilution	5%	5%	10%2	NA
Ave. Mining Recovery	95%	95%	90%2	100%

1 Based on 3 year historic metal price trends

2 Underground mining costs, dilution and mine recovery are based on stope type, either long hole (69% of design stopes) or cut and fill (31% of designed stopes) mining method.

3 Processing includes crushing, milling, site refining and dry stack tailings disposal.

4 Estimated based on current operations and may vary on an annual basis.

5 Recoveries for leach pad based on 2009 Metcon test work and recent Company production leach cycles for life of pad. During the period from 2010 through 2012 (solution ramp up), an average recovery of 57% Au and 30% Ag was achieved. Refer to Technical Report dated May 11, 2011 on SEDAR. Recoveries reflect partial 300 day leach cycle with pad leaching discontinued prematurely in Q1 2014. Leach pad CCD mill recoveries are based on in-situ remaining ounces on the pad.

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1.3   Mining Methods

The Santa Elena ore body varies in dip and thickness along strike and at depth. As a result, three well established underground mining methods have been selected for ore extraction. These mining methods are categorized in Table 1.4 below:

Table 1.4: Mining Method Selection Criteria

Orebody Geometry	Mining Method
Dip > 50 degrees, Thickness < 15m	Longitudinal Long hole Stoping
Dip > 50 degrees, Thickness > 15m	Transverse Long hole Stoping
Dip < 50 Degrees, Any Thickness	Mechanized Cut and Fill

In general, conventional mechanised mining methods have been selected.  The basis of the development of the mining methods and consequent equipment selection has been that SilverCrest will undertake production drilling, blasting and loading using a contractor for the waste rock and ore haulage to surface.  Initially a contractor will be retained to carry out mine development, with jumbo drill rigs purchased later in the mining life, after which development will be done in house. Approximately 69% of stoping will be by long hole method and 31% by cut and fill methods. Most long hole stopes are produced early in the mine schedule. Average stope width is 13.4 metres.

Conventional open pit mining will continue using a contractor until the third quarter of 2014 when open pit reserves are depleted. Mining of the heap leach spent ore will be completed by loader and conveyor to transport material to the plant.

1.4   Mining Schedule

The mining schedule estimates the tonnages to be mined from the underground, open pit and the existing heap leach facility to feed the expansion plan process plant at a nominal rate of 3,000 tpd.  Table 1.5 shows the combined schedule for the Santa Elena Expansion Project. The schedule is based on optimizing higher grade long hole stopes first, with more costly cut and fill mining left for later in the mine life.

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Table 1.5: Summary of Mining Schedule

Aspect of operations	2013	2014	2015	2016	2017	2018	2019	2020	2021	Total life of mine
Total tonnes from underground		127,707	392,412	535,520	646,088	708,359	684,604	580,407	245,225	3,920,323
Total tonnes from old heap leach		335,426	615,588	472,480	361,912	299,641	323,396	427,593	8,495	2,844,530
Total tonnes from open pit	882,040*	544,867	-	-	-	-	-	-	-	1,426,907
Total tonnes processed	882,040*	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	253,720	8,191,760
Total AU ounces sold		36,173	39,606	29,361	27,557	27,924	53,357	37,535	11,226	262,739
Total AG ounces sold		1,345,248	1,594,643	1,409,639	1,742,741	1,556,867	1,951,279	1,816,118	702,392	12,118,926

* 2013 open production has been included in the reserves but has not been included in the financial analysis for the Santa Elena expansion plan

Table 1-5 excludes approximately 882 kilo tonnes yet to be mined in the open pit during 2013, which will be placed on the heap leach pad.  By early 2014, these tonnes and estimated remaining ounces will be added to the schedule for additional mine life.

1.4.1   Underground Production

For the purpose of this PFS, an underground mining schedule has been developed for the stopes in the reserve model and for development required to access the stopes throughout the life of mine. The mining schedule results in grade and tonnage performance as shown in Figure 1.1. Peak production is reached in year 6, when both access ramps are actively producing.

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Figure 1.1: Summary of annual tonnes and grade for the life of mine

 

1.5   Recovery Methods

The ore from both underground and open pit resources will be processed by conventional cyanide leaching technology. In addition partially leached material from the existing heap leach operations will be blended with underground ore at a variable rate and reprocessed through the same plant.

Santa Elena ore (Open Pit, Underground and Leach Pad) contains an estimated 1.24 g/t Au and 75 g/t Ag and after crushing and grinding can be leached in cyanide to yield approximately 92% Au Recovery and 67.5% Ag recovery. Because of the relatively high level of silver in the ore (and hence solutions) there are advantages and benefits to using traditional CCD and Merrill-Crowe for metal recovery rather than CIL/CIP. The partially leached heap ore yields recoveries of approximately 52% Au and 29% Ag when crushed to 10 mm and processed on the heap leach (partial leach cycle to Q1 2014).  On re-leaching after grinding in the new plant, the balance of the metals are recovered to the level expected from new ore from open pit and underground.

The process plant has been designed to treat a nominal 3000 tonne per day (tpd) of ore, a mixture of freshly mined material and partially leached heap leach residue.  The plant has been designed to treat any proportion of these two types of feed.  Because this may include up to 100% fine crushed material from the heap leach pads it was considered that this would be unsuitable feed for a SAG milling circuit and it was elected to use a conventional crushing and grinding circuit.

1.6   Project Infrastructure

The Santa Elena open pit heap leach mine was constructed in late 2009 and 2010, and has been operational since 2010. There are a number of facilities currently in use at the Santa Elena site, and construction is currently underway for the new processing plant and underground mine.

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Much of the same infrastructure facilities utilized for the open pit mine will continue to be used for the expansion project, including, but not limited to, access roads, waste dumps, explosive magazines, office buildings, fuel storage facilities, power generation, primary crushing equipment, heap leach pads and solution collection ponds.  The material on the existing heap leach facility will be removed, adding additional space on the facility for rehandling of the tailings prior to transport to the waste dump as dry stack tailings. The additional facilities as part of the PFS include:

1.   A new CCD – MC processing facility, for which construction is currently underway.

2.   Upgrade of surface power generation facility.

3.   Construction of the underground decline, for which approximately 1,500 metres of development has been completed.

4.   A ventilation shaft with ventilation fans.

5.   A fresh air raise which will act as an escape way.

6.   Underground water recirculation facilities

7.   Underground electrical distribution system.

8.   Underground maintenance facilities.

9.   Dry stack tailings disposal facility (incorporated onto existing waste rock dump).

10.   Additional surface facilities such as mine dry and maintenance shops for the underground mine (maintenance shops already constructed).

1.7   Capital and Operating Costs

EBA has estimated total remaining capital costs for expansion of $88 million dollars including $9.3 million contingency, which excludes $24.2 million spent on the project from January 1st 2012 to 30 April, 2013.  The total remaining capital is made up of $48 million initial, required for construction of the plant, initial underground mining equipment purchases and preproduction underground development, $39.7 million sustaining capital, which includes development, mining equipment purchases and ongoing infrastructure requirements.  Capital costs for the project have been estimated for initial (prior to production from new facilities) and sustaining as shown in Table 1.6.  Note that the costs shown below exclude sunk costs, for construction completed on the process plant and for the underground decline.

EBA has estimated total operating costs at an average of $39 per tonne of ore processed.

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Table 1.6: Capital Cost Summary

Capital expense item	Estimated cost in US $
Total capital (including contingency) – Life of Mine	87,813,477
Initial capital (including contingency)	48,067,874
Mining equipment purchases	2,343,810
Development – UG	3,480,550
Process plant including EPCM	41,928,472
Electrical and services – UG	72,042
Indirect	243,000
Sustaining capital (including contingency)	39,745,603
Mining equipment purchases	5,144,587
Development	31,284,348
Electrical and services	2,283,346
Health and safety	143,442
Construction (incl. labour)	322,880
Indirect	567,000
Contingency amount as included in the above items	9,304,765
Budget spent from January 2012 to  April 30, 2013 (sunk costs)	24,263,480
Underground development completed as of April 30, 2013	5,300,000
Processing plant construction and mill equipment purchases completed as of April 30, 2013	18,963,479

Operating costs for the Santa Elena Expansion Project have been estimated for the underground mining, processing costs and general and administrative costs as summarised in Table 1.7.

Table 1.7: Operating Costs

Aspects of operations	Cost per tonne ore in US $
Average underground mining cost including ore development	$ 25  / tonne ore
Rehandling cost of heap leach ore	$ 0.62 / tonne ore
Processing of fresh ore	$ 20.50 / tonne ore
Processing of old heap leach ore	$ 18.50 / tonne ore
Handling of ore to tailings storage facility	$ 1 / tonne ore
G and A	$ 3.75 / tonne ore
G and A per year – based on Santa Elena operating costs for 2013	$3,650,000 /year
Average operating cost per ounce gold equivalent recovered*	$575 / oz.
Average operating cost per ounce silver equivalent recovered**	$11 / oz.

* Gold equivalent ounces calculated at gold to silver equivalency ratio of 1:52

** Silver equivalent ounces calculated at gold to silver equivalency ratio of 1:52

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1.8   Economic Analysis

EBA prepared an economic evaluation of the Santa Elena Expansion Project based on a pre-tax financial model and a post-tax model has been prepared by Tetra Tech.  For the economic analysis long-term consensus metal prices and exchange rate (as of April 30, 2013) used in the base case were as follows:

1.   Price of gold –                                 US $1,450/ troy oz.

2.   Price of silver -                                US $28 / troy oz.

3.   Closure costs of $5 million

4.   Sandstorm receives and pays SilverCrest $350/oz. plus applicable inflation on an estimated remaining 31,000 oz. of gold per agreement.

The pre-tax financial model was established on a 100% equity basis, excluding debt financing and loan interest charges.  The financial results of the base case and the two alternative cases are presented in Table 1.8. The economic analysis excludes the 2013 open pit heap leach production with an NPV of approximately $ 20 million based on projected production cash flows at a gold price of $1,250 and a silver price of $19.50.

Table 1.8: Base Case Economic Analysis Results

Aspects of Financial Analysis
Production
Gold Ounces Sold - post refiner	262,739
Silver Ounces Sold - post refiner	12,118,926
Revenue  US$(000)
Gross Sales	$684,931
Operating Expenses
Total Operating Costs1	$282,223
Freight & Refining	$5,579
Capital Expenses
Total Capital Costs2	$87,813
Pre-Tax Cash Flow
Total Cash Flow3	$302,481
Pre-tax financial results
Pre-tax NPV DCF @ 5%4	$223,670
Pre-tax IRR5	88%

1 Operating cost per AgEQ oz. sold varies between $9.6 and $12.6 over the life of mine

2 Excludes sunk costs, up to April 30, 2013, closure costs and working capital

3 Includes deductions of $5 million for closure costs and $1.8 million for working capital

4 Does not include 2013 revenue from open pit mining of $20 million

5 The IRR presented here do not reflect total project economics but reflect incremental project economics as they do not include sunk costs for the Santa Elena expansion prior to April 30 2013

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Sensitivities on the pretax financial results, for different gold and silver prices were undertaken as shown in Table 1.9 below.

Table 1.9: Comparison between the Base Case and the Spot Price for the Pre-Tax Financial Results

Item	Base Case	Spot Price
Gold Price (US$/oz.)	$ 1,450	$ 1,250
Silver price (US$/oz.)	$ 28	$ 19.5
IRR (%)**	88%	49%
DCF NPV @ 5.0% in millions**	$223.7	$108.7
Payback (production years)	1.1	1.7

* The IRR presented here do not reflect total project economics but reflect incremental project economics as they do not include sunk costs for the Santa Elena expansion prior to April 30 2013

** Post tax results have been included in section 22 for a 30% tax rate. The financial results exclude 2013 production revenue of an estimated $20 million.

The economic analyses consider SilverCrest delivering a remaining 31,000 ounces of gold (2014 and beyond) to Sandstorm Gold Ltd. (Sandstorm) at a price of US$350/oz under the Sandstorm Purchase Agreement executed on May14th, 2009. It does not include the option held by Sandstorm to participate in the future underground mine production which would require payment by Sandstorm of 20% (pro-rata of gold) of the capital costs with a subsequent increase of gold price for 20% of produced ounces to US$450/oz Au.  Sandstorm may elect to participate in the expansion, once SilverCrest provides a preliminary technical report that includes amongst other items, the required capital costs associated with the transition from open pit heap leaching to a conventional mill facility and underground mine.

1.9   Expansion Pre-Feasibility Recommendations

Further optimization of the mine schedule is warranted to investigate grade optimization versus stoping costs (long hole or cut and fill), potential to expand and accelerate increased underground production with a second ramp for the bottom of the pit and expand milling capacity. These optimizations will be addressed in detail after successfully announcing commercial production for the new mill facility in 2014.  Mineralization at Santa Elena is open in most directions with potential to further increase resources and reserves for increased production and mine life. Further infill and expansion drilling is recommended.

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2.0   INTRODUCTION AND TERMS OF REFERENCE

EBA Engineering Consultants Ltd. operating as EBA, A Tetra Tech Company has prepared this Pre-Feasibility Study (PFS) for SilverCrest Mines Inc. (SilverCrest, SVL) of Vancouver, British Columbia, Canada, for the Santa Elena Mine, located in central Sonora, Mexico.  Nusantara de Mexico S.A. de C.V., a 100% owned Mexican subsidiary of SilverCrest (SilverCrest for the purposes of this report), holds the rights to the Santa Elena Mine and is a legal operating entity of SilverCrest. The effective date of the report is April 30, 2013.

The Santa Elena Mine is currently producing gold and silver from an open pit and 2,500 tonne per day heap leaching operation, which commenced commercial production in July of 2011.  The Santa Elena Expansion Project involves combined processing of ore from the remaining reserves in the open pit, new reserves from underground development and reprocessing of spent ore from the existing heap leach pad in a new milling and processing plant.  Construction of the mill and plant facility is currently underway on the property.  An underground exploration decline is being developed and has intersected the low sulphidation Main Mineralized Zone at an elevation of 625 metres above sea level, which at this location is an estimated 14 m in true thickness.

The purpose of this report is to document a PFS completed for the Santa Elena Expansion Project and to support the revised Mineral Resource and Mineral Reserve Estimates announced on May 29, 2013.

EBA followed standard professional procedures in preparing the contents of this report. Data used in this report has been verified where possible and EBA has no reason to believe that the data was not collected in a professional manner. Technical data provided by SilverCrest for use by EBA and other consultants in this PFS is the result of work conducted, supervised, and/or verified by SilverCrest professional staff or their consultants.

This report conforms to National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101 and Form 43-101F1) and incorporates the Canadian Institute for Mining, Metallurgy and Petroleum Definition Standards for Mineral Resources and Mineral Reserves (CIM Definition Standards).

2.1   Report Authors and EBA Quality Control

This report has been completed by the following Independent Qualified Professionals:

•   Mike Tansey, P. Eng, Senior Mining Engineer, EBA Engineering Consultants Ltd.

•   James Barr, P.Geo., Senior Geologist, EBA Engineering Consultants Ltd.

•   Carlos Chapparo, P.Eng., Senior Geotechnical Engineer, EBA Engineering Consultants Ltd.

•   John Fox, P.Eng., Consulting Metallurgical Engineer, Laurion Consulting Inc.

•   Nick Michael, P.Eng, Senior Mineral Economist, Tetra Tech

SilverCrest is a Producing Issuer, as defined by National Instrument 43-101, and as such are entitled to Qualify and disclose material and technical information relating to the Santa Elena Mine with non-independent Qualified Persons.  This report includes Mineral Reserve Estimates for the open pit and spent ore on the leach pad which have been prepared by:

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•   N. Eric Fier, C.P.G., P. Eng, President and Chief Operating Officer, SilverCrest Mines Inc.

2.2   Site Visits

Site visits to the property have been undertaken by the report authors as described in Table 2.1 below:

Table 2.1: Independent QP’s in this study and date of their visit to the Santa Elena operations.

Author	Date of Site Visit
James Barr	Nov 26, 2010, May 12, 2011, May 10-11, 2012, Oct 13-14, 2012
Mike Tansey	Aug 16-18, 2011; Jan 23-25, 2012
Carlos Chaparro	7 – 12 July 2012
Nick Michael	N/A
John Fox	June 2013
N. Eric Fier	Numerous between January 2005 to April 2013

At site, discussions and meetings were held with;

•   Brent MacFarlane, Vice-President of Operations, SilverCrest Mine Inc.

•   N. Eric Fier, C.P.G., P. Eng, President and Chief Operating Officer, SilverCrest Mines Inc.

•   Dr. Salvador Aguayo, Vice-President of Development, Nusantara de Mexico

•   Ramon Gomez, Chief Geologist, Nusantara de Mexico

2.3   Effective Date of the Report

The effective date of the report has been set as April 30th, 2013, coinciding with the date when the leach pad drilling verification drilling was complete and all assay results were finalized at the lab.  The cut-off date for the Mineral Resource Estimate was March 15th, 2013, at which time all drill holes up to and including hole SE-13-145 were complete with final assays available from the lab.  All metallurgical work discussed in this report was completed in 30, April 2013.

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3.0   RELIANCE OF OTHER EXPERTS / DISCLAIMER

This report has been prepared by EBA for SilverCrest. The information, conclusions, opinions, and estimates contained herein are based on:

•   Information available to EBA at the time of preparation of this report,

•   Assumptions, conditions, and qualifications as set forth in this report,

•   Data, reports, and other information supplied by SilverCrest and other third party sources, and

•   Previous studies completed by EBA on the Santa Elena Mine.

Except for the purposes legislated under provincial securities laws, any use of this report by any third party is at that party’s sole risk.

4.0   PROPERTY DESCRIPTION AND LOCATION

The Santa Elena mine site is approximately 150 km northeast of the state capital city of Hermosillo, Sonora, Mexico near the intersection of 30º 01' north latitude, and 110º 09' west longitude (Figure 4.1). The community of Banamichi is located 7 km west of the property. The area is covered by the INEGI “Banamichi” 1:50,000 topographic map H12-B83.

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Figure 4.1: Santa Elena Property Location

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4.1   Mining Titles and Agreements

The property consists of six contiguous concessions with a total nominal area of 3,142.81 ha (Table 4.1, Figure 4.2). The concessions are registered with Mexico Mines Registry in Hermosillo and Mexico City in the name of Nusantara de Mexico S.A. de C.V. (Nusantara), a wholly-owned subsidiary of SilverCrest.  Nusantara filed the Santa Elena 7 concession which surrounds the five other concessions. All concessions were ground surveyed by a registered land surveyor at the time of staking.

Table 4.1: List of Mining Titles

Concession Number	Date	Concession Name	Owner	Size (ha)
192174	1983	Santa Elena	Nusantara de Mexico	24.19
178094	1983	Santa Elena No 4 Fraccion SE	Nusantara de Mexico	0.06
176544	1983	California Nusantara de Mexico	Nusantara de Mexico	18
221460	1995	Elena 5	Nusantara de Mexico	399.87
223533	2003	Santa Elena 6	Nusantara de Mexico	841
227239	2006	Santa Elena 7	Nusantara de Mexico	1,859.63
180494	1983	Santa Elena No 3 Fraccion SW Tungsteno de	Nusantara de Mexico	0.06
			Total	3142.81

Mining regulations in Mexico provide all concessions to be valid for a period of 50 years. Taxes are based on the surface area of each concession and are due in January and June of each year at a total annual cost of approximately $10,000. All tax payments have been paid to date.

On December 8, 2005, Nusantara had the right to acquire a 100% interest in the Santa Elena property from Tungento de Mexico SA de CV.  On August 14, 2009, Nusantara exercised the option to acquire 100% of the Santa Elena property.

On November 12, 2007, Nusantara signed an agreement with the Community of Banamichi (Ejido) for a 20 year lease on surface rights for a maximum of 841 ha with respect to exploration and exploitation. Lease payments were dependent (at Nusantara’s discretion) on the number of hectares required for a given year. A minimum of 285 ha was required for initial exploitation. Currently 431 ha are under lease for exploitation and exploration.  The annual cost per year ranged from approximately $55,000 to $160,000 dependent on the number of hectares required. Lease obligations have been met to date.

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Figure 4.2: Santa Elena Property Map

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4.2   Royalties

No known private or state royalties currently exist for the Santa Elena mine.

4.3   Sandstorm Gold Stream

On May 14, 2009, SilverCrest entered into a definitive Purchase Agreement with Sandstorm Gold Ltd. (Sandstorm) under which SilverCrest’s wholly-owned Mexican subsidiary, Nusantara de Mexico S.A. de C.V., agreed to sell 20% of future gold production from the Santa Elena open pit, or a total of 50,000 ounces gold, to Sandstorm in exchange for an Upfront Deposit of $12,000,000. The agreement also provides for ongoing per-ounce payments by Sandstorm equal to the lesser of $350 and the prevailing spot gold market price upon delivery of gold.  In addition, Sandstorm issued $3.5 million common shares to SilverCrest. The per ounce price of $350 is subject to an increase of 1% per annum commencing on the 3rd anniversary of the date that Santa Elena Project began commercial production (i.e. July 2014).

If SilverCrest decides to develop an underground mine on the Santa Elena Project, Sandstorm has the right to review the project, accept the terms of the agreement and if acceptable purchase 20% of the gold from the underground mine (includes reprocessing the leach spent ore) at a per ounce price equal to the lesser of $450 and the prevailing spot gold market price, subject to an increase of 1% per annum beginning on the 3rd anniversary from the date the underground mine begins commercial production. In exchange, Sandstorm will pay 20% of the upfront capital expenditures based on 20% of the pro rata gold to be produced, including sunk costs incurred to date (pre April 30, 2013). The Pre-Feasibility Study upfront capital costs combined with sunk costs incurred and approved by SilverCrest’s board will be the basis for Sandstorm’s review and potential investment.

4.4   Environmental Liabilities and Permitting

Currently, SilverCrest has maintained all of the necessary permits for exploration and exploitation at the Santa Elena mine site. A Manifestacion de Impacto Ambiental (MIA) was submitted to Secretaría de Medio Ambiente y Recursos Naturales (SEMARNAT) in early January of 2013 for amendment of the land use licence related to the underground expansion project and was approved in May 2013.  This amendment approval allows for tailings facilities that were not previously approved for the current operation.

A mining concession in Mexico does not confer any ownership of surface rights. However, use of surface rights for exploration and production can be obtained under the terms of various acts and regulations if the concession is on government land. The Santa Elena concessions are located on Ejido (community or co-op) land, and as of November 12, 2007, an agreement with the surface owners has been signed which allows SilverCrest access and authorization to complete exploration and mine operations activities.

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Table 4.2: Permits for Santa Elena

Permit	Mining Stage	Agency	Status
Environmental Impact Statement (MIA)	Construction/operation/abandonment	Secretary of environment and natural Resources (SEMARNAT)- State offices	in place
Land use change study	Construction/operation	SEMARNAT-General Department of Permitting for Forestry and soils (DGGFS)-State offices	in place
Land use license	Construction	Banamichi Municipality	in place
Explosive handling and storage permits	Construction	National Secretary of Defense (SEDENA). (Need approval from state and municipal authorities)	in place
Archaeological release letter	Construction	Construction Nation Institute of Archaeology and History (INAH)	in place
Water use concession title	Construction/operation/prior to utilization of water	National Commission of Water (CAN)-State offices	in place
Water discharge permit	Operation	CAN-(State offices)	in place
Construction permit	Construction	Municipality of Banamichi	in place

5.0   ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCUTRE AND PHYSIOGRAPHY

5.1   Accessibility

The Santa Elena property can be accessed easily year round by paved highways 90 km east from Hermosillo to Ures, then 50 km north along a paved secondary road to the community of Banamichi, then by a gravelled maintained road 7 km east to the mine site.

5.2   Climate

The climate is typical Sonoran desert, with a dry season from October to May. Average rainfall is estimated at 300 mm per annum. There are two wet seasons, one in the summer (July to September) and another in the winter (December). The summer rains are short with heavy thunderstorms whereas the winter rains are longer and lighter. Seasonal temperatures vary from 0ºC to 40ºC. Summer afternoon thunderstorms are common and can temporarily impact the local electrical service. Flash flooding is common in the area.

5.3   Local Resources

5.3.1   Water Supply

The main supply well currently used at the mine site, PSA-1, was installed and tested in September 2009 and a pump installed at approximately 109 metres depth.  Pumping tests indicated the well was located in a confined aquifer with a potential association to geothermal sources and an estimated sustained supply of 25 L/s, sufficient to support the open pit mining operation (Breckenridge, 2010).

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Well PSA-2 was constructed in the summer of 2011as a back-up well to support Santa Elena Mine’s primary water supply, Well PSA-1 (formally MW-1). The well is 171 meters deep, and screened from a depth of 72m to the base of the well. There is a steel casing that extends from surface to a depth of 72m. The casing and screen section are eight inches in diameter. The well was started by using an air rotary drill; however, due to the unstable and unconsolidated nature of sediment in this portion of the sand and gravel aquifer the well was completed using a mud rotary drill.

The well was pumped at a rate of 10.6 liters per second (lps) over a 72 hour period, and the final measured drawdown in the well was approximately 20 meters. During the pumping test the water level dropped from its static level at 68m to a depth of 90m within a few minutes. The water level stabilized at 90 meters after two hours of pumping. The testing suggests that the well could be pumped indefinitely at a rate of 10.6 lps with no increase in drawdown (GRE, 2011a). The safe yield of the well is 12.9 lps, based on the criterion that drawdown equivalent to 25% of the aquifer thickness is allowable (GRE, 2011a).

Due to the presence of the screen and the fact that it is open to a smaller aquifer depth, it is noted that Well PSA-2, although very productive, is less productive than Well PSA-1. The screen in general reduces the ability of water to enter the well, and the vertical-slot screen used in Well PSA-2 was not optimized for maximum yield. The pumping test indicates that there is minor hydraulic interference between the two water wells, and that the cumulative safe yield of the well field, with both wells pumping at the same time is 35 lps (GRE, 2011a).

5.3.2   Power

A minor amount of electrical line power available from nearby sources currently supplies municipalities and agriculture, but is insufficient for the Santa Elena operation. Power for current operations is provided by diesel generators. Future line power is possible, but requires permitting and a significant capital expenditure.

5.3.3   Community Services

The Santa Elena Mine is located near to the village of Banamichi for which accommodations and local food services are found.  The mining centre of Cananea is the closest sizable urban area (pop. est. 30,000), and is located approximately 100 km north by road from the Santa Elena mine site. Most services and supplies are available in Cananea.  Sonora’s capital city, Hermosillo, is located approximately 150 km southwest of the property, and is regarded as the main industrial hub for the majority of the local mining operations.  Services are available for heavy machine purchase and repair, materials fabrication, and engineering services. Alternatively, Tucson Arizona is approximately a 4 hour drive north across the international Mexican USA border from the Santa Elena mine site.

Northern Mexico has significant precious and base metal mines and there is a significant workforce of trained mining and processing personnel. Many of the trades and skills learned in the region would be transferable to the new operations. The nearby Cananea and La Caridad mines are amongst the largest mines in North America.

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5.4   Existing Infrastructure

The Santa Elena open pit heap leach mine was constructed in 2009 and 2010 with the following operational infrastructure:

§   7 km long main access road from paved highway and local community of Banamichi,

§    2,500 tpd ore open pit mine,

§   Waste dump with the estimated permitted capacity of 35 million tonnes,

§   3-stage crusher,

§   Two lined and certified leach pads,

§   Lined and certified barren and pregnant solution pond,

§   Lined and certified emergency pond designed for 100 year event,

§   Merrill Crowe plant and refinery,

§   On-site laboratory for production and exploration work,

§   Administration office,

§   Diesel generators (total 2.5 MW capacity, 4 generators)

§   Core storage and logging facility, and

§   All required piping, power and security.

Construction of infrastructure to support an underground mine expansion, new mill and processing plant facility, is currently underway on site.  The new mine infrastructure is described in detail in Section 18.0.

5.5   Physiography

The Santa Elena property is located on the western edge of the north trending Sierra Madre Occidental mountain range geographically adjacent to the Sonora River Valley. Property elevation ranges from 800 m ASL to 1,000 m ASL. The property is located on the range front at a low elevation in relation to the mountains immediately east and west respectively.

Vegetation is scarce during the dry season, limited primarily to juvenile and mature mesquite trees and cactus plants. During the wet season, various blooming cactus, trees and grasses are abundant in drainage areas.

6.0   HISTORY

This section has been adapted from the initial Mineral Resource report authored by Fier (2007), who conducted much of the original research into historical work on the property through archived company and government records.  EBA has not independently reviewed the information nor verified any reported historical grade or production values.  EBA feels these historic numbers are insignificant and irrelevant in scale to the current and future production at the Santa Elena Mine, and at least four primary levels (down to 75 metres below surface) of the historic underground workings have been removed within the shell of the current producing open pit.

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Earliest records of work on the property are from the late 19th to early 20th century when an English company by the name of Consolidated Fields operated a mining operation at Santa Elena until the onset of the Mexican Revolution in 1910.  It is estimated that the most extensive underground development on the property occurred during this period.  The development included a 450 ft. (137 meters) two compartment shaft, a single compartment 100 m inclined shaft, and eight to nine working levels at a spacing of approximately 15 m to 20 m with numerous cross-cuts and raises.  The two compartment shaft was caved near the surface prior to current development and the depth was not been verified.  Four of the levels (surface to 75 m in depth) and a total of approximately 1.5 km of development were accessed by SilverCrest in 2006.  Ongoing open cut mining has made the underground workings unsafe to enter.  Stoping in the upper accessible levels removed an estimated 57,000 tonnes determined by a void survey conducted by SilverCrest in 2010.  No detailed production records are available for the Consolidated Fields work; however, SilverCrest estimates that approximately 35,000 tonnes of the original tailings from Consolidated Fields’ operations remain onsite.  During the 1960’s, Industrias Peñoles S.A de C.V. drilled two or three holes on the property. No records are available for this drilling.

During the early 1980’s, Tungsteno de Baviacora (Tungsteno) mined 45,000 tonnes grading 3.5 g/t gold and 60 g/t silver from an open cut at Santa Elena.  This material was shipped for processing to the company’s flotation mill near Baviacora, approximately 30 km southwest of Santa Elena.  The 50 ton (short ton) per day mill was specifically built for processing tungsten ores from a nearby deposit from 1977 to 1983.  The tonnage from Santa Elena was supplemental to the tungsten production.  Very limited records from the production are available but the owner has stated that recovery was adequate for the Santa Elena tonnage but that some metal concentration still remains in the tailings on site.

After 2003, Tungsteno periodically surface mined high silica/low fluorine material from Santa Elena and shipped it to the Grupo Mexico smelter in El Tajo near Nacozari, approximately 60 km to the northeast.

During 2003, Sergio A. Trelles Monge, CPG and Qualified Person conducted an exploration program for Tungsteno at Santa Elena.  Sr. Trelles was not considered an “independent” QP for the purposes of this work.  The program consisted of the collection of 117 surface and underground samples.  A sample summary report is available for review but sample lengths and locations are not clear. In late 2003, Nevada Pacific Gold Inc. of Vancouver B.C. completed a brief surface and underground sampling program with the collection of 119 samples. A report was completed but subsequently misplaced.  Only the ALS-Chemex assay sheets and a rough location map were available for review. The property remained under control of Tungsteno.  The above data is considered historical and unreliable and has not been used in the current study.

In early 2004, Fronteer Development Group of Vancouver, B.C., completed an extensive surface and underground mapping and sampling program.  A total of 145 channel samples (89 underground and 56 surfaces) were collected and analyzed by ALS Chemex of Hermosillo, Mexico.  The property remained under control of Tungsteno.  This data has been used by SilverCrest for exploration purposes only and has not been used in this PFS.

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On December 8, 2005, SilverCrest, using its 100% owned subsidiary Nusantara, entered into an option agreement with Tungsteno to acquire a 100% interest in the Santa Elena property through staged option payments over 5 years. SilverCrest completed payments as per the terms of the agreement in August of 2009. SilverCrest’s Santa Elena open pit mine has been in commercial production of gold and silver since 2011, with a mining rate of approximately 2500 tonnes per day.

7.0   GEOLOGICAL SETTING

7.1   Regional Geological Setting

Santa Elena is located in northwestern Mexico where much of the geology can be attributed to the subduction and related volcanism of the Farallon Plate beneath the North American Plate.  The east-directed subduction of the Farallon Plate began approximately 200 Ma ago with the tectonic rifting of the supercontinent Pangea (Rogers 2004). The resulting NW-SE trending Sierra Madre Occidental extends from the USA-Mexican border to Guadalajara in the southeast, a distance of over 1,200 kilometers.

Delgado-Granados et al. (2000) propose that subduction of the Farallon Plate occurred at a relatively shallow angle, resulting in continental uplift across northern Mexico with accretionary terrains developing along the western fringes.  The shallow subduction is also thought to be responsible for the tectonics that produced the Laramide orogeny.

Continental arc volcanism culminated with the Laramide orogeny in the early to late Eocene (Alaniz-Alvarez et al., 2007).  The waning of compression coincides with the first part of Basin and Range extension (Wark et al., 1990; Aguirre-Diaz and McDowell, 1991, 1993).

The silicic volcanism is thought to be related to fractional crystallisation of mantle sourced basalts from subduction (Johnson, 1991; Wark, 1991).  Ferrari et al (2007) summarises five main igneous deposits of the Sierra Madre Occidental;

§   Plutonic/volcanic rocks: Late Cretaceous –Paleocene.

§   Andesite and lesser Dacite-Rhyolite: Eocene (Lower Volcanic Complex).

§   Silicic ignimbrites: Early Oligocene & Miocene (Upper Volcanic Complex).

§   Basaltic-andesitic flows: late stage of and after ignimbrites pulses.

§   Repeat and episodic volcanic events related to rifting of the Gulf of California (alkaline basalt and ignimbrite) emplaced to western flanks: Late Miocene Pliocene and Quaternary

During the final stages of the deformation period during the Paleocene – Early Eocene, E-W and ENE-WSW extension occurred in the Lower Volcanic Complex that now hosts many porphyry deposits of the Sierra Madre Occidental. These porphyry deposits are hosted in Middle Jurassic to Tertiary aged intrusions, located at Cananea, Nacozari and La Caridad. (Ferrari et. al., 2007).  This E-W and ENE-WSW extensional direction is similar in orientation to the Santa Elena vein.

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Early Oligocene extensional tectonics occurred along the eastern Sierra Madre Occidental flank forming the typical basin and range province. By early to mid-Miocene, extension migrated west into Northern Sonora and along the western flank of the Sierra Madre Occidental resulting in N-NW striking normal faults.  This extensional regime caused major deformation across the Sierra Madre Occidental exhuming pre-Cambrian basement rocks, especially in the Northern Sierra Madre Occidental (Ferrari et. al., 2007).

Northwest trending shear and fault zones appear to be an important control on mineralization in the Sonora region. Mineralizing fluids may have been sourced from Cenozoic intrusions. The structural separation along the faults formed conduits for mineral bearing solutions. The heat source for the mineralizing fluids was likely from the plutonic rocks that commonly outcrop in Sonora.

To the west of the Sierra Madre Occidental are the parallel ranges and valleys that show structural similarities to the extensional tectonic regimes of the Basin and Ranges Province to the east.  Elevations in the west are lower than the eastern Provinces, with transition to the Coastal plains and Gulf of California.

7.1.1   Local and Property Geology

The Santa Elena property is located at the northwest extent of the Sierra Madre Occidental.  The primary rock types observed on the property are Tertiary andesite and rhyolite flows (Figures 7.1)).  These units strike approximately N-S and dip 10º to 45º E-NE.  All the volcanic units in the immediate vicinity of the Santa Elena deposit exhibit propylitic to silicic alteration.  Within the main mineralized structure, widespread argillic alteration and silicification proximal to quartz veining is present. Chloritic alteration increases away from the mineralized zone in the andesite beds.

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Figure 7.1: Santa Elena Property Geology Map

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The Main Mineralized Zone (MME) is hosted within an E-W trending structure that cross-cuts the volcanic units.  The structure hosts an epithermal quartz calcite vein that has been mapped for approximately 1.2 kilometres in length, and ranges from 1 to 35 metres wide, averaging approximately 15 metres. The structure dips from 40º to 60º to the south, and has been drill-tested to approximately 600 m below surface.

Andesite and granodorite dykes have been identified at the Santa Elena deposit which are adjacent and sub-parallel to the MMZ. The location of the mineralizing heat source has not yet been linked to any local plutonic phases. However, a large intrusion that exists approximately 10 km east and north of Santa Elena may be associated with the mineralization.

The main structure (MMZ) hosts quartz and stockwork veining and displays typical epithermal textures, including: banded quartz; vuggy quartz; and brown-black bladed calcite (pseudomorph to quartz) with many of these textures intermixed with hydrobreccia reworking. Some fracturing is found at localized fault intersections and adularia has been identified in many of the core holes.

In late 2007, an induced polarization (IP), resistivity and magnetometer geophysical survey was completed by Pacific Geophysical Ltd. of Vancouver B.C. Results showed the Main Mineralized Zone is a resistivity high (silica) and IP low (minor sulphides) that can be traced for approximately 1.2 km along strike of the zone.

The oxidation-reduction (redox) boundary within the deposit is noted to persist to a depth of approximately 100 metres below surface, with a transitional horizon for an additional 20-30 metres, or 120-130 metres total depth.  Visually, the oxidized material is characterized by the presence of iron oxides, including limonite, jarosite, goethite and hematite along fractures and joints. The iron oxides are associated with gold and silver mineralization.  Below the redox boundary, the vein remains intact as a milky white quartz-calcite vein with rare traces of fine grained pyrite, or primary iron-sulphides.

7.1.1.1   Structural Geology

Arroyo Tinaja Fault

Regional interpretation from recent topographic mapping, regional geology maps and recent drilling suggests that the dry river valley, el arroyo Tinaja immediately east of the current open pit, may be associated with a large scale regional fault. This proposed regional fault is estimated to strike 033° and dip steeply to sub-vertically.  Ongoing work is being conducted to confirm the presence and character of the fault system where it potentially intersects the Santa Elena vein underground. Minor displacement has been identified along the fault based on surface mapping and subsurface drilling.

Northwest-Trending Structures

Northwest-trending splay and cross-cutting structures appear to influence mineralization at intersections with the main vein and along a NW-SE (striking approximately 135 with steep westerly dip) trend in the footwall of the vein.  These narrow quartz and calcite filled planar brittle structures appear to cross-cut and postdate the emplacement of the main vein at Santa Elena.  The structures display weak alteration haloes of fine grained pyrite in addition to the presence of  anomalous gold and silver concentrations (hole SE-12-98 5.72 gpt Au and 201 gpt Ag over 4.90 metres, down hole width).

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8.0   DEPOSIT TYPE

Mineralization at Santa Elena occurs as a series of replacement veins, stockwork and hydrothermal breccias typical of other low-sulphidation epithermal deposits found in the Sierra Madres and elsewhere in the world. Similar deposits include the La Colorada deposit in Sonora Mexico, Pinos Altos deposit in Chihuahau Mexico, El Peñón deposit in Chile, and those deposits occurring in the Midas and Oatman districts of Nevada and Arizona in the USA.  These deposits form in predominantly sub-aerial felsic volcanic complexes in extensional and strike-slip structural regimes.  Samples previously collected by various parties, including SilverCrest, show a geochemical signature consistent with a high calcium, high level, low-sulphidation system (Au-Ag-Sb-Pb-Zn-Ba-Ca-Mn).

The mineralization is the result of silica-rich fluids ascending into a structurally controlled, near-surface environment. Mineral deposition takes place as the fluids undergo cooling by fluid mixing, boiling and decompression. Brecciation of the mineralized zone appears to be due to explosive venting from an assumed intrusive at depth.  The Main Mineralized Zone is hosted in major E-W structure cross cutting the andesite units close to contact with Rhyolite,

8.1   Mineralization

The Santa Elena deposit contains gold and silver (typically as electrum) with minor lead, zinc and copper.  The structure consists of multiple banded quartz veins and stockwork with associated adularia, fluorite, calcite and minor sulphides. Bonanza ore shoots (greater than 500 gpt silver and 30 gpt gold) appear to be locally present (hole SE-12-106 with average uncapped grades of 35.94 gpt Au and 1,027 gpt Ag over 2.37 metres, not true width) but require more definition to determine their full extent.

The permeable nature of the fractured zones has allowed partial oxidation to locally occur to maximum depth of 400 metres below the surface in selective fractured zones.  Minor disseminated sulphides and possibly rhodonite have been noted at approximately 530 vertical metres and shows minor disseminated sulphides and rhodonite.

Metal zonation appears to exist with higher grades and thicker mineralized widths near the epithermal boiling zones, one of which daylights in the open pit area.  A trend of higher grades and thicker veining is apparent with a plunge of approximately 25° to the east.  Drill hole SE-12-74 intersected the vein at approximately 500 vertical metres depth with an average uncapped grade of 1.56 gpt Au and 133 gpt Ag over 7 metres (not calculated as true width) along this plunging trend from the current open pit operation.

Zonation also appears to correspond to northwest-trending structures that cross-cut the Main Mineralized Zone forming high grade shoots.  Vertical zonation shows gold content consistent with depth and silver content increasing. At the surface, the silver to gold ratio is 20:1. At 500 m below surface, the ratio is approximately 100:1.

Minor sulphides have been observed in a few locations within the mineralized zone.  The andesite in the hanging wall shows disseminated pyrite averaging 5%. Calcite is found in close proximity to pyrite and averages about the same. Select locations in the hanging wall show +30% finely disseminated pyrite spatially associated with +30% disseminated and veinlet calcite. Hydrothermal breccias exist in the hanging wall andesite proximal to the Main Zone with drill holes (SE 07-20 and 21) intersecting up to 200 m of breccia with a pyrite/calcite matrix. Selective sampling shows some anomalous lead, zinc and copper, suggesting a possible mineralized intrusive (porphyry) at depth. Initial results from the 2007 geophysical program indicate deep-seated IP highs.

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Alteration within the deposit is widespread and pervasive with the most significant being silicification, kaolinization, and chloritization.  Kaolin and alunite occur primarily along deeply weathered and oxidized structures and the fractured-andesite contact.  Limonite within the oxide zone consists of a brick-red colour after pyrite, brown goethite and local yellow jarosite.  Manganese occurs locally as pyrolusite and minor psilomelane near the surface.

Gangue minerals consist of quartz, calcite, adularia, chlorite and fluorite. Black calcite can be a significant gangue mineral found within the Main Mineralized Zone.  Analysis shows calcite content of up to approximately 15% which has a direct metallurgical benefit for processing ore.

9.0   EXPLORATION

Exploration carried out previous to SilverCrest’s acquisition has been discussed in Section 6.0 and is not repeated here except for clarity purposes.

In 2006 and the first half of 2007, SilverCrest completed an extensive exploration program at Santa Elena. This work included surface mapping and channel sampling, underground mapping, underground channel sampling and core drilling. The surface program was conducted in May 2006 under the direction of N. Eric Fier, CPG, P.Eng.

A total of 289 samples were collected and analyzed by ALS-Chemex in Hermosillo, Mexico and North Vancouver, BC.  More detail regarding this program can be found in the NI 43-101 Technical Report: ‘Reserve Update for the Santa Elena Open Pit and Preliminary Assessment for the Santa Elena and Cruz de Mayo Expansion Project, Sonora, Mexico’ (May 11, 2011).

In 2012 SilverCrest began construction of the underground portal and exploration decline.  Mapping was conducted along the decline for geotechnical purposes in addition to sampling.

10.0   DRILLING

10.1   Historical Drilling

No information of historic drilling was available to EBA.  The drill hole database and current Mineral Resource estimate include only SilverCrest drilling that has been completed since 2006.

10.2   Previous Drilling by SilverCrest

SilverCrest completed a drill program in early 2006 consisting of 19 NQ holes totalling 2,572 m. Drilling was completed by Major Drilling de Mexico, a subsidiary of Major Drilling Canada of Ontario using a Longyear 38 drill and associated support equipment. Drilling was conducted at 100 metre spacing with all holes but two drilled north at angles from -45 to -70º along the E-W trending strike of the mineralized zone.

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SilverCrest completed a core drill program in 2007 consisting of 21 holes totalling 1,951.4 m, and a drill program consisting of an additional 48 core holes (9,939 m), 4 geotechnical core holes (1,163 m) and 21 RC drill holes (4,308 m) in 2008. Drilling for both programs was completed by Cabo Drilling de Mexico, a subsidiary of Cabo Drilling Corporation of North Vancouver B.C., using a Longyear 38 drill. The 2006 to 2008 drilling was termed Phase II.

In 2009, SilverCrest completed a pre-production in-fill drill program consisting of 11 RC holes (577 metres). RC drilling was completed as pre-collars for coring and condemnation drilling in the proposed waste dump and leach pad areas.

From 2006 to 2009 a total of 144 combined drill holes have been drilled at Santa Elena for a total of 25,055.8 metres (Table 10.1).

Table 10.1: Summary of 2006-2009 Drill Holes

Year	Diamond Hole (DD)		Reverse Circulation (RC)
	No.	Length (m)	No.	Length (m)
2006	19	2579.94
2007	50	5816.53
2008	34	10983.94	21	4164
2009			20	1461.39
Total	103	19,380.41	41	5,625.39

More detail regarding this program can be found in the 2011 NI 43-101 Technical Report; ‘Reserve Update for the Santa Elena Open Pit and Preliminary Assessment for the Santa Elena and Cruz de Mayo Expansion Project, Sonora, Mexico (May 11, 2011).

10.3   2012-2013 Drilling

During 2012-2013, SilverCrest targeted delineation of shallow, below-pit mineralization and deep mineralization (Table 10.2) with additional drilling. Three drilling companies were contracted; Major Drilling de Mexico based in Hermosillo, Mexico, Guardian Drilling from Saskatchewan, Canada, and DrilCor based in Durango, Mexico.  A number of the deep drill holes were RC collared with diamond tails at depth in the mineralization. Figure 10.1 shows the location of the 2012 to 2013 drill holes.  Appendix D has a list of the 2012-2013 drill holes and locations.

This drilling focused on delineating and extending the areas along trend and down-dip of the vein. Other drilling was located off strike to explore for near parallel mineralization.

A total of 20 drill holes were collared using reverse circulation (RC) to expedite hanging wall drilling, then finished with diamond core from approximately 40-50 metres before the vein target depth. (Table 10.2).  This practice was discontinued due to significant deviation in the pre-collared holes. A total of 149 holes (48,980.18 m including RC with DD tails) were drilled during the 2012-13 program, including holes drilled from within the current pit (Table 10.2).

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Table 10.2: Summary of all 2012-2013 Santa Elena Drilling.

Year	Diamond Hole (DD)		Reverse Circulation (RC)		RC with DD tail
	No.	Length (m)	No.	Length (m)	No.	Length (m)
2012	79	25,727.93	20	1,890.9	14	6,338.1
2013	36	15,023.25
Total	115	40,751.18	20	1,890.9	14	6,338.1

Down hole surveys were performed on a total of 134 drill holes from the 2012-2013 drill program.  Surveys were taken at 50 metre intervals down the hole in addition to just below the casing and at the end of the hole.

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Figure 10.1: Map showing 2012-2013 Santa Elena Drilling

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Table 10.3 show significant drill hole intercepts from the 2012-2013 drilling campaign.

Table 10.3: Significant Drill Hole Intercepts from 2012-2013 Drilling

HOLE-ID	From (m)	To (m)	Interval (m)	Au (gpt)	Ag (gpt)
SE-12-39	152.18	166.10	13.92	2.03	153.55
SE-12-40	167.23	179.21	11.98	2.60	141.00
SE-12-41	150.63	157.63	7.00	3.40	223.41
SE-12-42	73.12	80.60	7.48	1.53	85.36
SE-12-44	90.90	97.65	6.75	2.48	72.44
SE-12-45	179.20	191.85	12.65	2.22	67.41
SE-12-46	169.41	178.89	9.48	1.60	105.34
SE-12-47	193.70	210.70	17.00	2.14	107.02
SE-12-48	174.25	194.00	19.75	1.43	111.76
SE-12-49	152.30	166.30	14.00	0.20	15.34
SE-12-50	222.85	231.00	8.15	3.54	220.74
SE-12-51	205.00	209.00	4.00	2.96	266.00
SE-12-52	183.51	186.80	3.29	1.92	92.83
SE-12-54	91.50	123.72	32.22	1.61	169.59
SE-12-56	205.20	213.45	8.25	4.74	133.38
SE-12-57	226.00	254.70	28.70	0.64	114.28
SE-12-60	416.54	420.91	4.37	0.31	111.43
SE-12-60	432.65	441.75	9.10	5.67	158.01
SE-12-61	453.88	466.38	12.50	1.11	62.41
SE-12-62	394.77	400.00	5.23	15.34	546.22
SE-12-63	261.96	269.00	7.04	1.14	97.15
SE-12-64	285.50	292.60	7.10	1.72	101.73
SE-12-65	234.60	245.28	10.68	1.49	156.90
SE-12-66	324.90	326.85	1.95	0.32	53.71
SE-12-67	208.57	230.00	21.43	1.70	136.09
SE-12-70	239.20	258.87	19.67	0.98	105.09
SE-12-71	289.53	300.06	10.53	1.00	98.13
SE-12-72	472.67	478.79	6.12	2.28	192.08
SE-12-73	489.00	502.01	13.01	1.55	324.83
SE-12-74	572.51	577.49	4.98	2.11	177.11
SE-12-75	439.70	447.95	8.25	2.82	107.49
SE-12-77	343.35	347.80	4.45	1.76	111.70
SE-12-79	307.00	309.10	2.10	3.40	237.97
SE-12-79	316.25	319.57	3.32	1.34	543.37

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HOLE-ID From (m) To (m) Interval (m) Au (gpt) Ag (gpt)

SE-12-80	248.65	265.34	16.69	0.94	95.56
SE-12-81	359.20	364.55	5.35	0.53	48.15
SE-12-82	423.59	426.05	2.46	0.96	65.93
SE-12-82	433.95	443.99	10.04	1.61	76.50
SE-12-83	379.60	387.90	8.30	167.20	2.76
SE-12-84	404.22	423.47	19.25	2.16	88.61
SE-12-85	413.20	415.95	2.75	3.81	64.21
SE-12-86	390.70	396.80	6.10	4.18	256.64
SE-12-87	326.17	330.47	4.30	0.21	20.73
SE-12-89	418.74	422.33	3.59	2.37	82.03
SE-12-89	429.06	444.57	15.51	0.65	53.77
SE-12-90	160.07	172.57	12.50	2.59	139.37
SE-12-90	191.13	197.82	6.69	3.00	175.67
SE-12-91	405.86	421.42	15.56	23.97	307.90
SE-12-92	258.32	261.49	3.17	4.03	637.27
SE-12-94	243.52	252.75	9.23	1.30	191.45
SE-12-94	296.95	311.84	14.89	1.84	206.54
SE-12-95	269.15	278.34	9.19	1.41	196.20
SE-12-96	263.25	267.40	4.15	0.71	48.89
SE-12-102	253.70	257.95	4.25	2.89	68.49
SE-12-102	266.50	269.70	3.20	1.02	179.07
SE-12-103	171.95	197.12	25.17	1.22	155.98
SE-12-104	197.48	202.54	5.06	0.52	78.19
SE-12-104	206.72	210.72	4.00	2.56	174.13
SE-12-104	217.19	231.48	14.29	1.12	104.12
SE-12-105	154.10	167.53	13.43	4.02	200.01
SE-12-106	172.33	189.07	16.74	3.02	213.70
SE-12-106	209.84	215.69	5.85	16.13	569.55
SE-12-107	206.15	209.00	2.85	1.01	85.43
SE-12-108	221.80	231.05	9.25	1.10	78.16
SE-12-109	242.85	251.00	8.15	1.32	153.44
SE-12-110	225.46	233.00	7.54	1.39	308.56
SE-12-111	507.91	514.87	6.96	0.74	51.14
SE-12-113	199.53	211.79	12.26	0.81	81.24
SE-12-114	219.93	248.10	28.17	1.59	115.55
SE-13-115-2	227.44	233.41	5.97	0.41	54.70
SE-13-115-2	247.52	256.35	8.83	0.70	56.69
SE-13-115-2	260.30	277.45	17.15	0.66	69.92

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HOLE-ID From (m) To (m) Interval (m) Au (gpt) Ag (gpt)

SE-13-116	451.47	455.47	4.00	5.87	281.50
SE-13-117	176.00	190.00	14.00	0.98	198.64
SE-13-117	234.12	237.86	3.74	1.12	123.73
SE-13-117	266.37	276.82	10.45	0.78	76.70
SE-13-118	188.16	204.13	15.97	1.05	111.98
SE-13-118	334.95	339.11	4.16	0.90	70.80
SE-13-119	215.10	240.90	25.80	1.24	187.65
SE-13-119	243.82	246.63	2.81	1.47	99.74
SE-13-121	449.25	458.93	9.68	2.86	126.93
SE-13-122	249.50	262.50	13.00	1.06	144.83
SE-13-122	270.68	276.42	5.74	1.06	63.56
SE-13-123	433.48	438.28	4.80	4.85	101.07
SE-13-124	271.63	276.23	4.60	0.65	81.47
SE-13-124	283.84	287.07	3.23	0.77	70.13
SE-13-125	474.80	474.80	0.00	0.95	117.03
SE-13-126	446.31	446.31	0.00	1.46	169.11
SE-13-127	169.84	193.61	23.77	1.68	89.54
SE-13-128	469.73	478.01	8.28	0.08	138.87
SE-13-128	516.20	529.31	13.11	1.43	170.51
SE-13-129	236.00	239.49	3.49	0.96	78.17
SE-13-130	403.08	409.09	6.01	7.21	143.72
SE-13-131	162.67	179.99	17.32	0.52	57.33
SE-13-131	269.73	274.87	5.14	3.94	139.70
SE-13-132	224.21	229.24	5.03	0.84	71.92
SE-13-133	437.21	445.87	8.66	10.36	159.80
SE-13-134	481.12	485.87	4.75	0.55	57.63
SE-13-134	490.64	495.85	5.21	1.03	141.26
SE-13-135	265.14	266.72	1.58	0.34	53.55
SE-13-136	458.53	478.10	19.57	1.08	84.44
SE-13-137	235.58	240.07	4.49	0.55	90.14
SE-13-138	464.42	473.20	8.78	0.49	75.99
SE-13-139	253.11	263.88	10.77	0.69	34.84
SE-13-139	287.25	292.91	5.66	0.83	44.15
SE-13-140	215.30	234.10	18.80	0.77	125.51
SE-13-141	448.35	457.44	9.09	4.16	189.08

Also in 2012, 10 trenches and subsequent bulk composite samples were excavated using an excavator to an average depth of 5 metres on the leach pad. Sampling was to test spent ore metallurgy for estimated recovery rates through the milling process. Assay results are presented in Table 10.4.

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Table 10.4: 2012 Trench and Bulk Composite Samples

Muestra	Ley de Cabeza
	Au g/ton	Ag g/ton	Cu G/ton
Patios 1 Gastado	0.75	33.60	43.00
Patios 2 Gastado	0.22	14.20	28.10
Patios 3 Gastado	0.86	32.83	53.35
Patios 4 Gastado	0.64	33.60	46.18
Patios 5 Gastado	2.17	77.40	64.50
Patios 6 Gastado	1.02	33.23	43.18
Patios 7 Gastado	0.57	19.80	38.50
Patios 8 Gastado	2.50	68.90	59.03
Patios 9 Gastado	0.43	40.55	51.68
Patios 10 Gastado	0.36	24.20	48.50

10.3.1   Heap Leach Drilling

In early May 2013, SilverCrest completed 25 large diameter drill holes, totalling 359 metres, on the current leach pad for the purposes of validating production data (Figure 10.2).  This drill program was completed by Layne Drilling, a Mexican company. Drill equipment used to complete the drill program was a reverse circulation Becker Hammer with inner tube of 6’’ for sampling. Drill holes were spaced on a 25 by 25 metre grid, and drilled to 5 to 20 metres deep. Enough material and space was left between the collar of the hole and the liner to prevent any damage to the liner, and were stopped several meters above the surveyed location of the liner. All drilling operations were performed without the addition of water or other additives and full sample was collected per each interval drilled without any rejects on the drill site. A total of 164 samples were collected for analysis.

Figure 10.2: Drilling setup on the Heap Leach Pad

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10.4   Core Recovery

Core recovery for the 2012-13 drilling program was measured in relation to marker blocks and was generally good (Figure 10.3). RQD was also measured and recorded during core logging (Figure 10.3). The holes include SE-12-39 to SE-13-150 and SERC-12-22 to SERC-12-35.

Figure 10.3: Core Recovery and RQD of Santa Elena 2012-2013 Drilling

10.5   Drill Surveying and Topography

Surveying of drill hole collars was first completed by handheld GPS and later resurveyed by the registered on site surveyor.

Eagle Mapping of Vancouver B.C. completed an aerial survey of the property in 2007.  The resultant digital elevation model from this survey provides the basis for the site topography for the project.  Where previous drill collars were not located by the surveyor, their elevation in the database has been manually fit to the digital elevation model.

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11.0   SAMPLING METHODS AND APPROACH

11.1   Historical Sample Collection Methods

Knowledge of the sampling methods for work completed prior to 2004 is limited.  No data from this period has been used in previous NI-43 101 Technical Reports as sampling completed by Tungsteno and Nevada Pacific was inadequately documented and insufficient to determine the approach.

In 2004, Fronteer completed surface and underground sampling at Santa Elena. Written documentation on sampling methods is limited but investigations by SilverCrest confirmed the underground sample locations.  The sampling method was found to be of contradictory “discontinuous” channels and this approach was considered adequate at the early exploration stage.

11.2   Previous SilverCrest Sample Collection Methods

The 2006 sampling by SilverCrest consisted of continuous surface channel sampling along exposed road cuts and outcrops. The underground verification channel sampling program consisted of semi-continuous horizontal sampling of identified Fronteer sample locations.  The samples were collected over selected intervals, placed in plastic bags and periodically shipped to ALS-Chemex in Hermosillo Mexico for preparation, with sample pulps shipped to and analysed by ALS-Chemex, North Vancouver, BC.

The 2006, 2007 and 2008 core drilling procedure included the collection and labelling of the drill core. After logging and identifying the mineralized zone, core was selected for splitting and sampling. The 2008 RC drilling program consisted of collecting chips and cataloguing.

11.3   2012-2013 Sample Collection Methods

The 2012 and 2013 drilling program included procedures for the collection and labelling of the drill core.  A total of 15 drill holes (Table 11.1) where first drilled by RC methods and finished with diamond core tails with a further 4 drilled purely as RC of HQ size drill core (63.5 mm diameter).  Although RC cuttings were not retained a number of samples from the hanging wall were sampled. (Figure 11.1)

Table 11.1: Table showing RC Cutting Sample Length (From ‘SERC’- series drill holes) and frequency from the hanging wall

Length (m)	Frequency
0.5	0
1	9
1.5	12
2	91
2.5	11
3	490
3.5	0
3.5+	1
Total	614

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The drill core was recovered and stored in vinyl boxes, each of which contains approximately 2.25 metres of core.  Drill runs were identified in the field by drillers using markers in the core boxes at 3 metre intervals.  These intervals were validated by SilverCrest geologists.  Recovered drill core was boxed by the drillers on-site. The core boxes were collected and delivered twice daily to the on-site core logging facility where the core was logged and sampled by SilverCrest technical staff.  Core is currently stored onsite for future viewing and reference.

Core logging procedures included review of the core quality and recording of recovery, lithological, geotechnical and mineralogical data within standardized company logging forms. After characterizing the mineralization, SilverCrest geologists marked the start and end of each interval for sampling. The range in drill core sample lengths can be seen in Figure 11.1 which shows lengths ranging from 0.11 to 36.7 metres (the latter was checked in supplied drill logs as being correct) and mode of approximately 2 metres. Not all drill holes were entirely sampled and only samples are presented in Figure 11.1 below.  This did not include RC sample intervals.  The average sample length used in the 2013 resource is 1.74 metres.

Figure 11.1: Santa Elena Sample Length Frequency used in 2013 Resource Estimate

Sample intervals were recorded on the core box with sample tags. The intervals were marked on the drill core which was cut in half by a SilverCrest technician using a diamond saw blade.  Half of the core was sealed in a sample bag with the corresponding sample tag.  The other half of the core sample was returned to the core box for company record and future viewing.  Sample numbers, intervals, and descriptions were recorded on the standardized drill logs.

SilverCrest inserted Certified Reference Material (CRM’s), blanks and duplicates samples at regular intervals into the sampling stream.  In addition internal laboratory QA-QC procedures were followed.

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11.4   2012-2013 Laboratory Analytical Methods

For the 2012-2013 sampling, two independent analytical laboratories were used for sample analyses: Nusantara de Mexico S.A. de C.V (Nusantara), an on-site grade control laboratory for Santa Elena operations; and ALS Chemex. Nusantara either prepared and analysed samples, or prepared and transported samples to ALS Chemex in Chihuahua or Hermosillo for further preparation before being sent to ALS-Chemex in Vancouver for analysis.

EBA notes that some pulp samples from the Santa Elena on-site lab were received by the ALS preparation labs.

11.4.1   Nusantara Lab Preparation and Analysis

Nusantara analysed a number of samples using Gravimetric Fire analysis and Atomic Absorption Fire Assay for Au and Ag.  The following sample preparation procedures have been translated from Spanish and are summarized for this report.

Received samples are dried, weighed and crushed to 90% passing the minus 150 mesh. Gold is analyzed using fire assay fusion with atomic absorption finish and by gravimetric methods.  Silver is analyzed using an aqua regia digestion and atomic absorption finish.

Nusantara’s internal QA-QC includes adding one duplicate, one reference and one blank to every 27 samples. All samples analysed by AAS outside the range of calibration are retested.  There are measures in place for oven time and contamination by chemical additives.

Table 11.2: Number and Preparation Method of Samples sent to ALS-Chemex from Santa Elena Mine

Sample Type	No. of Samples
Drill Core	4,884
Pulp	2,985
RC Cuttings	426
Total	8,295

Table 11.2 outlines the number and form of samples transported to ALS-Chemex from Nusantara for analysis.  A total of 2,985 pulp samples are recorded as having been received by ALS-Chemex in Chihuahua or Hermosillo.

11.4.2   ALS-Chemex Lab Preparation and Analysis

ALS-Chemex labs in Chihuahua and Hermosillo prepared the samples using the Prep-31 procedure typically used for rock and drill core.  The samples were dried, weighed, and crushed to obtain 70 % passing a 2 mm (Tyler 9 mesh, US Std. No.10) screen.  A split of up to 250 g was taken and pulverized to >85 % passing a 75 micron (Tyler 200 mesh, US Std. No. 200) screen.

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ALS-Chemex in Vancouver, BC, analysed samples using the ME-ICP41 analytical procedure.  Samples were digested with aqua regia in a graphite heating block.  After cooling the resulting solution was diluted to 12.5 mL with deionized water, mixed and analyzed by inductively coupled plasma-atomic emission spectrometry (ICP-AES) and reported 36 element grades including silver (ME-ICP41).  The analytical results are corrected for inter-element spectral interferences.

Samples containing grades of silver exceeding 100 ppm were redigested in aqua regia and run using ore grade detection limits with ICP (Ag-OG46), and where grades exceeded the detection limit 1,500 ppm, silver was analyzed using fire assay (Ag-GRA21) with AA finish.

Gold was analyzed using fire assay fusion with AA finish (Au-AA23) and default overlimit analysis using gravimetric determination (Au-GRA21).  A nominal weight 30+/-3 grams was used for the analysis.  Silver analysed by the GRA21 finish has a lower detection limit of 5 ppm to an upper limit of 10,000 ppm with gold having a lower detection limit of 0.05 ppm and an upper limit of 1000 ppm.

No ALS laboratory SG testing was undertaken.

11.5   Heap Leach Sampling, Preparation and Analysis

All sampling was carried out by SilverCrest’s geologists and sampling protocols adopted the following procedures:

§   Plastic bags were placed in a tray in the vertical outlet of the cyclone and into a container to avoid loss of material (Figure 11.2)

§    Full interval was sampled and samples were taken at multiple orders according to the depth of the hole. For holes with a length of 10 and 20 metres, samples were taken every 2 metres. Holes with length of 15 metres, samples were collected every 3 metres and only one 5 metre sample was collected for holes with 5m length.

§   All bags were labelled with the corresponding depth.

§   The samples were delivered to the Santa Site lab for splitting to pulverization and additional splitting to generate aliquot for analysis.

A very meticulous sampling procedure allowed for a very good sample recovery. All samples were handled by geologists at Santa Elena site. Samples were sent to Santa Elena lab for analysis. Analytical method for gold included Fire Assay finishing in AA as well as gravimetric analysis for comparison purposes and for silver an Aqua Regia digestion finishing in AA. Blanks and Certified Reference Material (“CRM”) were inserted by exploration personnel prior to the sampling preparation at the Santa Elena Lab to carry out a QA/QC protocol in the preparation and analysis of the samples collected by the drilling program on the pad. The results did not indicate deviations from the blanks and CRM assay values.

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Figure 11.2: Heap Leach Sampling

11.6   QP Statement

EBA has observed the core and RC sampling methods on site and have reviewed the leach material sampling protocol and believes the methods meet an industry standard of practice.  EBA feels the sample collection methods were carried out in a methodical and unbiased method sufficient to support a database for estimation of Mineral Resources and Reserves.

12.0   DATA VERIFICATION

12.1   Historical Data Verification

Records for historical accounts of data verification and quality control measures were not available to EBA.  No sampling material remains from the programs, as such, no verification sampling was undertaken.  Historical data prior to the 2006 SilverCrest drilling campaign is not included in the current geological database.

12.2   Previous SilverCrest Data Verification

The following description of historical verification has been summarised from the Technical Report on the Pre-Feasibility Study for the Santa Elena Project, Sonoro Mexico, by Scott Wilson Roscoe Postle Associates, (SWRPA, August 11, 2008) and the Technical Report on the Cruz de Mayo Property, Sonora, Mexico, by N. Eric Fier and C. Stewart Wallis (December 10, 2007). These technical reports were independent and NI 43-101 compliant.

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In April of 2006, SWRPA collected select samples for verification, including an underground continuous channel sample and quarter splits of drill core and sent to ALS Chemex in Hermosillo with a regular shipment of core samples. Overall, the grade comparisons are considered to be within acceptable ranges.

In May 2006, SilverCrest collected 15 underground channel samples to verify the sampling results of Fronteer. The SilverCrest silver assays showed a bias of up to 60% lower than the corresponding Fronteer assays with gold assays 50% to 100% lower. SWRPA considered it acceptable at this stage of property development to use the Fronteer data in the resource estimate.

Gravimetric silver grades were consistently higher compared to both Fronteer and the SilverCrest silver fire with AA finish results. The result lends support to the higher values. The fire assay with AA results was used in the resource estimate as they were more similar to the Fronteer results which were also used

In addition to the underground sampling, SilverCrest also completed silver geochemical analysis on 289 surface samples for fire assay AA finish and fire assay gravimetric analysis. Results show an overall 20.3% increase in silver grade using silver gravimetric assays. AA silver results were used in the resource estimation and are considered conservative for grade estimation. QA-QC was completed by duplicate analyses on 16 of the above 298 samples at ACME Laboratories in Vancouver on ALS-Chemex pulps from core sampling and preparation.  Although the ACME results have a higher detection limit, the limited results on the duplicate pulps show consistent correlation of grades between laboratories.

During the 2008 drilling approximately every 20th sample was duplicated in a different laboratory for QA/QC purposes. The comparison for 2008 drill sample results show average gold and silver results to be similar and within acceptable limits for QA/QC.

12.3   2012-2013 Data Verification

12.3.1   Site Visits

James Barr, P.Geo, visited the Santa Elena mine between May 10-11, 2012, and October 13-14, 2012 during the 2012-2013 drilling campaign at which time rock exposure in the open pit and exploration underground decline was inspected, sample collection and logging procedures were reviewed, verification samples were collected (Table 12.1, Figures 12.1 and 12,2) and recommendations for sampling quality control measures were made.

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Table 12.1: Drill Core Grade Verification Sampling (Oct. 2012)

Hole	From	To	Source	Samples	Au (g/t)	Ag (g/t)
SE-12-98	97.4	98.86	SVL	155228*	7.87	274
			EBA**	500427	6.07	281
SE-12-54	111.85	112.85	SVL	176561	10.75	233
			EBA**	500432	9.9	184
SE-12-39	158.34	159.55	SVL	176392	4.08	339
			EBA**	500428	4.42	296
CRM	CDN-ME-19		CDN	n/a	0.62 +/- 0.062	103 +/- 7
			EBA**	500430	0.673	109
SE-12-50	227.9	228.35	SVL	176937	4.29	918
			EBA**	500429	3.09	794
SERC-12-23	0.32	30.8	SVL	176251	0.32	30.8
			EBA**	500431	0.304	38.3
*Results reported from Nusantara lab
**EBA results reporting from ALS certificate HE12244041

Figure 12.1: Grade Verification Sampling, Silver

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Figure 12.2: Grade Verification Sampling, Gold

12.3.2   SilverCrest Certified Reference Material Insertion

Insertion of certified reference materials (CRM) at regular intervals was completed by SilverCrest staff during the 2012-2013 Santa Elena drill program. A list of the CRMs used by SilverCrest were provided to EBA along with a database of the insertion records, however, the exact CRM used was not noted in the database. Reported laboratory grades for the CRMs were plotted to make a broad association between the value and the CRM.  Ellipses in Figure 12.3 defines the ±2 standard deviations for Au and Ag, and the red dots represent certified mean values reported for each CRM. Blue dots represent returned laboratory values for the CRM insertions.

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Figure 12.3: Classification of Santa Elena CRM’s Material

Table 12.2 shows the count and mineralization values and ranges to ±2 standard deviations of the CRM’s used. An additional 10 “STD” in the database could not be related to known CRM material.  Table 12.2 outlines the standards that were sourced by SilverCrest from CDN Resource Laboratories Ltd in Vancouver and Rocklabs in New Zealand.

Table 12.2: Certified Reference Material Reporting Values used in the Santa Elena 2012-2013 Drilling by SilverCrest

CRM	Au gpt	Ag gpt	Count	Source
CDN-CM-17	1.37+/-0.13	14.4 +/- 1.4*	56	CDN Laboratories
CDN-ME-5	1.07+/-0.14	206.1+/-13.1**	21	CDN Laboratories
CDN-CM-19	0.620+/-0.062	103+/-7	18	CDN Laboratories
OxD 87	0.417+/-0.026	Not Reported	9	Rocklabs

*4 Acid with ICP or AA

**FA with Gravimetric

Samples identified as CRM are plotted (light blue) in Figures 12.4 to Figure 12.10 below relative to the mean (red line) and ±2 standard deviations (purple lines) reported by CDN Laboratories for each CRM.

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Figure 12.4: CRM: CDN-CM-17 - Gold

Figure 12.5: CRM: CDN-CM-17 – Silver

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Figure 12.6: CRM: CDN-ME-19 – Gold

Figure 12.7: CRM: CDN-ME-19 - Silver

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Figure 12.8: CRM: CDN-ME-5 – Gold

Figure 12.9: CRM: CDN-ME-5 - Silver

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Figure 12.10: CRM: OxD87 - Gold

12.4   SilverCrest Blank Material Insertions

SilverCrest inserted 114 blank samples in a random fashion and near to expected high grade samples during the 2012-2013 drilling program, each blank was labelled “Blank” or “Blanco” in the drill hole data base.  The material was collected from barren exploration drill core collected from unaltered andesite in the hanging wall to the vein.

Figures 12.11 and 12.12 show the 114 samples of blank material plotted with the +1 standard deviation (dashed) and lower detection limits (red line) of 0.005 (Au) and 0.2 (Ag).

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Figure 12.11: SilverCrest 2012-2013 Reported Blank Material Au (gpt)

Figure 12.12: SilverCrest 2012-2013 Reported Blank Material Ag (gpt)

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Blank sample results show a range of results (0.005 to 0.11 gpt Au and 0.2 to 13.5 gpt Ag) which can most likely attributed to contamination of the source material. The majority of blank samples are near the limit of detection for both Au and Ag.

EBA feels it is unlikely that the results indicate contamination of the samples occurred on site or during laboratory sample preparation.

12.5   ALS-Chemex and Nusantara Laboratory Comparison

For verification of Santa Elena results against ALS-Chemex 751 samples were sampled at both Santa Elena and ALS-Chemex from the 2012-2013 Santa Elena drilling (Section 10.3).  Samples analysed at Santa Elena prior to sending to ALS Chemex facilities in Hermosillo or Chihuahua were at the drill core, pulp or RC cutting stage of preparation (Section 11).

Samples up to 250 grams of pulverized material, with 85% passed a 75 micron screen, were transferred internally to ALS-Chemex in Vancouver, BC for analysis.

A total of 175 samples within the resource are using results analysed from Santa Elena.

Comparison of the duplicate analytical results are plotted on scatter plots to observe any bias/trends between the two labs (Figures 12.13 and 12.14). Table 12.3 shows the key statistics of the comparison results.

Figure 12.13: Au (gpt) Comparison between Santa Elena and ALS-Chemex

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Figure 12.14: Ag (gpt) Comparison between Santa Elena and ALS-Chemex

Table 12.3: Key Statistic Comparison of Au and Ag for Nusantara and ALS-Chemex.

	Au (gpt)		Ag (gpt)
	Nusantara	ALS	Nusantara	ALS
Mean	0.46	0.49	29.52	30.38
Standard Deviation	1.64	1.92	81.51	94.64
Minimum	0.00	0.01	0.00	0.20
Maximum	20.37	25.60	1,453	1,840

Reasonable reproducibility from the Nusantara lab to the ALS lab is indicated by the plots above.  Some gold grades produce a potential slight positive bias to ALS at high grades. Silver grade reproduction is highly variable but overall seems to co-relate with an even spread across the 1:1 unity line (red). Overall, the data reported from the Nusantara lab appears to be reproducible at the certified ALS lab and sample data reported from this source are considered to be reliable.

12.6   Heap Leach Data Verification

Blanks (Table 12.4) and Certified Reference Material (“CRM”) (Figures 12.15 and 12.16) were inserted during the sampling preparation at the Santa Elena Lab to monitor and have a QA/QC protocol in the sampling and analysis of the samples collected by the drilling program on the pad.  The results did not indicate deviations from the blanks and CRM assay values reported in the blanks and CRM assays certificates (Table 12.5).

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Table 12.4: Heap Leach Blank Samples Results

Hole ID	Sample No.	Au (gpt)	Ag (gpt)
LP-13-01	LP 13 01 11 BCO	0.02	0.44
LP-13-02	LP 13 02 11 BCO	0.02	0.53
LP-13-03	LP 13 03 16 BCO	0.03	1.10
LP-13-04	LP 13 04 11 BCO	0.02	0.36
LP-13-05	LP 13 05 16 BCO	0.02	0.75
LP-13-06	LP 13 06 16 BCO	0.02	0.66
LP-13-07	LP 13 07 21 BCO	0.01	2.00
LP-13-08	LP 13 08 21 BCO	0.01	0.19
LP-13-09	LP 13 09 21 BCO	0.03	0.04
LP-13-10	LP 13 10 21 BCO	0.01	1.54
LP-13-11	LP 13 11 21 BCO	0.02	0.21
LP-13-12	LP 13 12 21 BCO	0.02	0.32
LP-13-13	LP 13 13 21 BCO	0.02	0.18
LP-13-14	LP 13 14 21 BCO	0.01	0.23
LP-13-15	LP 13 15 21 BCO	0.03	0.08
LP-13-16	LP 13 16 21 BCO	0.01	0.85
LP-13-17	LP 13 16 21 BCO	0.01	0.30
LP-13-18	LP 13 16 21 BCO	0.02	0.55
LP-13-21	LP 13 16 21 BCO	0.01	0.68
LP-13-22	LP-13-22-01 BCO	0.01	0.54
LP-13-23	LP-13-23-01 BCO	0.01	0.65
LP-13-24	LP-13-24-01 BCO	0.02	0.66
LP-13-25	LP-13-25-01 BCO	0.02	1.47

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Figure 12.15: Distribution of CRM Au Grades from Leach Pad Drilling Insertions

Figure 12.16: Distribution of CRM Ag Grades from Leach Pad Drilling Insertions

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Table 12.5: Tabulated CRM Insertion Results for the Heap Leach Drilling

Hole ID	Sample No.	Au (gpt)	Ag (gpt)
LP-13-02	LP-13-02/1	1.39	26.25
	Standard CDN-ME-16	1.48 + - 0.14 g/t Au	30.8 +- 2.2 g/t Ag
LP-13-12	LP-13-12-15/1	1.16	27
	Standard CDN-ME-16	1.48 + - 0.14 g/t Au	30.8 +- 2.2 g/t Ag
LP-13-13	LP-13-13-9/1	0.58	98
	Standard CDN-ME-19	0.620 + - 0.62 gt/ton Au	103 + - 7 g/t Ag
LP-13-14	LP-13-14-11/1	0.7	90.5
	Standard CDN-ME-19	0.620 + - 0.62 gt/ton Au	103 + - 7 g/t Ag
LP-13-21	LP-13-21-17/1	0.60	86
	Standard CDN-ME-19	0.620 + - 0.62 gt/ton Au	103 + - 7 g/t Ag

13.0   MINERAL PROCESSING AND METALLURGICAL TESTWORK

13.1   Previous Mineral Processing and Metallurgical test work

The following sections were documented in the previous Technical Report for Santa Elena (May, 2011) and have been extracted from the Technical Report on the Pre-Feasibility Study for the Santa Elena Project, Sonoro, Mexico, by Scott Wilson Roscoe Postle and Associates (SWRPA) (August 11, 2008). This technical report is NI 43-101 compliant. All test work covers the Santa Elena project. No test work been done on the Cruz de Mayo project.

13.1.1   Previous Metallurgical Test Work Summary For Santa Elena

There has been varied metallurgical test work done on the Santa Elena property over the last twenty five years. The test work was reviewed in 2008 by Mr. Geoff Allard (Allard), P.E., of Allard Engineering Services of Tucson, Arizona.

The Allard audit overall findings are summarized below:

§   No fatal flaws exist with regard to metallurgy,

§   Ore is amenable to heap leaching with cyanide,

§   Average overall recovery of 67% for gold and 34% for silver (varied for depth) is consistent with available test results and acceptable for Pre-Feasibility level,

§   Ore is suitable for anticipated design,

§   Top size of 3/8 inch for heap feed is appropriate,

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§   Test work does not adequately reflect proposed flow sheet and design criteria,

§   Significant additional test work is required to firm up design criteria for a feasibility level study,

§   There is inverse dependence of recovery on particle size,

§   Short column testing (less than proposed lift height of five metres) cannot address leach cycles and percolation issues, and

§   Some reagent consumptions cannot be identified from the existing test work. In these instances, estimates have been used.

§   As previously stated in their August 2008 PFS, SWRPA concur with Allard’s findings.

13.1.2   Previous Metallurgical Test Work Description

Various metallurgical tests have been carried out on the Santa Elena property. Relevancy of test work carried out during the period 1983 to 2003 was very difficult to ascertain, as either the work was not applicable or sample locations/references were not readily discernible. Table 13.1 below lists these reports and comments by Allard.

Table 13.1: Previous Metallurgical Reports and Comments

Metallurgical Reports	Comments
Western Testing Laboratories - October 1983	Column tests conducted on Santa Elena tailings, not applicable
Comision de Formento Minero of Hermosillo (CFM), September 1984	Column tests, sampling location unknown, supporting  information only
Western Testing Laboratories - June 1985	Sample reference unclear, rejected for inclusion
Comision de Formento Minero of Hermosillo (CFM), June 1985	Well documented column test but sample reference  unclear, supporting information only
Comision de Formento Minero of Hermosillo (CFM), September 1986	Well documented bottle roll tests, column test and flotation test but sample reference unclear, supporting information only
Universidad de Sonora - February 2003	Bottle roll and column tests, minimal information on test procedures and sample location, supporting information only

Further test work started in 2006 and continued through February 2008. Summaries of the test work and Allard’s audit comments follow below.

In July 2006, Nusantara completed six bottle roll tests on representative samples collected from the Santa Elena Mine. This work was completed by Sol y Adobe Ingenieros Asociados S.A. de C.V. (Sol & Adobe) in Hermosillo, Mexico, in association with the University of Sonora. Results of a three-day leach at minus 10 mesh are shown in Table 13.2.

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Table 13.2: Santa Elena Bottle Roll Results

Sample No.	Calculated head Gold	Calculated head Silver	% Gold Recovery	% Silver Recovery	NaCN Consumption	NaOH Consumption
	g/t	g/t	%	%	kg/t	kg/t
1	0.73	39.37	75	58	0.8	4.3
2	0.39	19.42	67	43	0.5	3.3
3	4.29	139.51	73	53	1.74	5.9
4	2.49	79.5	77	20	1.1	4.8
5	0.81	68.58	73	48	0.7	2.9
6	2.96	45.03	69	31	1.3	11.9

The conclusions made by Sol & Adobe indicated an average 73% recovery for gold and 42% recovery for silver. Reagent consumption was modest at 1.0 kg/t of cyanide consumption. Lime consumption appeared to be high at 5.52 kg/t. Sol & Adobe recommended further bottle roll tests to optimize cyanide and lime consumptions on average grade composites (Sol & Adobe, 2006).

Allard made the following comments on this test work:

§   Poor correlation between assay head and calculated head,

§   Overstated lime consumption,

§   Some calculation errors.

Société Générale de Surveillance de México S.A. de C. V. (SGS) carried out bottle roll tests starting in 2007 as part of a column test campaign. Three ore composites developed from underground channel sampling from the Santa Elena Mine named S (surface), SL (second level) and FL (fourth level) were prepared, and tests were done on ground (-100 mesh) (Table 13.4) samples for 96 hours.

Table 13.3: Santa Elena Bottle Roll Results for Ground Composites

Bottle Roll Results for S, SL and FL Ground Composites
Sample ID	Test no	Gold Extraction	Silver Extraction	NaCN	Lime
		%	%	kg/t	kg/t
S	1	81.5	55	0.41	0.57
S	2	82.7	56.3	0.48	0.57
SL	1	92.3	58.1	0.45	0.56
SL	2	92.8	60	0.38	0.58
FL	1	94.6	39.6	0.34	0.6
FL	2	93.9	40.1	0.38	0.55

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Table 13.4: Santa Elena Bottle Roll Results for Coarse Composites

Bottle Roll Results for S, SL, and FL Coarse Composites
Sample  ID	Test  no	Particle Size Inches	Gold Extraction	Silver Extraction	NaCN	Lime
			%	%	kg/t	kg/t
S	1	1/4"	47.3	17.6	0.18	0.575
S	2	3/8"	43.6	15.1	0.19	0.601
S	3	1/2"	40.7	11.2	0.25	0.585
SL	1	1/4"	39.3	11.1	0.22	0.59
SL	2	3/8"	37	9.5	0.22	0.565
SL	3	1/2"	33.7	8.7	0.19	0.565
FL	1	1/4"	35.2	10.6	0.13	0.57
FL	2	3/8"	32.9	8.9	0.13	0.56
FL	3	1/2"	25.3	7.3	0.3	0.572

Cyanide and lime consumptions were reported as low. The gold and silver extractions were reported as acceptable. SGS recommended using higher cyanide concentrations, higher reaction time and better control in the production of fine particles while crushing, to improve the metals recoveries.

Allard comments that the bottle roll tests showed the dependence of recovery on particle size. Since the procedure to generate the coarse samples (as reported in the test work) may not be representative of a normally crushed sample, the tests can be used as an indicator of particle size dependence but should not be used for recovery projections.

Column testing was carried out by SGS on the composites using minus 3/8 inch material, except for one sample which was minus ¼ inch material. The results obtained after 77 days of leaching are summarized in Table 13.4. The column tests were run for 58 days, the ore was allowed to rest for 14 days, and then leaching commenced for another 19 days.

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Table 13.5: Santa Elena Column Test Results for Particle Size Composites

Column Test for Different Composites at a Particle Size of 3/8"
Sample ID	Gold Extraction	Silver Extraction	NaCN	Lime
	%	%	kg/t	kg/t
S-A	69.4	38.9	0.775	7.82
S-B	70.8	35.7	0.559	5.68
S (1/4")	72.1	36.9	0.577	6
SL-A	67.1	32.3	0.571	7.3
SL-B	69.2	29.4	0.588	6.8
FL-A	61.2	33.7	0.718	5.53
FL-B	57.7	33.3	0.601	7.66
Average	66.8	34.3	0.627	6.68

From these column tests, the gold extraction averaged 66.8% and silver extraction, 34.3%. Cyanide consumptions are moderate, averaging 0.627 kg/t ore, while lime consumption averaged 6.68 kg/t ore.

Allard made the following comments on the column tests:

§   Circulation of leach liquor through columns without cyanide to “neutralize” ore is poor practice,

§   Increase of cyanide concentration after 11 days from 0.3 g/L to1.0 g/L did not improve leaching kinetics as stated in the SGS report,

§   Columns operated in closed circuit without recovering gold or silver from column effluent. This affected dissolution and diffusion,

§   pH was maintained at abnormally high range,

§   Test column height (1.71 m) was inadequate to determine leach cycle for proposed 5 m lifts, and

§   Tests can be used for recovery indications but questionable test aspects would tend to reduce recovery.

Subsequent test work, consisting of grinding to 75% minus 75 micron and leaching for 120 hours, was done on the samples of the column test tails from the previous tests. As well, work indices were determined for the composites. The results after 120 hours of leaching are shown below in Table 13.6.

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Table 13.6: Santa Elena Grinding/Leaching Test Work

Grinding/Leaching Tests on Column Test Work Tails – Santa Elena
Sample ID	Gold Extraction	Silver Extraction	NaCN	Lime
	%	%	kg/t	kg/t
S-A	95.1	67.35	1.05	0.874
S-B	94.8	65.8	1	0.839
SL-A	88.1	59.1	1	0.839
SL-B	85.7	62.2	0.85	0.864
FL-A	78	61.5	1.25	0.92
FL-B	81.7	59.7	1.2	0.89
Average	87.2	62.6	1.06	0.871

The results of test work on the previous column test tails returned average gold and silver recoveries of 87.2% and 62.6%, respectively. Allard commented that cyanide and lime consumptions were high at 1.06 kg/t and 0.87 kg/t, respectively.

The crushing work indices are summarized below in Table 13.7.

Table 13.7: Crushing Work Indices

Composite	Work Index kWh/t
S-A	14.65
S-B	14.56
SL-A	14.04
SL-B	13.75
FL-A	14.11
FL-B	14.3

Allard commented on the consistency of the indices between samples.

There was one test done to determine specific gravities and bulk densities that has not been audited by Allard, but is included here. In September 2007, SGS determined crushing work indices, specific gravities and bulk densities of three composites from bulk sampling: S = Surface, SL = Second Level and FL = Fourth Level (SGS, 2007). There is no record of procedures. The results are shown in Table 13.8.

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Table 13.8: Composites Work Indices, Specific Gravities and Bulk Densities

Composite	Work Index	Specific Gravity	Bulk Density
	kWhr/t	t/m3	t/m3
S	14.77	2.68	2.49
SL	14.77	2.65	2.35
FL	14.7	2.69	2.47

13.2   Recent Mineral Processing and Metallurgical test work

Appendix C contains a metallurgical report of test work carried out by Inspectorate  Mining and Metals in their Richmond, BC facility on samples from Santa Elena and a limited number of samples from Cruz de Mayo.  Inspectorate also generated slurry samples for testing at Pocock Industrial in Salt Lake City for thickening and filtration characterization.  Additional test work was carried out in Sonora at the University of Sonora.

The broad conclusions from the Inspectorate test work can be summarized as follows.

§   A grind of 80% passing (P80)100 µm is suitable for treating Santa Elena ore.

§   A cyanide strength of 1g/L is suitable for leaching Santa Elena ore

§   A leach time of 72 hours is indicated by the standard test procedure but the leaching rate can be accelerated significantly if the leach solution is changed after an initial period of 10 hours and shortened to under 40 hours as well as improving silver recoveries.

§   A gold recovery of >92% and silver recovery of 68% is indicated.

§   Ore from Cruz de Mayo is harder and requires a finer grind than Santa Elena ore, and still gives poorer results.

Additional findings from recent test work is a Bond Ball Mill (metric) work index of 14.  The filtration test work indicated that filtration of Santa Elena tailings was highly susceptible to the flocculent selected.

Open pit heap leach metallurgical test work is based on previous work and approximately 3 years of production data which has been publicly announced and filed on SEDAR. Overall, actual average heap leach recoveries are estimated for ore under a 300 day leach cycle at 57% gold and 30% silver. Since the leach pad will be decommissioned pre-maturely (before 300 day cycle is completed for most of 2013 ore placement), the estimated final open pit heap leach recoveries are 52% gold and 29% silver.

13.2.1   Leaching Cyanidation of Samples from Santa Elena Deposit

Bottle roll cyanide leaching was mainly performed on Santa Elena NIVEL 4 whole ore at various grind sizes and different cyanide dosages to assess gold and silver recovery potentials.

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Baseline Bottle Roll Tests on Ground Whole Ore (Underground)

Baseline bottle roll tests (C1 to C6) were conducted on ground whole ore to assess the grind- and cyanide-sensitivity of gold and silver recoveries as well as kinetics and reagent requirements.  The leach tests were carried out for 72 hours at 40% solids at target P80 grinds of 120 mm and 74 mm. Three cyanide dosages of 0.5-, 1.0- and 1.5 g/L cyanide were tested at each grind, and pH ≥10.0 was maintained with hydrated lime.

Composites C1 to C12 were collected as bulk samples on the 4th level (est. 80 m below surface) of the old underground working which would be near the bottom of the operating open pit. The baseline test results are summarized in Table 13.9, whilst Au and Ag leach profiles are plotted in Figures 13.1 and 13.2.

Table 13.9: Baseline Cyanidation Performance

	Test Conditions		Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
Test
No	P80 Size,	NaCN,	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
	um	g/L
C1	121	0.5	3.31	119.1	2.66	113.0	94.2	42.4	0.16	65.10	0.78	0.08
C2	119	1.0	3.31	119.1	2.98	128.9	96.1	59.2	0.12	52.60	1.30	0.11
C3	119	1.5	3.31	119.1	2.78	125.9	95.5	64.5	0.12	44.70	1.55	0.10
C4	75	0.5	3.31	119.1	2.73	115.0	96.5	47.3	0.10	60.60	0.90	0.10
C5	73	1.0	3.31	119.1	3.01	125.7	94.2	61.3	0.17	48.70	1.29	0.07

As shown in Table 13.9, the gold responded much better to cyanidation than the silver in these exploratory baseline tests, averaging 95.6% Au versus 56.6% Ag. In general, residue gold and silver grades decreased with finer grinding and higher cyanide concentration.  In the range of 0.5-1.5 g/L cyanide, higher cyanide strength benefited silver extraction, but not gold.

Cyanide consumption was in the neighborhood of 1.3 kg/t at a level of 1 g/L NaCN. In general, less than 0.1 kg/t lime were required to maintain slurry pH ≥10.  Increasing the cyanide strength from 0.5 g/L to 1.5 g/L doubled the NaCN consumption.

Leach kinetics showed that both gold and silver dissolution was fast in the first 8 hours of retention and subsequently slowed down. Finer grind and higher cyanidation strength resulted in faster metal dissolution and better extraction. Gold and silver extractions might improve by extending the residence time to over 72 hours without replacing leach liquor.

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Figure 13.1: Gold Leach Kinetics on Milled NIVEL 4 Comp.

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Figure 13.2: Silver Leach Kinetics on Milled NIVEL 4 Comp.

Leach Optimization

The optimization tests (C7 to C12) were conducted on the NIVEL (Level) 4 Composite at a moderate grind of 100 µm at various cyanide dosages with leach solution replacement at various leach intervals.  Test conditions and leach results are summarized in Table 13.12, while leach profiles are plotted in Figures13.3 and 13.4.

In comparison to the baseline test results, leaching with solution replacement resulted in better silver extraction, and faster gold and silver leach kinetics.  By replacing the leach solution, up to 70% silver recovery is achievable, and residence time can be shortened to 30 hours for gold. Double the number of solution exchanges and NaCN dosages as well as the addition of oxidants resulted in negligible gains in gold and silver extractions. In general, approximately 96% Au and 70% Ag extraction can be achieved at 40% solids in 1.0g/L NaCN solution with single solution replacement at the target grind of 100 µm.

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Table 13.10: Leach Optimization Test Summary

Test No		Test Conditions		Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
	NaCN, g/L	Oxidant	Sol'n Replacement	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C7	1.0		@ 24hrs	3.31	119.1	2.58	131.8	95.6	67.6	0.11	42.7	1.23	0.17
C8	1.0	0.25kg/t Pb(NO3)2	@ 24hrs	3.31	119.1	2.63	126.9	95.8	70.3	0.11	37.7	1.31	0.16
C9	1.0	0.1kg/t H2O2	@ 24hrs	3.31	119.1	2.80	133.7	95.5	67.9	0.13	42.9	1.31	0.15
C10	1.5		@ 24hrs	3.31	119.1	2.87	135.7	96.0	70.2	0.11	40.4	1.24	0.16
C11	1.5		@ 14- & 24hrs	3.31	119.1	3.00	125.9	96.3	67.0	0.11	41.6	1.97	0.12
C12	2.0		@ 14- & 24hrs	3.31	119.1	2.85	127.6	96.5	69.2	0.10	39.3	2.68	0.16

Figure 13.3: Optimization Gold Extraction Kinetics

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Figure 13.4: Optimization Silver Extraction Kinetics

Confirmatory Leach on New Santa Elena Feed

Confirmatory leach tests (C15 to C17) were carried out on another Santa Elena NIVEL (Level) composite (Comp. 4 ± 0 Nivel) with mine water (also referred as site water in the test spreadsheets) provided by the client. Leach pulp density increased from 40 wt.% for the baseline tests to 45%. During the leach tests, the pregnant leach solution was replaced with fresh mine water after 10 hours leach. Leach conditions and results are summarized in Table 13.12.

Results showed that over 97% gold extraction is achievable by leaching the Comp. 4 ± 0 Nivel in 1.0g/L cyanide at 45% solids by replacing partial or full leach liquor. Increasing cyanide dosage to 1.5g/L did not result in significant improvement of gold recovery. Silver extraction was ~43% in all three cases.

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Table 13.11: Confirmatory Test Comp. 4 +/- 0 Nivel

Test No		Test Conditions		Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
	NaCN, g/L	Oxidant	Sol'n Replacement	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C7	1.0		@ 24hrs	3.31	119.1	2.58	131.8	95.6	67.6	0.11	42.7	1.23	0.17
C8	1.0	0.25kg/t Pb(NO3)2	@ 24hrs	3.31	119.1	2.63	126.9	95.8	70.3	0.11	37.7	1.31	0.16
C9	1.0	0.1kg/t H2O2	@ 24hrs	3.31	119.1	2.80	133.7	95.5	67.9	0.13	42.9	1.31	0.15
C10	1.5		@ 24hrs	3.31	119.1	2.87	135.7	96.0	70.2	0.11	40.4	1.24	0.16
C11	1.5		@ 14- & 24hrs	3.31	119.1	3.00	125.9	96.3	67.0	0.11	41.6	1.97	0.12
C12	2.0		@ 14- & 24hrs	3.31	119.1	2.85	127.6	96.5	69.2	0.10	39.3	2.68	0.16

Figure 13.5: Confirmatory Extraction Kinetics

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Figure 13.6: Confirmatory Silver Extraction Kinetics

Baseline Bottle Roll Tests on Spent Ore

During 2012, 10 representative bulk samples from trenching were collected from spent ore on the leach pad and analysed using mill design specifications. Analyses consisted of bottle roll leach tests for 72 hours at approximately 33.3% solids at target P80 grinds of 74 µm. The cyanide dosage of 1.0 g/L cyanide was tested at a pH ≥10.0 maintained with hydrated lime. Test work was completed at the on-site Santa Elena lab which has a fully functional wet lab with extensive external QA/QC reviews (refer to Section 12 – Data Verification).

Results in Table 13.11 show average recovery rates of 91.4% gold and 72.5% silver which is consistent with results tested on underground bulk samples and bottle roll tests used for mill design. Recovery rates of 92% gold and 67.5% silver are used for all milling ores including spent ore from the leach pad.

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Table 13.12: Average Recovery Rates for Underground Bulk Samples

Muestra	Recuperacion			NaCN Consumido kg/ton	Cal Agregada kg/ton
	Au %	Ag %	Cu %
PR-Patios 1 Gastado	89.01	74.21	26.60	0.38	1.30
PR-Patios 2 Gastado	84.10	67.43	22.60	0.56	1.40
PR-Patios 3 Gastado	91.90	75.36	13.93	0.44	1.50
PR-Patios 4 Gastado	94.40	86.32	18.92	0.74	1.20
PR- Patios 5 Gastado	96.19	85.78	28.71	0.86	1.70
PR-Patios 6 Gastado	92.44	54.71	10.85	0.44	1.10
PR-Patios 7 Gastado	94.82	65.39	20.09	0.78	1.60
PR-Patios 8 Gastado	93.25	84.00	24.73	0.58	1.00
PR-Patios 9 Gastado	91.88	66.93	22.42	0.64	1.50
PR-Patios 10 Gastado	86.32	65.71	18.89	0.56	1.20

Three additional tests were completed on composites 4, 6 and 8 (Table 13.12) to check for reproducibility and graphically report results in Figures 13.7 to 13.9.

Table 13.13: Results of Additional Testing on Samples 4, 6 and 8.

Muestra	Recuperacion			NaCN Consumido kg/ton	Cal Agregada kg/ton
	Au %	Ag %	Cu %
PR-Patios 4 Gastado-Reemplazo	94.89	87.24	21.36	1.06	1.60
PR-Patios 6 Gastado-Reemplazo	93.32	56.27	12.59	0.74	1.50
PR-Patios 8 Gastado-Reemplazo	92.66	84.79	27.46	0.74	1.60

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Figure 13.7. Additional Testing of Composite Sample #4.

Figure 13.8. Additional Testing of Composite Sample #6.

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Figure 13.9. Additional Testing of Composite Sample #8.

Leach Tailings Production

All remaining Santa Elena feed samples, including 20kg of Composite Nivel 4 and 5.5kg of Comp. 4 ± 0 Nivel, were leached to produce sufficient tailing samples for thickening studies at Pocock Industrial. The two feed samples were processed separately due to their difference in color. The leach tests (C18-C19) were carried out at 48% solids with synthetic mine water in 1.5g/L NaCN at a target P80 grind of 100 mm. The leach results are summarized in Table 13.13.

Table 13.14: Bulk Leach Summary

Test No	Sample ID	Test Conditions				Extraction		Residue Grade		Consumption (kg/t)
		P80 Size, µm	% Solids	NaCN, g/L	Retention	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C18	Comp. 4 ± 0 Nivel	97	48	1.5	20 hours	94.4	42.4	0.06	21.2	0.52	0.13
C19	Nivel 4 Comp.	111	48	1.5	8 hours	84.9	51.8	0.29	54.9	0.35	0.12

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13.2.1   Filtering and Thickening Test work

Pocock Industrial carried out a series of tests on leach residue produced by Inspectorate to determine design parameters for plant design.  A flocculent screening programme was carried out and it was determined that flocculent performance was highly variable.  The best performance was achieved using a Hychem Flocculent (N301) for both thickening and a filter aid.  Using this flocculent in thickening over 70% solids could be achieved from feed densities in the 20-25% solids range with an indicated thickener area (for a conventional thickener) of 0.11m2/tpd.  This same flocculent used as a filter aid could produce an 18% moisture cake at a filtration rate of 777kg/m2/hr.

14.0   MINERAL RESOURCE ESTIMATES

14.1   Introduction

Three individual Mineral Resource Estimates are presented in this report, a new estimate for spent ore material on the existing leach pads has been prepared by Eric Fier, Chief Operating Officer of SilverCrest and non-Independent Qualified Person, and an updated estimate for the material outside of, and below, the ultimate open pit has been prepared by James Barr, P.Geo, Senior Geologist with EBA and an Independent Qualified Person.  No resource estimate has been completed for the open pit at Santa Elena as all the material has previously been declared as mineral reserves, as discussed in Section 25 of this report.  A description of the methods used to define the open pit grade model is included below.

14.1.1   Previous Underground Resource Estimates

An initial resource estimate for underground resources at Santa Elena was completed by SilverCrest with the assistance of Dassault Systemes Software International and reviewed by EBA in January, 2011.  The resource was released by EBA in May 2011 as part of a Preliminary Economic Assessment in the report “NI 43-101 Technical Report, Reserve Update for the Santa Elena Open Pit and Preliminary Assessment for the Santa Elena and Cruz de Mayo Expansion Project, Sonora Mexico (amended date May 11, 2011)”.  The estimate was based on 146 exploration drill holes (totalling 20,970.4 metres) and 2,106 samples for modelling gold and silver into the block model using GEMS software.  The underground resources include all mineralized blocks outside of the final open pit1 and subjected to a 1.77 AuEQ2 gpt cut-off.

The underground Mineral Resources reported in 2011, and shown in Table 14.1 below, are superceded by the current resource estimate described in this report.

 

1 Note:  the final open pit used in the 2011 resource estimate was updated in the current report, the current ultimate pit is much smaller than previously. 

2 AuEQ from the 2011 resource estimate was based on $USD 1,000/oz Au and $USD 18/oz Ag, and metallurgical recoveries of 93.5% Au and 75.7% Ag.

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Table 14.1: Previous Santa Elena Underground Resources (EBA, May 2011)

Classification	Tonnes	Gold (gpt)	Silver (gpt)	Contained Gold (troy/oz)	Contained Silver (troy/oz)
Indicated	991,100	1.83	109	58,330	3,476,960
Inferred	1,487,900	1.53	87	92,470	5,250,190
* Based on a cut-off grade of 1.77 AuEQ using $USD 1,000/oz Au and $USD 18/oz Ag, and metallurgical recoveries of 93.5% Au and 75.7% Ag, and subject to grade capping of 400 gpt Ag and 12.32 gpt Au.

Previously reported in-pit Mineral Resources have been converted to Reserves and are superceded by the current report.  Current in-pit reserves are discussed in Section 15 of this report.

14.2   Basis of Current Resource Estimates

Interpretation used for the geological and grade modelling of the Santa Elena deposit was prepared by EBA and has incorporated information through discussions with SilverCrest personnel and interpretations used previously for the property.  The models used in this Mineral Resource Estimate have been completed by EBA using Dassault Systemes GEMS v6.5 software.

14.2.1   Gold Equivalent (AuEQ) Calculations and Assumptions

The Santa Elena deposit demonstrates mineralogical correlation between gold and silver concentrations.  Based on current metal prices, EBA applied criteria to define the AuEQ, defined below, to be applied for both interpretation and reporting purposes.

Inputs for metal prices were based on trailing 3-year and 5-year datasets of average monthly metal prices for gold and silver at the end of March 2013, based on the drill hole data cut-off date of March 15th, 2013. Figures 14.1 and 14.2 show the 3 year monthly average trends with the average composite values.  From the values shown in Table 14.2, a ratio of 70:1 silver to gold was selected for the AuEQ using the formula shown below.

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Figure 14.1: 3 Year Trailing Average Monthly Gold Price with 3 and 5 Year Composite Averages

Figure 14.2: 3 Year Trailing Average Monthly Silver Price with 3 and 5 Year Composite Averages

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Table 14.2: Metal Prices and Gold to Silver Ratio Estimates

	Gold* ($USD/oz)	Silver* ($USD/oz)	Au:Ag Ratio**
3-year	$1,531	$29	72
5-year	$1,292	$24	73
Estimated value			70:1
* As of end of March, 2013, when resource model being assembled ** Using metal recovery assumptions of 92% for Au and 67.5% for Ag

14.2.2   Specific Gravity used in the Resource Estimation

Routine in-situ specific gravity (SG) testing was conducted by SilverCrest personnel throughout the 2012-2013 drilling programs.  A total of 846 measurements from drill core were collected with a standard scale and bucket measuring apparatus. Results were tabulated in the corresponding drill hole logs.  Of these, 766 SG samples were labelled with a rockcode (Table 14.3) and have been used in average SG calculations.

A specific gravity value of 2.60 was selected for mineralized material based on the mean value attained for the quartz, and vein related, samples.

Table 14.3: SG Tally Rock Type, Rock Codes used in SG Analysis

Rock Type	Rockcode	Number
Andesite	And	416
Andesite Breccia	And Bx	57
Andesite Porphyry	And Porf	34
Andesite-Quartz	And Qz	70
Diorite	Dio	3
Porphyry-Quartz-Feldspar	PQF	62
Porphyry-Quartz-Feldspar + Quartz	PQF qtz	15
Quartz Vein	Qz vein	83
Rhyolite	Rhy	26
Total		766

SG samples have been grouped based on logged lithology from the Santa Elena database, averaged and plotted. Figure 14.3 shows the range (whiskers), ±1 standard deviation (box) and mean (x) of each series of SG results based on rock codes from Table 14.3.

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Figure 14.3: Santa Elena Specific Gravity Mean Value and ±1SD Ranges

Table 14.4 shows the average (mean) for each rockcode used in the Santa Elena SG analysis. Rockcodes are defined in Table 14.3. Values of 2.65 (tonnes/m3) for underground waste material and 2.60 for mineralized vein related material were selected for use in the geological and resource model.

Table 14.4: Santa Elena SG Analysis and Average SG per Rockcode

Rockcode	Number
And	2.66
And Bx	2.65
And Prof	2.63
And Qz	2.66
Dio	2.57
PQF	2.62
PQF qtz	2.55
Qz vein	2.61
Rhy	2.65
Average	2.62

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14.3   Geostatistics

The overall drill hole database for Santa Elena was reduced down to only include data within the interpreted mineralization and cut-off grade used for the Mineral Resource Estimate.  This reduced the total number of core samples in the database from 12,723 down to a value of 3,678.  Figures 14.4 and 14.5 shows the raw gold and silver grades with in the Santa Elena vein. Table 14.5 summarizes raw descriptive statistics for the dataset.

Table 14.5: Descriptive Statistics for Raw Samples within Vein

	Au	Ag
Number of samples	3,678	3,678
Minimum	0	0
Maximum	267	2,840
Mean (gpt)	1.88	112.46
Median (gpt)	0.70	58.60
Variance	51.77	35,033.8
Standard Deviation	7.20	187.17
Coefficient of Variation	3.82	1.66
97th percentile (gpt)	9.20	500.5
98th percentile (gpt)	11.17	634.5
99th percentile (gpt)	17.7	889

Figure 14.4: Log-histogram of Raw Gold Grades within Santa Elena Vein

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Figure 14.5: Log-histogram of Raw Silver Grades within Santa Elena Vein

14.3.1.1   Grade Compositing of Raw Samples

Raw samples were composited to one metre lengths based on the high frequency of this sampling interval, and to accommodate areas where the vein narrowed. Table 14.6 summarizes the resulting effect of compositing on the population.

Table 14.6: Descriptive Statistics for 1 metre Composite Samples within Vein

	Au	Ag	Au (capped)	Ag (capped)
Number of samples	2676	2676	2676	2676
Minimum	0	0	0	0
Maximum	213	2,047.41	12	600
Mean (gpt)	1.99	104.77	1.73	99.4
Median (gpt)	0.79	60.5	0.79	60.5
Variance	31.07	23,488	6.01	14,176
Standard Deviation	5.57	153.26	2.45	119.06
Coefficient of Variation	2.81	1.46	1.41	1.19
97th percentile (gpt)	9.76	432.25	9.76	432.25
98th percentile (gpt)	11.31	529.70	11.31	529.70
99th percentile (gpt)	23.38	940.68	12	600

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14.3.1.2   High Grade Capping

Inspection of the uncapped composite data indicates the populations are positively skewed with some anomalous outliers.  Histogram distributions and log-probability plots for gold and silver grade distributions were used for visual determination of anomalous values and as the basis of selection for high grade metal capping. High grade caps were applied to the composited samples prior to interpolation of the block model using values of 12 gpt for gold and 600 gpt for silver.  Figures 14.6 and 14.7 below highlight the capped gold and silver values on cumulative log-probability plots.

Figure 14.6: Cumulative Log-probability Graph for Composited Gold Grades

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Figure 14.7: Cumulative Log-probability Graph for Composited Silver Grades

14.3.1.3   Cut-off grade calculation

The calculation used to define the cut-off grade that was used in the model is:

This calculation gave a cut-off value of 1.4 gpt AuEQ for the grade model used in developing the grade solids.

14.4   Underground Mineral Resource Estimate

The underground Mineral Resource Estimate described in this section has been prepared for mineralized material outside of the ultimate pit design and located within the Main Mineralized Zone, or one of the sub-parallel Cholugo and Cholugo Dos zones.  The underground Resource Estimate has been completed by EBA to conform to the guidelines set forth by National Instrument 43-101 and incorporates terms as defined by the Canadian Institute of Mining, Metallurgy and Petroleum Standards on Mineral Resources and Reserves: Definition Standards and Best Practice Guidelines. The interpretation, assumptions and parameters used in preparing this estimate are described in the following sections.

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Drill core data was used as the basis for geological interpretation with cut-off grade and for the construction of mineralized solids at a 1.0 gpt AuEQ cut-off created as triangulations using Dassault Systemes GEMS v6.5 software. The analytical data was interpolated into a block model using the Ordinary Kriging (OK) method and was constrained by the mineralized solids.

The geological grade model has been built referencing the WGS 84 zone 12 coordinate system, the origins and bounds of the block model are listed in Table 14.7, no rotation has been applied. Topographic data used in the model was collected by Eagle Mapping in November of 2010 using a fixed wing aircraft to create a digital terrain model (DTM) accurate to 2 metre resolution. Drill hole collar elevations surveyed by SilverCrest personnel using a handheld GPS were generally 10-14 metres higher than the DTM surface. Elevations were later corrected using the mine survey equipment and personnel to match the Eagle Mapping DTM.

Table 14.7: Block Model Origin and Dimensions of the Santa Elena Model

Database	Direction	Origin	Block Size (m)	Number of Blocks	Dimension (m)	Rotation (deg)
Vein, Underground	East (rows)	580,850	5	245	1,225	0
	North (columns)	3,321,100	5	238	1,190	0
	Elevation (levels)	900	5	130	650	0

Note: origin is based on the UTM WGS 84 projection, and is located at the upper northwest corner of the model.

14.4.1   Drill Holes Used in Underground Mineral Resource

This 2013 resource estimate is based on the sampling from an additional 137 SilverCrest core holes (44,726.53, metres) drilled during 2012-2013. Figure 14.8 shows the total number of drill holes used the 2013 resource estimation.  Drill holes for resource were selected from 2006 up to the effective date of April 30th, 2013. Table 14.8 displays the number of drill holes available for resource estimation.

Table 14.8: Drill Holes (all years) used for 2013 Resource Calculations.

Drill Type	No. of Holes Drilled	Total Meters
Diamond	194	51,961.46
RC	48	4,934.29
RC with Diamond Tail	14	6,386.1
Total	256	63,281.85

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Figure 14.8: Location of Drill Holes used in 2013 Mineral Resource Update

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All drill intercepts used for resource estimation were core.

14.4.2   Grade Model Used in the Interpretation

Interpretation used for the geological grade model was prepared by EBA through discussions with SilverCrest personnel, independent field observations and review of documentation.  The block model and resulting resource estimate was completed by EBA using Dassault Systemes GEMS geological modelling software.

The Santa Elena deposit is a semi-continuous quartz and calcite filled brittle structure that strikes approximately 080 degrees and dips approximately 60 degrees to the south with a predominant south easterly directed plunge on mineralization with a rake of approximately 25 degrees.  Local variations in both the strike and dip in the vein orientation is seen across the vein and may be associated with local pinch and swelling in the mineralized thickness.

A grade solid was generated on cross and plan section using raw drill hole assay grades and a 1.00 gpt AuEQ3cut-off threshold within an interpreted mineralized solid.  Hard boundaries are suited to the nature of epithermal vein mineralization and directly from evidence seen in drill hole mineralization and sample results at Santa Elena.  Because the grade shell includes two splays in the deeper eastern section it was subdivided into three internal domains to reflect current interpretation of flexure and re-orientation of the mineralization along strike. The flexural zones connecting these domains were modelled separately to allow for smooth and continuous interpolation of metal grades within the mineralization.  In the footwall of the zone, two additional narrow zones identified at mid to shallow depths directly beneath the current open pit operation were modelled and named Cholugo and Cholugo Dos (Figure 14.10).  A unique code was applied to each domain for modelling purposes, described in Table 14.9 below.

 

3 Calculated using an Au:Ag ratio of 70:1 as described in Section 14.2.4.

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Table 14.9: Description of Mineralized Domains for Santa Elena

Block Model Geological	Block Model Rockcode	Description
DORE PIT	700	Main Mineralized Zone (MMZ) material within ultimate pit
DORE_W	703	MMZ; easterly striking and southerly dipping. Western domain.
DORE_FLD	705	MMZ; easterly striking and southerly dipping mineralization in inflexion locations (x3)
DORE_M	702	MMZ; easterly striking and southerly dipping. Middle domain.
DORE_C	703	MMZ; easterly striking and southerly dipping. Central domain.
DORE_E	704	MMZ; easterly striking and southerly dipping. Eastern domain.
V_EASTFW	721	Splay occurring on the footwall of the MMZ.
V_WESTHW	720	Splay occurring on the hanging wall of the MMZ.
DORE_CL	706	Steeply dipping vein converging with the MMZ at mid-level below pit with limited (data based) strike length.
DORE_CL2	707	Steeply dipping vein converging with the MMZ at mid-level below pit with limited (data based) strike length.

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Figure 14.9: Geological Model showing Vein Domains, looking north

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Figure 14.10: Geological Model showing Vein Domains, looking ESE

The composite intervals described in Section 14.3 were calculated within the boundaries of the modelled grade shell using raw assay data, and then subjected to additional geostatistics and analysis as described below.

14.4.3   Interpolation Method and Variography

The model is based on verified drilling completed by SilverCrest between 2006 and 2013 which has provided near continuous drill intersection at approximately 40 metres spacing within the mineralization.  Grade distribution and the drill sample density supports variography analysis to estimate the orientation and anisotropy of elliptical search parameters for modeling the mineralization.

Initially, variograms were created for both gold and silver in each of the 3 subdivided domains to determine internal ellipse orientations and then for the overall vein to determine a reasonable search range.  The variograms were verified with field observation and matched well with expected grade trends within the Main Mineralized Zone.  Slight variance was noted between the ranges of gold and silver variograms, so the gold variogram parameters were selected as the controlling values of the primary elliptical interpolation pass.

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A second set of ellipse parameters was generated for a second interpolation pass, which were set at 130% of the range of the primary parameters.

Table 14.10 below summarizes the ellipse and variogram parameters generated from the primary variogram analysis for the separate domains.

Table 14.10: Ellipse and Variogram Parameters based on Experimental Variogram Analysis for the Santa Elena Model

Metal	Domain	Ellipse	Major	Semi-major	Minor	P. Azi	P. Dip	Int. Azi	Nugget	Total Sill
Au	WEST	ZAU_001	90.278	69.608	29.204	307.289	46.379	73.234	0.454	5.272
	CENTER	ZAU_101C	90.278	69.608	29.204	208.843	-49.833	87.938	1.205	3.647
	EAST	ZAU_101E	90.278	69.608	29.204	275.857	8.34	14.225	0.4322	6.007
	FW_SPLAY	ZAU_101E	90.278	69.608	29.204	275.857	8.34	14.225	0.4322	6.007
	HW_SPLAY	ZAU_101E	90.278	69.608	29.204	275.857	8.34	14.225	0.4322	6.007
	FLEXURES	ZAU_001	90.278	69.608	29.204	307.289	46.379	73.234	0.454	5.272
Ag	WEST	ZAU_001	90.278	69.608	29.204	307.289	46.379	73.234	0.454	5.272
	CENTER	ZAU_101C	90.278	69.608	29.204	208.843	-49.833	87.938	1.205	3.647
	EAST	ZAU_101E	90.278	69.608	29.204	275.857	8.34	14.225	0.4322	6.007
	FW_SPLAY	ZAU_101E	90.278	69.608	29.204	275.857	8.34	14.225	0.4322	6.007
	HW_SPLAY	ZAU_101E	90.278	69.608	29.204	275.857	8.34	14.225	0.4322	6.007
	FLEXURES	ZAU_001	90.278	69.608	29.204	307.289	46.379	73.234	0.454	5.272
Metal	Domain	Ellipse	Major	Semi-major	Minor	Rot Z	Rot X	Rot Z
Au	Cholugo	CHEL_AU	150	150	150	5	75	310	n/a	n/a
	Cholugo D	CHEL2_AU	150	150	150	0	90	310	n/a	n/a
Ag	Cholugo	CHEL_AU	150	150	150	5	75	310	n/a	n/a
	Cholugo D	CHEL2_AU	150	150	150	0	90	310	n/a	n/a

Metal grade values were interpolated into the block model from the one metre composite dataset using the variogram and ellipse parameters applied to and Ordinary Kriging algorithm.  A minimum of 2 composites, to a maximum of 15, were required to interpolate block grades with no more than 5 composites reporting from any one drill hole.

14.4.4   Block Classification

Mineral resources have been classified based on the CIM Definition Standards for Mineral Resources and Mineral Reserves. The category the resources have been assigned is based on the confidence in geological information available relating to the mineral deposit, the quantity and quality of data available, the level of detail of the technical and economic information which has been generated and the interpretation of the data and information.  EBA confirms that the resources for the Santa Elena deposit meet the test of being a reasonable prospect for consideration of economic extraction based on the success of currently open pit production and continuity of mineralization along defined structural trends.

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Interpolation of gold and silver grades used the variogram and search parameters outlined above as the basis for estimation.  Following the interpolation, a script was run to assign classification scheme to each block to represent a level of confidence for the reported grades.

An Inferred classification has been applied to the block model to target peripheral blocks that indicate the presence and continuity of mineralization, but lack the density of data necessary for confirmation.  Inferred blocks within the Santa Elena deposit lie within the maximum variogram range of 117 metres and have grade support from at least 1 drill hole and a minimum of 2 individual composite grades.  All mineralized blocks within the Cholugo and Cholugo Dos veins were assigned as Inferred.

An Indicated classification has been applied to the block model to target portions of the mineralized body where data density confirms the presence and continuity of mineralization with a moderate level of confidence. A factor of 0.9 was applied to the total sill of the experimental variogram to determine the range with a level of confidence to support classification of Indicated Mineral Resources as seen below in Figure 14.11.  This resulted in a value of 60 metres.  All mineralized blocks with grade support from at least 2 drill holes and a minimum of 10 reporting composites within a distance of 60 metres were assigned to the Indicated category.

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Figure 14.11: Factor applied to Total Sill of the Major Experimental Variogram Range

14.4.5   Underground Mineral Resource Estimate

Mineral Resource Estimates have been completed for the underground portion of the Santa Elena deposit (MMZ), including the co-trending Cholugo and Cholugo Dos zones.  The estimates have been prepared by EBA using to the Standards of Disclosure for Mineral Projects as documented in National Instrument 43-101 and adhering to the Canadian Institute of Mining, Metallurgy and Petroleum Best Practice Guidelines.  A summary of the estimates are presented with variable cut-off grades in Table 14.11 including the reserve estimate from Section 15, and in Table 14.12 as mineral resources excluding the reserves.

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Table 14.11: Updated Underground Mineral Resource Estimate including Reserves for Santa Elena (Effective April 30th, 2013)

Area	Classification*	AuEQ** Cut-off (g/t)	Tonnage	Au	Ag	Contained Au Ounces	Contained Ag Ounces
Santa Elena	Indicated	1.50	5,443,000	1.83	125	321,000	21,828,000
Santa Elena	Indicated	1.40	5,632,000	1.79	123	324,000	22,183,000
Santa Elena	Indicated	1.30	5,816,000	1.75	120	328,000	22,505,000
Santa Elena	Inferred	1.50	787,000	1.50	170	38,000	4,305,000
Santa Elena	Inferred	1.40	815,000	1.47	166	38,000	4,357,000
Santa Elena	Inferred	1.30	874,000	1.40	159	39,000	4,463,000
Cholugos	Inferred	1.50	699,000	1.67	147	37,000	3,305,000
Cholugos	Inferred	1.40	737,000	1.61	143	38,000	3,378,000
Cholugos	Inferred	1.30	782,000	1.55	138	39,000	3,457,000
Total	Indicated	1.40	5,632,000	1.79	123	324,000	22,183,000
Total	Inferred	1.40	1,552,000	1.54	155	76,000	7,735,000
*Classified by EBA, A Tetra Tech Company and conforms to NI 43-101 and CIM definitions for resources. Mineral Resources have been estimated from geological evidence and limited sampling and must not be considered to imply direct mineability. The reported baseline Mineral Resource for each zone is based on a 1.4 gpt AuEQ cut-off and is highlighted in light blue.  Grade capping has been applied at 12 gpt Au and 600 gpt Ag, **Based on Au:Ag ratio of 70, incorporating metal price assumptions of $1,450/oz Au, $28/oz Ag and using metallurgical recoveries of 92% Au and 67.5% Ag. *** Tonnage and contained ounces have been rounded to the nearest thousand.

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Table 14.12: Updated Underground Mineral Resource Estimate Excluding Reserves for Santa Elena (Effective April 30th, 2013)

Area	Classification*	AuEQ** Cut-off (g/t)	Tonnage	Au	Ag	Contained Au Ounces	Contained Ag Ounces
Santa Elena	Indicated	1.50	2,055,000	1.73	117	114,000	7,753,000
Santa Elena	Indicated	1.40	2,143,000	1.69	115	116,000	7,919,000
Santa Elena	Indicated	1.30	2,235,000	1.64	112	118,000	8,082,000
Santa Elena	Inferred	1.50	725,000	1.42	173	33,000	4,026,000
Santa Elena	Inferred	1.40	752,000	1.39	168	34,000	4,075,000
Santa Elena	Inferred	1.30	811,000	1.33	160	35,000	4,181,000
Cholugos	Inferred	1.50
Cholugos	Inferred	1.40	737,000	1.61	143	38,000	3,378,000
Cholugos	Inferred	1.30	782,000	1.55	138	39,000	3,457,000
Total	Indicated	1.40	2,143,000	1.69	115	116,000	7,919,000
Total	Inferred	1.40	1,490,000	1.50	156	72,000	7,453,000
*Classified by EBA, A Tetra Tech Company and conforms to NI 43-101 and CIM definitions for resources. Mineral Resources have been estimated from geological evidence and limited sampling and must not be considered to imply direct mineability. The reported baseline Mineral Resource for each zone is based on a 1.4 gpt AuEQ cut-off and is highlighted in light blue.  Grade capping has been applied at 12 gpt Au and 600 gpt Ag, **Based on Au:Ag ratio of 70, incorporating metal price assumptions of $1,450/oz Au, $28/oz Ag and using metallurgical recoveries of 92% Au and 67.5% Ag. *** Tonnage and contained ounces have been rounded to the nearest thousand.

Figure14.12 below shows the block model for the Santa Elena underground resource, inclusive of reserves. Blocks displayed are above 1.4 gpt AuEQ and are classified as Indicated.  Also shown are the drill holes used in resource estimation (Appendix D).

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Figure 14.12: Oblique Perspective of Santa Elena Block Model, Indicated Resource (AuEq >1.4 gpt)

14.4.6   Underground Validation

Estimated block grades were visually compared to drill hole grades along horizontal and vertical sections using the Santa Elena GEMS model (Figures 14.13 to 14.16).  In addition, Nearest Neighbor (NN), Inverse Distance Weighted methods, (IDW5 and IDW3) were run to compare against the Ordinary Kriging method (OK) used in the actual resource estimate.  Average Au and Ag grades were reported for 25 metre vertical and horizontal slices of the resulting verification and plotted for comparison against the average input composite grades within the same sections.  The results were compared to check for potential global biases of the primary OK calculation method.  In general, an overall grade smoothing trend is observed in all methods.  The smoothing effect in areas with discreet high grade peaks in the composite samples data effectively accounts for the declustering of data within the block model in higher grade areas which are represented by high number of densely populated sampling points. OK has declustered the data but a directional bias as part of the OK algorithm highlights high grade trends in areas with high density of sample points which is not seen in the IDW runs.

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Figure 14.13: Vertical Section Comparison of Model Runs and input Au grades

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Figure 14.14: Vertical Section Comparison of Model Runs and input Ag Grades

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Figure 14.15: Plan Section Comparison of Model Runs and input Au Grades

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Figure 14.16: Vertical Section Comparison of Model Runs and input Ag Grades

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Overall, the four methods depict a close approximation to the overall grade trends shown by the composite sampling data.  The individual methods show repeatability and do not indicate that biases have been introduced during the modelling process.  The OK method is thought to support the local variation in grade and is felt to be an adequate representation of the grade distribution throughout the deposit.

14.4.7   EBA Statement on Mineral Resource Estimates

EBA has compared and reviewed the available datasets for the block model resources and feels that the block models used to report the mineral resource estimate as stated above are valid and represent the source data in a reasonable manner.

14.5   Heap Leach Mineral Resource Estimate

The Leach Pad Resources (spent ore) have been estimated by SilverCrest utilizing daily production, grade and mass records obtained at the crushing plant when conveying mineralized material to the pad. These records represent a complete production dataset since start-up of the operations at Santa Elena. A detailed leach pad survey was completed in April 2013 to provide volumes of material placed on the pad. This volume was compared with weightometer data collected on the crushing plant to validate tonnes on the pad using well known specific gravity on the leach pad of 1.85 g/cm3.

A block model from validation drilling was completed on 25 holes spaced on a 25 metre by 25 metre grid as described below. The validation results support the estimation by the Company utilizing daily production data.

Table 14.13: Validation of Heap Leach Resource Estimate

	Tonnage	Au	Ag
	x 1000	Grade(g/t)	Grade(g/t)
Tonnage and grade based on daily production	2,844.568	0.65	33.3
Tonnage and grade based on block model	2,853.949	0.54	30.10

14.5.1   Heap Leach Resource Validation

The drilling database used for the grade interpolation includes 25 large diameter reverse circulation Becker Hammer, Tandem Mounted hydraulic drill holes for a total length of 355 metres. A total of 169 samples were integrated in an Access/ Dassault Systemes database together with the most recent topographic survey of the heap leach pad. The drill hole data is spaced on a 25 metre grid on the pad (Figure 14.17).

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Figure 14.17: Leach Pad Verification Drill Hole Location Map

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A block model comprising the volumetric of the current leach pad area was built with the following specifications:

Table 14.14: Leach Pad Block Model Origin and Dimensions

Database	Direction	Origin	Block Size (m)	Number of Blocks	Dimension (m)	Rotation (deg)
Leach Pad	East (rows)	580,570	100	7	700	0
	North (columns)	3,320,970	100	3	300	0
	Elevation (levels)	850	6	20	120	0

Original samples were not composited, maintaining the original length. An inverse distance squared methodology was applied with a single domain and isotropic search ellipse (200 meters -same search distance in all directions).  This provides a close to polygonal estimation approach for the resource estimation due to small area and limited number of samples.  The grade interpolation parameters included no grade capping. A minimum of 2s and a  maximum of 12 composites were used to interpolate block grades, with no more than 3 composites reporting from any one drill hole in multi stage search radiuses to populate all blocks. Visual validation was performed to assess the results and final resource numbers were benchmarked against production data for validation purposes.

Table 14.15: Leach Pad Material Mineral Resource Estimate (Effective April 30th, 2013)

Area	Classification*	AuEQ** Cut- off (gpt)	Tonnage	Au	Ag	Contained Au Ounces	Contained Ag Ounces
Leach Pad	Indicated	0.00
Leach Pad	Indicated	0.50	2,844,530	0.65	33.3	59,420	3,048,200
Leach Pad	Indicated	1.00
Total	Indicated	0.50	2,844,530	0.65	33.3	59,420	3,048,200
*Classified by SilverCrest Mines Inc. and conforms to NI 43-101 and CIM definitions for resources. Mineral Resources have been estimated from drill hole data and validated using company production data and sampling. The estimate must not be considered to imply economic mineability. The reported baseline Mineral Resource for the leach pad material is based on a 0.5 gpt AuEQ cut-off and is highlighted in light blue.  No grade capping has been applied. **Based on Au:Ag ratio of 70, incorporating metal price assumptions of $1,450/oz Au, $28/oz Ag and using metallurgical recoveries of 92% Au and 67.5% Ag.  All leach pad spent ore is considered above cutoff grade of 1.0 gpt AuEQ.

14.6   Open Pit Mineral Resource Estimate

The updated open pit Resources have been estimated by SilverCrest utilizing Company production data (blastholes), exploration drilling data and computer modelling. All Indicated Resources in the operating open pit have been converted to Reserves as of December 31, 2012.

The Open Pit Mineral Resource estimates are based on verified information from historical and recent SilverCrest sources. Solid boundaries for the mineralization were constrained at 0.20 gpt AuEQ cut off and the open pit limits defines the mineralized material dated as December 31, 2012. The Resource estimation was completed using Gemcom GEMS software applying Ordinary Kriging. Raw data (blast holes, core and RC samples) were composited to a 5 metre length for blast holes and 2 metre length for in-pit core and RC samples. These composites were assessed for spatial and geostatistical purposes. A capping of 8 gpt Au and 300 gpt Ag was applied to core and RC data only. The composite data was interpolated into a 5 metre x 5 metre x 5 metre block size model using Ordinary Kriging in two passes. The first pass using only blast hole data with an average spacing of 4 metres and interpolating grades to a maximum distance of 15 metres from the sample location. The second pass used only core and RC drilling data spaced on average 50 metres to interpolate 5 metre x 5 metre x 5 metre blocks located in distances beyond 15 metres from the sample location. The blasthole resource model was compared against the production data for verification purposes.

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15.0   MINERAL RESERVES ESTIMATES

15.1   Introduction

Three individual Mineral Reserve estimates are presented in this report and are combined into the mine schedule and economic analysis that comprises the Santa Elena Expansion Project.

An updated Reserve estimate for the Santa Elena open pit and a new Reserve estimate for spent ore material on the existing leach pads have been prepared by Eric Fier, CPG, Chief Operating Officer of SilverCrest and non-Independent Qualified Person, and has accounted for production up to the Effective Date.  A new Reserve estimate for the proposed underground mine has been prepared by Mike Tansey, P.Eng, Senior Mining Engineer with EBA and an Independent Qualified Person.

On the basis of resources classified as Indicated in Section 14, EBA has estimated Santa Elena underground diluted and recoverable Reserves. SilverCrest have estimated leach pad and open pit Reserves (Table 15.2).

15.2   Previous Reserve Estimates

An updated reserve estimate for open pit mining operations at Santa Elena was previously completed by SilverCrest, with the assistance of Dassault Systemes Software International, and reviewed by EBA in January, 2011.  This Reserve estimate was released by EBA in May 2011 in the report “NI 43-101 Technical Report, Reserve Update for the Santa Elena Open Pit and Preliminary Assessment for the Santa Elena and Cruz de Mayo Expansion Project, Sonora Mexico (amended date May 11, 2011)”.

The open pit Reserves reported in 2011, and shown in table 15.1 below, are superseded by the current open pit Reserve estimate described in this report.   Significant variance from the current open pit Reserve estimate is attributed to the open pit design which was recently reduced in size from the 2011 pit.

Table 15.1: Previous Santa Elena Open Pit Reserves (January 2011)

Classification	Tonnes	Gold (gpt)	Silver (gpt)	Contained Gold (troy/oz)	Contained Silver (troy/oz)
Probable	4,794,790	1.81	75.9	278,560	11,711,000
* Represents the diluted and mine recoverable Reserves based on $USD1,000/oz Au and $USD18/oz of Ag, cut-off grade of 0.38g/t AuEQ with applied metallurgical recoveries.  Capped grades at 12 gpt Au and 400 gpt Ag.  Based on a bulk density of 2.60 t/m3.   No previous Reserve Estimates were completed for underground or leach pad.

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15.3   Current Reserves Overview and Assumptions

The classification of Reserves follows the Standards of Disclosure for Mineral Projects as documented in National Instrument 43-101 and adheres to the Canadian Institute of Mining, Metallurgy and Petroleum Best Practice Guidelines. Indicated Resources demonstrating economic viability are assigned the Probable category. No Measured Resources have been estimated, and therefore none of the Reserves estimated will be classified as Proven.  The Mineral Reserves (open pit, underground, leach pad) in table 15.2 are based on the following parameters where applicable:

1.   Gold price of $ 1450 / troy ounce (3 year trailing average based - effective date 30th April 2013)

2.   Silver price of $28 / troy ounce (3 year trailing average based - effective date 30th April 2013)

3.   Silver grade was converted to a gold equivalent (AU EQ) using a 70:1 Ag (g/t) to Au (g/t) ratio

4.   Mining recovery of 85% for long hole stoping

5.   Mining recovery of 95% for mechanised cut and fill mining methods

6.   Global AU EQ cut-off grade of 1.5 g/t for stope design

7.   Specific cut off grades applied to each stope based on type of stoping and stope dimensions (see table 15.4).

8.   Total cost per tonne of $13.9 for open pit heap leach including waste, processing and G&A

9.   Total Cost per tonne of $28 for open pit milling

10.   Total cost per tonne of $ 61 (undiluted) for long hole stoping methods including stope development, processing and G and A costs

11.   Total cost per tonne of $ 65 (undiluted) for mechanised cut and fill stoping methods including stope development, processing and G and A costs

12.   Total cost per tonne of $22 for leach pad spent ore milling

13.   Ore density of 2.6 tonnes / m3

Table 15.2: 2013 Santa Elena Mineral Reserve Estimates

Classification*	Tonnes	Au gpt	Ag gpt	Au oz	Ag oz
SANTA ELENA UNDERGROUND DILUTED AND RECOVERABLE RESERVES**
PROBABLE	3,920,510	1.57	108.1	198,350	13,637,010
SANTA ELENA OPEN PIT RESERVES (As of April 30, 2013)***
PROBABLE	1,426,710	1.52	66.8	69,890	3,064,980
SANTA ELENA LEACH PAD RESERVES****
PROBABLE	2,844,530	0.66	29	60,390	2,653,810
TOTAL RESERVES	8,191,760	1.25	73.4	328,640	19,355,800
Classification	Tonnes	Au gpt	Ag gpt	Au oz	Ag oz

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Note: Tonnes and ounces are rounded. Underground and Leach Pad Reserves are based on 3 year historic metal price trends of US$28/oz silver, US$1450/oz gold and metallurgical recoveries of 92% Au and 67.5% Ag with a metal ratio of Ag:Au at 70:1. All Mineral Resources and Reserves conform to NI 43-101 and CIM definitions for Resources and Reserves.

*Open Pit and Leach Pad Probable Reserves were Classified by SilverCrest.  Underground Reserves and Resources were classified by EBA.

**Underground Probable Reserve is based on a cut-off grade of 1.47 gpt AuEq with an average 11% dilution and 90% mine recovery.

***Open Pit Reserve is based on a cutoff grade of 0.20 gpt AuEq in a constrained pit shell with applied capping of 8 gpt Au and 300 gpt Ag.

****Leach Pad Reserve based on production and drill hole data for volumetrics and grade model using a cutoff grade of 0.5 gpt AuEq. No capping was applied.

*****Underground Resources are based on 1 gpt AuEq grade shell and cutoff grade of 1.4 gpt AuEq with applied capping of 12 gpt Au and 600 gpt Ag.

15.4   Basis for Underground Mineral Reserve

The underground Mineral Reserve tonnages are based on the block model used for Indicated Resources. The block model has been developed using 5 metre x 5 metre x 5 metre blocks based on a grade shells modeled into a grade (mineralization) model created using drill hole intersections of greater than 1.0 g/t AU EQ. The Block model represents estimated Au and Ag grades interpolated using capped composite samples calculated by Ordinary Kriging (OK) into each block (refer to section 14 for details of resource modelling). The block model is clipped to the topography and/or the currently planned ultimate pit. Reserves were calculated using the grade shell model, restricted laterally by the grade (vein) model, and running various stope length and height iterations to obtain the optimal stope dimension.

EBA estimated reserve grade and tonnage through cost approximation and thereby determination of a cut-off grade for AUEQ based on gold price and metallurgical recovery (see section 13 and 17). Rock engineering criteria and analysis results (see section 16.2) were used to determine provisional stable stope spans, using a set stope height of 45 m.  The stope height was based on a conceptual model for long hole stoping, with two sublevels in each stope, separated by 22.5 m.  Mineral reserves were initially estimated at a cut-off of 1.5 g/t Au Eq for all stopes. Specific cut-off grades were then determined by mining method after mining costs were established by an iterative optimization process.  Within the stope shapes created, Indicated Resources above and below the cut-off grades were then considered for definition of undiluted Reserves. Other material within the stope shapes, including Inferred Resources, are considered as zero grade dilution, resulting in internally diluted Reserves. Optimization was undertaken by excluding low grade areas from the stopes and through placement of ground stability rib and sill pillars outside of high grade areas. Figure 15.1 shows the stope shapes generated for the Reserve definition.

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Figure 15.1: Stope Shapes defined for use in Reserve Definition

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External dilution (outside of the modelled stope) is added to the grade and tonnage to determine the Probable Reserves.  The external dilution is based on the type of stoping being undertaken, rock engineering design spans and resultant dilution, minimum mining widths and consideration for blast damage.

Table 15.3 below summarizes the Santa Elena underground reserves by tonnage contained in stopes, recoverable pillars, and development drifts situated in ore.  Table 15.4 shows the breakdown of the reserves, with the estimation of dilution and mining recovery on a stope by stope basis. Table 15.4 also shows the reserves sourced from pillars and ore development.

Table 15.3: Breakdown Summary of Santa Elena Underground Reserves

Area	Category	Tonnes	Grade Au (gpt)	Grade Ag (gpt)
Stopes	Probable	3,785,816	1.58	108
Pillars	Probable	109,699	1.36	122
Drifts	Probable	24,518	0.76	63
Total	Probable	3,920,518	1.57	108

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Table 15.4: Santa Elena indicated Stope Tonnage & Grade

Stope ID	Type	Internally Diluted Ore - From Stope Shapes				External Dilution And Mining Recovery Parameters					Diluted Ore Tonnage And Grade						Cut-Off
		Stope Tonnes	AuEQ Grade	Au Cap Grade	Ag Cap Grade	Mining Recovery	External Dilution Percentage	AuEQ Dilution Grade	Au Dilution Grade	Ag Dilution gpt	Net Tonnes	Stope Length	Tonnes Stoping	Net Grade AuEQ	Net Grade Au	Net Grade Ag	Cut-off
	Units	Metric Tonnes	gpt	gpt	gpt	%	%	gpt	gpt	gpt	metric tonnes			gpt	gpt	gpt	gpt
	Calculation	A	B	C	D	G	H	I	J	K	L=A x G + H x A x G			= (A x B + A x H x I) x G / L	= (A x C + A x H x J) x G / L	= (A x D + A x H x K) x G / L
	Notes/source	GEMS Model				Mining Parameters by QP		GEMS Model			Calculation						Costing
710-1	LH	69,477	4.26	2.39	130.82	85.0%	15.0%	0.23	0.13	6.52	67,914	125	52,314	3.74	2.10	114.61	1.65
655-1	LH	3,262	2.86	1.84	88.69	85.0%	28.1%	0.41	0.06	24.58	3,551	45	3,551	2.32	1.45	74.64	1.65
655-2	LH	19,430	1.85	1.19	46.48	85.0%	14.5%	0.15	0.06	6.25	18,912	53	12,298	1.64	1.04	41.38	1.65
655-3	LH	68,023	2.54	1.47	74.44	85.0%	15.0%	0.38	0.25	8.91	66,492	60	59,004	2.26	1.31	65.89	1.50
655-4	LH	44,578	5.62	3.11	176.27	85.0%	15.0%	0.34	0.18	11.02	43,575	43	38,208	4.93	2.72	154.71	1.50
655-5	LH	137,989	5.98	3.42	180.86	85.0%	15.0%	0.30	0.14	11.45	134,885	105	121,781	5.24	3.00	158.76	1.50
655-6	TVLH	208,387	3.20	1.46	122.44	90.0%	10.0%	0.34	0.09	17.60	206,303	145	206,303	2.94	1.33	112.91	1.43
655-7	LH	181,527	2.08	0.98	78.79	85.0%	15.0%	0.37	0.16	14.84	177,443	125	161,843	1.86	0.87	70.45	1.28
597-1	LH	49,648	4.81	2.81	159.83	85.0%	27.3%	0.22	0.07	10.82	53,708	120	38,732	3.83	2.22	127.90	1.65
597-2	LH	137,487	3.71	2.12	111.16	85.0%	15.0%	0.56	0.30	18.39	134,393	96	122,413	3.30	1.88	99.06	1.50
597-3	LH	188,369	4.11	2.20	134.42	85.0%	15.0%	0.32	0.15	11.74	184,131	120	169,155	3.62	1.94	118.42	1.50
597-4	TVLH	207,563	2.81	1.11	118.83	90.0%	10.0%	0.32	0.13	12.81	205,488	100	205,488	2.58	1.02	109.19	1.46
597-5	LH	73,086	2.27	1.18	76.17	85.0%	15.0%	0.34	0.12	15.14	71,441	43	66,075	2.02	1.04	68.21	1.22
597-6	LH	172,058	2.19	0.88	105.33	85.0%	15.0%	0.42	0.19	16.37	168,187	102	155,457	1.96	0.79	93.73	1.22
536-1	CF	112,971	4.06	2.23	133.41	95.0%	5.0%	0.11	0.04	4.73	112,688	0	112,688	3.87	2.12	127.29	1.59
536-2	LH	147,956	3.60	1.41	153.59	85.0%	15.0%	0.27	0.16	7.86	144,627	80	134,643	3.17	1.24	134.58	1.22
536-3	LH	128,602	3.06	1.41	115.61	85.0%	15.0%	0.28	0.14	9.51	125,708	80	115,724	2.70	1.25	101.77	1.22
536-4	LH	124,514	2.61	1.05	109.39	85.0%	15.0%	0.39	0.18	14.82	121,713	75	112,353	2.32	0.94	97.05	1.22
536-5	TVLH	262,753	2.33	0.96	96.36	90.0%	10.0%	0.30	0.12	12.63	260,126	100	260,126	2.15	0.88	88.75	1.43
536-6	LH	35,994	1.78	0.50	105.42	85.0%	15.0%	0.22	0.06	11.27	35,184	0	35,184	1.58	0.45	93.14	1.65
471-1	CF	153,808	4.84	3.86	114.30	95.0%	3.3%	0.17	0.07	6.70	150,988	0	150,988	4.69	3.74	110.83	1.53
471-2	CF	57,771	5.36	3.73	113.66	95.0%	3.6%	0.18	0.08	7.01	56,842	0	56,842	5.18	3.61	109.99	1.53
471-3	CF	194,022	5.87	3.56	161.85	95.0%	4.0%	0.17	0.06	7.50	191,693	0	191,693	5.65	3.42	155.91	1.53
471-4	CF	46,026	2.75	1.28	102.53	95.0%	6.7%	0.31	0.09	15.09	46,640	0	46,640	2.60	1.21	97.07	1.59
471-5	LH	54,512	2.22	0.97	87.78	85.0%	15.0%	0.31	0.14	11.96	53,285	95	41,429	1.97	0.86	77.89	1.65
471-6	LH	67,625	1.91	0.63	89.88	85.0%	15.0%	0.36	0.18	12.05	66,103	56	59,115	1.71	0.57	79.72	1.50
471-8	TVLH	223,073	1.36	0.51	56.77	90.0%	10.0%	0.21	0.06	10.53	220,842	150	220,842	1.26	0.47	52.56	1.46
401-1	CF	32,822	3.93	2.92	163.82	95.0%	5.0%	0.19	0.11	6.00	32,740	0	32,740	3.75	2.79	156.31	1.59
401-2	CF	169,232	5.18	2.80	167.10	95.0%	4.0%	0.36	0.18	13.16	167,201	0	167,201	5.00	2.70	161.18	1.59
401-3	CF	158,590	1.98	0.98	70.53	95.0%	5.0%	0.45	0.15	20.90	158,194	0	158,194	1.91	0.94	68.17	1.59
401-4	CF	62,318	1.92	1.10	62.74	95.0%	6.3%	0.10	0.03	5.42	62,902	0	62,902	1.81	1.04	59.37	1.59
327-1	CF	22,856	3.63	1.54	165.37	95.0%	7.1%	0.26	0.14	7.83	23,264	0	23,264	3.41	1.45	154.86	1.68

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Stope ID

Type

Internally Diluted Ore - From Stope Shapes

External Dilution And Mining Recovery Parameters

Diluted Ore Tonnage And Grade

Cut-Off

Stope Tonnes

AuEQ Grade

Au Cap Grade

Ag Cap Grade

Mining Recovery

External Dilution Percentage

AuEQ Dilution Grade

Au Dilution Grade

Ag Dilution gpt

Net Tonnes

Stope Length

Tonnes Stoping

Net Grade AuEQ

Net Grade Au

Net Grade Ag

Cut-off

Units

Metric Tonnes

gpt

gpt

gpt

%

%

gpt

gpt

gpt

metric tonnes

gpt

gpt

gpt

gpt

Calculation

A

B

C

D

G

H

I

J

K

L=A x G + H x A x G

= (A x B + A x H x I) x G / L

= (A x C + A x H x J) x G / L

= (A x D + A x H x K) x G / L

Notes/source

GEMS Model

Mining Parameters by QP

GEMS Model

Calculation

Costing

327-2	CF	148,618	3.60	1.34	173.05	95.0%	7.1%	0.37	0.13	16.64	151,272	0	151,272	3.39	1.26	162.62	1.68
327-3	CF	68,480	2.85	0.94	133.62	95.0%	3.6%	0.44	0.17	18.84	67,380	0	67,380	2.77	0.91	129.66	1.53
Subtotal Stopes Only		3,833,427	3.34	1.74	119.25	90.5%	10.2%	0.32	0.14	12.62	3,785,816			3.07	1.58	108.02	1.47
Rib pillar 1	Rib pillar	36,908	3.21	1.64	110.06	80.0%	15.0%	0.34	0.15	12.86	32,528	0	20,131	2.96	1.50	101.65	1.50
Rib pillar 2	Rib pillar	21,882	5.24	2.51	191.17	80.0%	15.0%	0.55	0.23	22.34	20,131	0	57,039	4.63	2.22	169.15	1.65
Sill pillar 1	Sill pillar	61,999	3.00	1.12	132.22	80.0%	15.0%	0.32	0.10	15.45	57,039	0	77,466	2.65	0.98	116.99	1.46
Subtotal Pillar Extraction		120,788	3.47	1.68	149.89	80.0%	13.2%	0.37	0.14	15.91	109,699			3.11	1.36	122.01	1.51
Subtotal Stoping		3,954,000	3.34	1.74	120.10	90.4%	10.3%	0.32	0.14	12.74	3,896,000			3.07	1.58	108.40	1.47
Ore drifting outside stopes	Stope access ore drifts in the vein	24,518	1.67	0.76	63.40	100.0%	0.0%	0.00	0.00	0.00	24,518	0	20,131	1.67	0.76	63.40	0.55
Total Stopes & Ore Drifts		3,978,518	3.33	1.74	119.75	90.5%	10.2%	0.32	0.14	12.66	3,920,518			3.06	1.57	108.12	1.47

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15.4.1   Underground Mining Methodologies Consideration for Underground Reserves

Three primary mining methods were considered for estimation of costs and thereby mineral reserves, namely; longitudinal long hole stoping, transverse long hole stoping, and mechanized cut and fill (refer to Section 16.3 for explanation of mining methods).

Longitudinal long hole stoping has the highest estimated external dilution for the three methods in, with an average of 15% with select few stopes as high as 28% (due to the width of the vein being lower than the minimum mining width).  This method constitutes the lowest overall operating cost and has been applied to all stopes with widths less than 15 m (hanging wall to footwall) and a footwall dip angle of greater than 50 °.

For stopes with dip angles greater than 50 ° and width greater than 15 m, transverse long hole stoping has been selected. This method requires more waste footwall development and is thereby marginally higher cost than longitudinal long hole stoping.  Lower dilution has been included for transverse stopes, as the spans of the stope are limited to 20 m widths at any one point; see section 16 for mining method descriptions. Table 15.5 shows the breakdown of mining methods utilized throughout the life of the mine. Calculations are based on tonnages mined and mining method employed in each stope.

Table 15.5: Mining Methods % Utilization throughout LOM

Mining Methods Percent Utilization throughout LOM
Mining Method	% Utilization
Long hole Stoping	69%
Cut & Fill	31%

Mechanised cut and fill stoping has been applied to all areas with footwall dip angle of less than 50° from the horizontal.  This method has the highest cost, due to footwall development requirements and productivity constraints due to reduced size of cuts taken at any point and the need to backfill .  External dilution is estimated to be less for mechanized cut and fill due to hanging wall exposures being limited to a single cut at any point. In addition, sloughage of the hanging wall  will be controlled by the backfill and through placement of support in the form of rock bolts and/or split sets in the active cut.

15.5   Open Pit Reserves (SVL)

All open pit Mineral Resources are declared as Reserves (table 15.6). Open pit production from January 1 to April 30, 2013 was subtracted from the December 2012 Resource estimation.

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Table 15.6: Santa Elena Open Pit Mineral Reserves (Effective December 31, 2012)

Classification*	Tonnes	Au gpt	Ag gpt	Au oz	Ag oz
SANTA ELENA OPEN PIT RESERVES (As of April 30, 2013)***
PROBABLE	1,426,710	1.52	66.8	69,890	3,064,980
*Open Pit and Leach Pad Probable Reserves were Classified by SilverCrest. ***Open Pit Reserve is based on a cutoff grade of 0.20 gpt  AuEq in a constrained pit shell with applied capping of 8 gpt Au and 300 gpt Ag.

15.6   Heap leach Reserves (SVL)

Section 14.4 details the estimation of heap leach Indicated Resources.  Based on mining practicality, no cut-off grades between 0 and 1 gpt AuEQ were applied. All average grades for the leach pad spent ore are above 1gpt AuEQ, therefore all tested resource tonnes on the pad are considered reserves.

All spent ore material on the leach pad will be reprocessed through the processing facility once operational in the first quarter of 2014. Only the material currently on the pad and leached (300 day cycle completed or estimated full cycle) has been declared as Reserves. Approximately 882,000 tonnes of open pit material is planned to be placed on the pad during the remainder of 2013 and will undergo only partial leaching before being re-processed through the new facility. Once this material is placed on the pad and leaching has been discontinued, it will be declared as leach pad Reserves for processing and metal recovery in 2014. Extensive metallurgical test work including ongoing operations data show that all declared Reserves are amenable to conventional leaching either by heap leach technology or standard CCD milling with a Merrill Crowe recovery system for doré bar production.

Table 15.7: Heap Leach Material Mineral Reserve Estimate (Effective April 30th, 2013)

Area	Classification*	AuEQ** Cut-off (g/t)	Tonnage	Au	Ag	Contained Au Ounces	Contained Ag Ounces
Leach Pad	Probable	0.50	2,844,530	0.65	33.3	59,420	3,048,200
*Classified by SilverCrest Mines Inc. and conforms to NI 43-101 and CIM definitions for resources. Mineral Resources have been estimated from drill hole data and validated using company production data and sampling. The estimate must not be considered to imply economic mineability. The reported baseline Mineral Resource for the leach pad material is based on a 0.5 gpt AuEQ cut-off and is highlighted in light blue.  No grade capping has been applied. **Based on Au:Ag ratio of 70, incorporating metal price assumptions of $1,450/oz Au, $28/oz Ag and using metallurgical recoveries of 92% Au and 67.5% Ag. All leach pad spent ore is considered above cutoff grade of 1gpt AuEQ

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16.0   MINING METHODS

16.1   Introduction

This section includes a geotechnical analysis, explanation of mining methods and mine design methodology completed for the Prefeasibility Study based on the information available and with the assumption that further investigation and detailed engineering will be completed prior to or during development. In practice the actual conditions will vary based on actual ground conditions encountered. As more drilling is completed and the mine is developed, the mining methods, stope layouts and rock engineering criteria will be modified in favour of actual conditions.

Conventional open pit mining will continue using a contractor until the third quarter of 2014 when the open pit reserves will be depleted. Mining of the heap leach spent ore will be completed by loader and conveyor to transport material to the plant.

The analysis was completed using the basis of assumptions for geological model described in Section 14, and mining methods were selected based on the results of preliminary geotechnical and rock mechanic analysis and to account for variations in vein thickness and orientation.  A detailed geotechnical analysis was conducted to determine specific parameters for each mining method.

16.2   Geotechnical Analysis for Underground Mining

16.2.1   General

The geotechnical analysis of underground production stoping at Santa Elena is based on geotechnical data collected from sixteen exploration drill holes, six of which were oriented, advanced as part of the 2012 subsurface exploration program and test work conducted on core samples at the University of Sonora, Hermosillo, Mexico.

This section addresses the present work conducted for the proposed Santa Elena Expansion Project and provides a summary of the range of variance of the rock mass conditions (by means of RMR76 and NGI-Q classification systems) at the following five zones:

§   Ore zone (referred herein as Zone D1).

§   Country rock above the hanging wall, from 5-6 m to 50-60 m above the upper ore zone/country rock interface (referred here in as Zone D2).

§   Country rock above the hanging wall, between the upper ore zone/country rock interface and about 5-6 m above it (referred here in as Zone D3).

§   Country rock below the foot wall, from 5-6 m to 50-60 m below the lower ore zone/country rock interface (referred here in as Zone D4).

§   Country rock below the foot wall, between the lower ore zone/country rock interface and about 5-6 m below it (referred here in as Zone D5).

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16.2.2   Geotechnical Data Collection

16.2.2.1   Drill Program

Table 16.1: 2012 Borehole Summary
Hole ID	Northing (m)	Easting (m)	Surface Elevation (m)	Azimuth (°)	Plunge (°)	Total Length (m)	Oriented Core Data
GT-12-09	3321358.7	81322.9	784.8	360	60	280.35	Y
GT-12-10	3321349.6	581448.9	776.8	360	70	310.50	Y
GT-12-11	3321348.7	581449.5	777.0	315	70	344.45	Y
GT-12-12	3321449.8	581599.1	808.1	315	45	240.35	Y
GT-12-13	3321383.0	581450.1	784	360	70	280.70	Y
SERC-12-32	3321351.7	581551.5	790.3	045	70	380.00	Y
SE-12-41	3321385	581450	784	360	45	234	N
SE-12-47	3321404	581525	793	015	60	228.45	N
SE-12-48	3321403.4	581525.8	788.9	010	50	215.15	N
SE-12-50	3321466.4	581607.9	808.6	345	78	278	N
SE-12-51	3321467.0	581608.0	808.7	015	70	275	N
SE-12-66	3321391.7	581579.7	802.0	358	86	359	N
SE-12-71	3321391.7	581579.7	802.0	320	82	332	N
SE-12-72	3321182.3	581734.3	853	030	65	600	N
SE-12-90	3321358.7	581323.0	784	020	55	218.75	N
SE-12-91	3321200.5	581640.6	853	013	59	455	N

The geotechnical investigation consisted of:

§   Geotechnical data collection on 16 exploration drill holes advanced as part of the 2012 subsurface exploration program.  The coordinates and elevations of the 16 drill holes are presented in Table 16.1.

§   Geological and geotechnical logging of the drill holes is listed in Table 16.1.  This work was conducted by Silver Crest’s mine geologist prior to splitting the core.  The geotechnical logging includes collection of geotechnical parameters to estimate the Rock Mass Rating (RMR) system (Bieniawski 1976) and the Norwegian Geotechnical Institute’s NGI Q-System (after Barton et al. 1974). EBA undertook QA/QC of the rock core during site visits at the Santa Elena project.

§   Six (6) of the 16 drill holes with core orientation (refer to Table 16.1).

§   Collection of rock samples for geotechnical laboratory purposes.

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16.2.3   Rock Mass Characteristics

EBA evaluated the rock mass quality using the NGI-Q and RMR systems for the 16 drill holes listed in Table 16.1.  For this purpose, the geotechnical logs were processed in raw spread sheets.  During this process, EBA reviewed (QA/QC) the geotechnical logs against the core photos and made corrections where appropriate.

The rock mass was subdivided into five geotechnically distinct zones as described in Section 16.2.1.

16.2.4   Design Parameters

16.2.4.1   Rock Mass parameters

The cumulative percentage distributions of the NGI-Q and RMR values were calculated for the 5 zones. The cumulative percentage distribution plots provide the means of progressively estimating the likelihood that an RMR or NGI-Q value will be equaled or exceeded. For this Prefeasibility study the NGI-Q and RMR values at 50% were chosen. The purpose of adopting the 50% design values for Prefeasibility design is to provide stope dimension and reinforcement that are not too conservative or too optimistic, and therefore serve for reasonable economic estimation.

Table 16.2 summarizes the NGI-Q and RMR values used for the preliminary design.

Table 16.2: RMR’ and NGI-Q’ Values for Design – at 50%
Zone	Elevation	RMR’76	NGI-Q’
D1	Above El. 580 m	75	29
	Below El. 580 m	76	11
D2	From surface to bottom of planned workings	52	4
D3	Above El. 580 m	67 (711)	12 (211)
	Below El. 580 m	52	3
D4	Above El. 580 m	70	13
	Below El. 580 m	67	13
D5	Above El. 580 m	67	7
	Below El. 580 m	72	14
1 Two values are presented; (i) the first value considers the data from all the drill holes, and (ii) the second (in brackets) does not consider three of the drill holes where potential faults were encountered.

In addition to this assessment set out in Table 2, NGI-Q values were assessed for the 655 Level. Core runs were isolated over a 45m run length at this level to give an individual rock mass characterization for stoping at this level. NGI-Q values of 15 and 6 were calculated for zones D1 and D3 for stopes at this level.

16.2.4.2   Rock Strength

Rock strength is based on laboratory testing results from University of Sonora together with a small sample of test work undertaken at University of British Columbia, Vancouver. The test work completed at Sonora consisted of simple Uniaxial Compressive Tests (UCS) without strain measurements, Brazilian Tests and Point Load Tests (PLT). The test work at UBC consisted of simple UCS and Brazilian tests for comparison. The results of the laboratory testing at both Universities is presented in Appendix E.

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The average ore body strength from the test work undertaken at the two institutes is 87 MPa.

The strength of the hanging wall for the prefeasibility analysis is taken as 92 MPa. This is based on the average point load testing data from University of Sonora. Three UCS tests were carried out in hanging wall material at University of Sonora giving an average of 95 MPa which ties in closely with the PLT results.

16.2.5   Design Approach

16.2.5.1   General

The design approach presented in this report consists of empirical and analytical methods using the rock mass quality approach.  In this respect, it should be noted that as with all the empirical methods, assumptions should be validated once mining starts by comparing actual with anticipated results.  Below is presented the empirical method used for the assessment.

16.2.5.2   Design Assumptions

The following assumptions have been made during the geotechnical design approach at Prefeasibility level for Santa Elena.

§   Modelling of the orebody and mine openings was undertaken using RocScience Examine 2Dtm software. As such, the rockmass is assumed isotropic and homogenous. See geotechnical recommendations in section 26.2.

§   Hoek-Brown parameters and deformation characteristics are generic dependent on rock type as established from the RocScience RocLabtm software program.

§   The orebody was split into three sections – West, Central and East and an average dip and strike was assigned to each.

§   Stopes are assumed to be of fixed width along strike with average dip and strike dependent on their location within the orebody.

§   Stope vertical height is set at 45 m.

§   No stress data is available at site. Stress is assumed to be depth dependent based on overburden. No variation in the horizontal to vertical stress ratio was incorporated at prefeasibility level.

§   Based on the ground water report undertaken by Global Resource Engineering (2011) and observations undertaken during ramp development the excavations were assumed to be dry or have minor inflow. A second hydro-geological assessment is recommended to confirm the initial findings of Global Resource Engineering.

16.2.5.3   Tabulation of Results

Tables 17 through 23 in Appendix E show the stope configuration summaries for each level for longitudinal and cut and fill mining. They indicate a stable span without support, a stable span with support and a measure of anticipated dilution with greater span lengths. Effective cable bolt support has been considered when designing maximum spans for stable spans with support. These stope design summary tables present the results of the analysis with respect to the rectangular stope design and are referred to in the body of this summary.  They are split into West, Central and East locations at each level dependent on average strike and dip of the orebody.

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These reference tables are meant for general design purposes only at a Prefeasibility level as the stope spans assume continuous width and dip profiles. As actual stopes vary in width and dip across their span, further geotechnical study of individual stopes is advised beyond Prefeasibility level.

Actual stope and pillar performance should be consistently monitored during mining and a database created. This will allow for any modifications of initial design to be made based on actual mining of the orebody.

Dilution levels for span lengths greater than those recommended from Potvin’s method are indicative only and it is strongly recommended that a site based dilution data base be established to monitor sloughage based on stope spans and rockmass quality. Further rockmass classification through mapping and logging can assist in defining stope dimensions to a higher level of detail. Actual stope and pillar performance should be consistently monitored during mining. This will allow for any modifications of initial design to be made based on actual mining of the orebody.

16.2.5.4   Methodology of Stability Graph for Stope Design

The stability graph method (developed by Mathews et al. (1981) and modified by Potvin (1988) and Nickson (1992) is widely used by designers to qualitatively assess the stability of open stopes.

The stability number N is defined as:

N = Q' x A x B x C                                                                                             (1)

Where A is the stress factor (ratio of intact rock strength to induced stress), B is rock defect orientation factor, and C is the gravity factor related to orientation of design surface.

See Table 16.3 for summary of NGI-Q’ values.

Table 16.3: NGI-Q' Values for Stope Design

Zone	Elevation	NGI-Q’ (50% level of cumulative frequency)
Mineralized Vein	Above El. 580 m	29
	Below El. 580 m	11
Waste Rock	Above El. 580 m	12
	Below El. 580 m	3

In addition to this NGI-Q values were assessed for the 655 Level. Core runs were isolated over a 45m run length at this level to give an individual rock mass characterization for stoping at this level. NGI-Q values of 15 and 6 were calculated for zones D1 and D3 for stopes at this level.

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16.2.6   Span Length Design – Longitudinal and Transverse

The following key mining parameters / assumptions were used as the basis of the initial stope design:

§   Non-man entry to production level.

§   With an initial assessment of the effect of both hanging wall and the stope back it was found that the hanging wall will govern the stope size. Therefore, the hanging wall effect was considered throughout the design process.

§   Cable bolted stopes and stable stope spans were assessed.

§   Two scenarios were considered for stope length design in longitudinal mining: When the hanging wall is excavated in waste rock, and when mineralized vein is left in place as part of the hanging wall.

§   Stope widths of 7 metres, 12 metres, 17 metres and 22 metres were considered.

Sloughage was considered using the ELOS graph developed by Clark and Pakalnis (1997). This seeks to quantify the degree of dilution anticipated for a given span length and corresponding shape factor.

It must be noted that this graph was developed for weak rock masses. There is no measure of time dependent behaviour but nevertheless this approach provides an indication as to predicted dilution at this stage and should be used in conjunction with observational data during mining. Acceptable sloughage levels of 15% or less were also calculated for stope design utilizing the ELOS criteria for the stability graph after Clark and Pakalnis (1997).

16.2.7   Stope Back Width Design

The stability graph method was utilized for designing the “stable” back length similar to hanging wall stope length design. It can be seen in Tables 17 through 23 in Appendix E that the hanging wall stability is the overall controlling factor for stope stability.

At this Prefeasibility level it may be advisable to consider supporting stope backs greater than 10 metres width with cable bolts at the 536 Level and below due to stress effects. However, as spans will likely vary across levels and also in individual stopes, the magnitude and orientation of the in-situ stress regime is recommended beyond Prefeasibility level to verify stress effects on the mine openings and hence the need for localized support in individual stopes.

16.2.8   Cut and Fill

The dimensions for the cut and fill mining method were estimated using the span design curve (Lang, 1994).

A maximum span of 15 metres is recommended in current cut and fill mining levels. The spans refer to spans which may not require intensive ground support such as timber sets, cable bolts or post pillars,  but  still requires local support within the immediate back to confine potential blocks (loose) which may open due to nearby blasting and/or redistribution caused by subsequent mining activity. 2 metre by 2 metre pattern bolting is recommended in slashed stopes with spans less than these stable back span figures. Due to the nature of the ore body, spans will vary across levels and also in individual stopes. An assessment of the need of support required for individual stopes needs to be undertaken at a stage beyond the scope of this PFS.

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With a 5 metre cut and a maximum stope length of 100 metres dilution from blast damage only along the hanging wall should be expected.

16.2.9   Transverse Long Hole Stoping

With reference to Table 16.4 assuming a 20 metre stope width and a maximum length of 30 metre the stope back should remain stable. Loose from the stope back may be created during the last 10 metre of stope production at 471 m and 536 m Level. Timely mucking and backfill placement is recommended. Stope sidewalls should remain stable. A 10% dilution can be expected from hanging wall sloughage at the 471 m and 536 m Levels. However, as the hanging wall forms the stope end, dilution will be dependent upon the speed at which the stope is blasted and mucked. If low grade mineralized vein forms the hanging wall then expect blast damage only.

Table 16.4: Transverse Stope Design

Level	Location	Average ore dip	Average ore strike	Q'	N	Stope Back	N	Stope Vertical Sidewall	N	Hangingwall ELOS (%) waste	N	Hangingwall ELOS (%) mineralized vein
655	West	55	90	15	11	Stable	45	Stable	6	5	31	Blast damage
	Central	62	72	15	21	Stable	84	Stable	7	4.5	22	Blast damage
	East	47	90	15	11	Stable	45	Stable	5	6	25	Blast damage
597	West	55	90	29	22	Stable	87	Stable	11	3	60	Blast damage
	Central	62	72	29	41	Stable	163	Stable	15	2	43	Blast damage
	East	47	90	29	22	Stable	87	Stable	11	3	49	Blast damage
536	West	55	90	11	8	Transition	33	Stable	3	7	23	Blast damage
	Central	62	72	11	15	Stable	62	Stable	4	7	16	Blast damage
	East	47	90	11	8	Transition	33	Stable	3	7	19	Blast damage
471	West	55	90	11	8	Transition	30	Stable	3	7	23	Blast damage
	Central	62	72	11	14	Stable	56	Stable	4	7	16	Blast damage
	East	47	90	11	8	Transition	30	Stable	3	7	19	Blast damage

16.2.10   Pillar Design

16.2.10.1   Sill Pillar Design

For Prefeasibility Study purposes, sill stress was estimated in Examine 2D in four cases of stope width 7 metre, 12 metre, 17 metre, and 22 metre.

A constant sill pillar thickness was assumed across the entire mining level for practical mining purposes. This constant thickness corresponds to the maximum ore width analyzed. Additionally blast damage of 1 metre was also assumed on each side of the pillar.

Table 16.5 shows the corresponding sill thickness stability above each mine level with the greatest ore span considered having the most effect on stability due to the ratio of width to height.

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Table 16.5: Sill Pillar Stability

Level	Sill thickness (m)	W/H Ratio		Av. Pillar Stress		Pillar Stress/UCS		Pillar Stability Graph
		7m ore width	22m ore width	7m ore width	22m ore width	7m ore width	22m ore width	7m ore width	22m ore width
655	10	1.1	0.4	8.5	10.5	0.10	0.12	Stable	Stable
597	13	1.6	0.5	12	13.5	0.14	0.16	Stable	Stable
536	16	2.0	0.6	21.3	15.8	0.24	0.18	Stable	Stable F.o.S 1.4
471	20	2.6	0.8	17	20.4	0.20	0.23	Stable	Stable F.o.S 1.4
401	25	3.3	1.0	20.6	20	0.24	0.23	Stable	Stable F.o.S 1.4
328	28	3.7	1.2	26	22	0.30	0.25	Stable	Stable F.o.S 1.4

16.2.10.2   Crown Pillar Design

Underground mining will result in a temporary surface crown pillar below the open pit, as well as permanent crown pillars in some areas.  The minimum thickness for crown pillar design was assessed as 25 metres using the empirical crown pillar stability graph (Carter, 1990) taking a typical Q value of 11.6 for the ore at near surface stress conditions. The maximum span of the orebody was taken as 25 metres giving a thickness/span ratio of 1.

16.2.10.3   Rib Pillar Design

Rib pillar widths have been assessed based on two-dimensional analysis utilizing brittle Hoek Brown parameters and the pillar stability chart as used for sill pillar analysis.

For two dimensional modelling a horizontal cross section was taken across the pillar midpoint in order to incorporate the pillar shape factor in the stress analysis. A constant stress field dependent on depth was imposed onto the cross section.

For efficiency purposes at this stage of prefeasibility level study the rib pillar widths were modelled at the mid-point of the mine at 536 m level. Five “fixed” stope span length classes of 25, 50, 75, 100 and 125 metres were modelled. An average pillar sub-vertical height of 56 metres was used during the analysis based on average dip of the ore body across the West, Central and East domains. Rib widths of 15 metres, 18 metres, 20 metres, 25 metres, and 30 metres, respectively, were found to give an adequate factor of safety for these span lengths, as shown in Table 16.6. Stress analysis was conducted on a pillar width 2 metres less than the final pillar width to take into consideration pillar blast damage. An in-situ stress testing and tri-axial strength testing regime is recommended to refine pillar dimensions beyond Pre-Feasibility level.

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Table 16.6: Rib Pillar Stability

Level	Span Length	Pillar Height	Pillar Width	Width/Height	Average Stress	Pillar Stress/UCS	Pillar Stability Graph
536	25	56	15	0.27	13.8	0.16	F.o.S* 1.6
	50	56	18	0.32	17	0.20	F.o.S* 1.4
	75	56	20	0.36	18.5	0.21	F.o.S* 1.4
	100	56	25	0.45	20	0.23	F.o.S* 1.4
	125	56	30	0.54	20	0.23	F.o.S* 1.4

*Factor of Safety

16.3   Proposed Underground Mining Methods

The Santa Elena ore body varies in dip and thickness along strike and at depth. As a result, three well established underground mining methods have been selected for ore extraction. These mining methods are categorized in table 16.7 below:

Table 16.7: Mining Method Selection Criteria

Orebody Geometry	Mining Method
Dip > 50 degrees, Thickness < 15m	Longitudinal Longhole Stoping
Dip > 50 degrees, Thickness > 15m	Transverse Longhole Stoping
Dip < 50 Degrees, Any Thickness	Mechanized Cut and Fill

In general, conventional mechanised mining methods have been selected.  The basis of the development of the mining methods and consequent equipment selection has been that SilverCrest will undertake production drilling, blasting and loading using a contractor for the waste rock and ore haulage to surface.  Initially a contractor will be retained to carry out mine development, with jumbo drill rigs purchased later in the mining life, after which development will be done in-house.  The roles played for operation of the underground mine will be as follows:

1.   SilverCrest will undertake:

a.   General mine infrastructure development,

b.   Mine management and planning,

c.   General underground maintenance and provision of supplies,

d.   Ore drilling and blasting,

e.   Ore loading,

f.   Ore development and tunnelling (after purchase of a drilling jumbo)

2.   Haulage contractor will undertake:

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a.   Haulage of all waste rock and ore to surface.

b.   Backhaul of all backfill rock and tailings to underground.

3.   Development contractor will undertake:

a.   Initial ramp, ore drift and other tunnelling underground

16.3.1   Longitudinal Long Hole Stoping

Longitudinal long hole stoping is a sublevel bulk mining method involving a narrow, steeply dipping vein with competent ore and host rock. Santa Elena will employ longitudinal long hole stoping when the ore has a width less than 15m. Figure 16.1 below shows typical layout and dimensions for sublevel long hole mining.

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Figure 16.1: General Sublevel Stoping Operation

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The stope design done for the purpose of this PFS considers two single drifts for drilling and blasting the ore to develop sublevels every 22.5 m, with a drift developed at the base of the stope for mucking the blasted ore.  This results in stopes of 45 m in height.  The stope spans along strike are dependent on the rock engineering criteria for managing hanging wall dilution.

The long hole mining concept takes advantage of the dip of the footwall, such that the broken ore will run to the base of the stope towards the mucking points. The stope inclination will vary between 50o and 62o depending on which level and domain the mining is taking place.

The sill pillars separate the stoping blocks vertically with the sublevels further dividing each block. Rib pillars will be left between adjacent stopes, separating the stoping blocks horizontally. Stope spans are based on geotechnical analysis by depth and geotechnical domain, see section 16.1.

A detailed geotechnical study based on geotechnical drilling and rock mass assessment was conducted and the results reported the recommended stope dimensions for each mining level and the different domains. These parameters along with the Indicated Resources where used to develop the stope layout.  Where possible, the stopes have been placed in high grade areas, with sill and rib pillars placed in the lower grade areas.  EBA has considered that rib pillar extraction is reasonably possible where rib pillars are located close to the access ramp to limit footwall development, have high grade and with a strike width to pillar thickness ratio of more than 2:1  to ensure that there is reasonable expectation of pillar stability prior to extraction.

Stoping will begin by drilling and blasting the lower sublevel to initiate ore flow and provide an open face into which ore will be blasted. The mining will then continue on the upper level and will retreat towards the cross cut intersection from the ramp, from which the stope is accessed.

16.3.2   Transverse Long Hole Stoping

In areas with a proposed stope thickness of greater than 15 metres, the transverse long hole mining method will be used.  When mining transversely, access drifts will be driven perpendicular to the strike of the ore body from the initial footwall developments. The drawpoints will also be oriented perpendicular to strike and will be accessed from crosscuts which extend from the footwall access drifts to the stope hanging wall. Mining occurs by retreating from the hanging wall to the footwall, initiating ore flow by blasting from the footwall towards free face created against the hanging wall.  To mine the stope, it is divided into primary, secondary and possibly tertiary stopes.  The primary stopes are the first to be mined, with secondary and tertiary stopes mined after backfilling. Note that for the purpose of the PFS, cemented rock fill is being considered to prevent backfill from contaminating the ore while mining the secondary and tertiary stopes.

The primary and secondary stopes will span from the footwall to the hanging wall and are planned to be 20 m wide. Figure 16.2 shows the primary stopes in blue and secondary stopes in green. Once the primary stopes are mined out, they are backfilled using a cemented rock fill, which will be sourced from development and surface waste rock dumps. The secondary and tertiary stopes are then mined between the filled stopes.  The overall stope heights are set at 45 m to allow levels to coincide with longitudinal stopes. A sublevel interval of 22.5 m has been considered, as in the case of longitudinal long hole method.

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Figure 16.2: General Transverse Long Hole Stoping Method

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16.3.3   Mechanized Cut and Fill

Where the dip of ore body results in stopes with a footwall inclination of less than 50°, long hole stoping will not be suitable due to the increased length of blastholes, potential blast hole inaccuracy and the potential for ore to hang up on the footwall.  EBA has thus considered mechanized cut and fill mining for these areas.  The lower east portion of the reserve model, as seen in figure 15.1, has been assessed to have conditions conducive to cut and fill mining.

Though cut and fill stoping can consider large vertical spans, rib pillars and sill pillars have been included in the stope layout to separate stoping blocks and to limit the hanging wall spans for each cut.  Once each cut is mined out, it will be filled with cemented rock fill, which can include tailings and waste rock from ongoing underground development.  Dilution from hanging wall geotechnical failure will be minimal with cut and fill mining because there will be limited exposure to the hanging wall during mining.  Support will be required for the exposed cut back (roof) of each cut during mining, for safety of personnel operating machinery and working in the stopes.

The ore will typically be mined in 5 m high cuts, with each cut silled out from hanging wall to footwall. Each cut will be accessed by attack ramps from the footwall development drifts. Once the cut has been filled all the way back into the attack ramp, the brow of the ramp is slashed into, and another cut can begin. This process will continue until the attack ramp grade becomes too steep for ore haulage. The orebody is then accessed from a higher footwall development. Figure 16.3 below represents the cut and fill method discussed here.

Trucks will backhaul fill material from tailings and waste rock stockpiles on surface, when rock is not available from underground development. This material will be mixed with water and 5% cement to create a consolidated fill, and be placed in the stopes via scoop tram.

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Figure 16.3: General Mechanized Cut and Fill Stoping Method

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16.3.4   Rib and Sill Pillar Recovery

Wherever economically feasible, sill and rib pillars will be recovered. Specifically, the rib pillars between the 597-3 and 597-2 stopes and the 597-2 and 597-1 stopes have been estimated to contain high grade ore and are accessible with minimal additional development, see figure 15.1 The sill pillar above the 597-4 stope also includes high grade ore, which could be extracted based on backfilling of the 597 level stope.

16.3.5   Mining Equipment

Table 16.8 shows the mining equipment selected for the underground mine.  The purchasing of the equipment has been scheduled based on when the equipment will be required in the capital cost estimate schedule (appendix F), with consideration for equipment delivery lead times.   Note that the mine trucks and one jumbo drilling are based on contractor equipment hire.

Table 16.8: Mining Equipment List

Description	Number	kW	Ownership
Mine trucks (20 ton)	6	250	Contractor
Long hole drill (Sandvik)	1	130	SilverCrest
Scoop trams (Sandvik)	3	200	SilverCrest
Jackleg/stoppers (based on 200 cfm air consumption)	4	37	SilverCrest
Jumbo drills (cut and fill stopes) DD321-40 (110kW)	2	110	Contract and SilverCrest
Anfo loader (will require compressor)	1	NA	SilverCrest
Scissor lift 92 HP	1	69	SilverCrest
Grader - 110 HP	1	82.1	SilverCrest
Kubota/s	3	14	SilverCrest
Crew transport	2	96	SilverCrest
Service truck	1	96	SilverCrest
Shotcrete Pump (Airplaco Shotcrete Mixer/Pump)	1	34	SilverCrest
Underground diamond drill/Gopher (HydraCore Gopher)	2	37	SilverCrest

For other equipment and machinery related to mining refer to section 18.

16.4   Mine Design

For the purpose of completion of this Prefeasibility study, the layout of the Santa Elena underground was completed using Dassault Systemes GEMS 6.5 software. This allowed for the digitization and visualization of the mine stope layout as it evolved through the design process.

16.4.1   Pillar Layout

The underground layout for Santa Elena was developed by initial placement of the sill pillars according to Indicated Resources above 1.5 g/t Au equivalent, rock engineering criteria, the depth of the open pit and the desired crown pillar vertical thickness. Refer to section 16.2 for the geotechnical analysis and geotechnical design parameters and resulting sill pillar thickness.  Subsequently, rib pillars were positioned on each level according to the rock engineering analysis, using the maximum allowable open span for each level as the distance between rib pillars. Where possible low grade areas where targeted for rib pillar placement.

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16.4.2   Stope Layout

Solids were digitised in between the pillars, delineating the mineable area of the underground Indicated Resources. These solids provided the initial stope shapes to calculate internally diluted grade and tonnage. These initial reports highlighted stopes which could be feasibly mined where grades were higher than cut-off grade for the planned mining method, and allowed the exclusion of stopes that had grades below the economic cut-off grade. Where stopes where found to have grades lower than the cut-off grade of 1.5 g/t AUEQ or the stopes grades where only marginally higher than the cut-off grade, the stopes were clipped or adjusted to exclude some low grade materials on the stope edges. In some instances entire stopes were excluded from the reserves as the stope grades were below the cut-off grade and there was too little material to justify reasonable extraction.

The layout of the cut and fill stopes was undertaken to create shapes above cut-off grade in the lower east portion of the mine where the deposit shallows out in dip. More flexibility in the stope shape is possible for the cut and fill stopes, due to the mining method, though as far as possible stope shapes were kept to blocks with strike lengths of a maximum of 100 m to limit the hanging wall spans within the cuts.  Sill pillars are not necessary but have been left to allow the upper most stopes to be mined first This creates an opportunity for stoping to be undertaken concurrently with ongoing development, and thereby the use of waste development muck for cut and fill stope backfill.

The stopes that are situated directly underneath the open pit include material that is considered crown pillar for the duration of mining. These stopes are scheduled for extraction at the end of the mine life. These stopes will be mined by long hole method, but will only require a single development drift on the bottom sublevel for haulage. The other access points for these stopes will be via drop-raises through the bottom of the open pit.

Figures 16.4 and 16.5 show the PFS arrangement of stopes, coloured by mining method. In yellow are the longitudinal long hole stopes, in orange are the transverse longhole stopes and in green are the cut and fill stopes. The black regions are the pillars that were identified as recoverable

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Figure 16.4: Longitudinal View of Stopes, Looking North

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Figure 16.5: 3D View of Stope Layout, Looking Northwest

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16.4.3   Development Layout

Figure 16.6 and 16.7 show the development layout developed for the PFS.

Access to the underground stopes will be via a main ramp, and a secondary ramp called the pit ramp, which will be developed from the bottom of the ultimate pit at an elevation of 675 m a.m.s.l. The main ramp will continue from the current exploration drift which is being driven from surface, east of the open pit. The ramps will be ventilated using a shaft that is driven within the centroid of the ramp with the fresh air carried by bag ducting to the face. A main exhaust ventilation shaft will be developed to the east of the main ramp.  Details of the preliminary ventilation design are in section 16.7 of this report.

The stopes will be accessed from crosscuts driven from the ramps towards the ore body from the closest point on the ramp to the ore body on each level. From these crosscuts, the sublevel developments will be driven.

For longitudinal stopes, the sublevel development drifts are situated within the ore, meaning this material will be sent for processing and recovery of gold and silver if above the marginal cut-off grade of 0.5 g/t AU EQ (to ensure the ore will at minimum pay the processing costs). The stopes will be developed on three sublevels: one abutting the upper sill pillar, one abutting the lower sill pillar, and one halfway between the sills. The lower development will be used as a loading and haulage drift, and the upper two will be used as drilling and blasting drifts.

The transverse stopes will be developed similarly to the longitudinal stopes, except the development drifts must be in the footwall waste rock, and the stopes will be accessed from perpendicular crosscuts. These drifts will serve as entry points through the ore body to the hanging wall, and mining will retreat to the footwall. The drifts into the transverse stopes will be silled out to allow more accurate drilling of long holes between sublevels.

The cut and fill stopes access will be developed in the footwall waste rock. The stopes will be accessed using attack ramps from these main development drifts. As a sublevel is finished, the backfill provides a working floor from which to mine the next cut.

Table 16.9 shows a summary of the linear metres of development over the life of mine, by development aspect.

Table 16.9: Linear Meters of development over LOM

Aspect of underground development	Total length in meters
Waste Development	10,454
Ore Development	6,300
Ramps	5,300
Main ventilation shafts	800

Figure 16.6 and figure 16.7 shows the developments in relation to the planned stopes.  Stope numbering is shown in figure 15.1 and the list of stopes with tonnage and grades is shown in Table 15.4.

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Figure 16.6: Development and Stope Layout, Looking North

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Figure 16.7: 3D View of Stope and Development Layout, Looking Northwest

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16.5   Backfill

For the transverse long hole stoping and mechanized cut and fill mining methods, backfilling is required.  For the transverse long hole stoping the backfill will be comprised of a combination of waste rock mixed with 5% Portland cement by weight to create a consolidated fill.  For cut and fill stoping the backfill does not need to be cemented; however, for the purpose of the PFS cement has been included in the backfill costs to provide for the opportunity to extract rib and sill pillars between cut and fill stopes if it is determined it is safe to do so.

The waste rock will either be transported from surface waste dumps or from adjacent mining stopes. Rock fill will be run of mine and therefore the overall sizing curve will be short on fines, however detoxified mill tailings will also be available for use in backfilling.

16.5.1   Backfilling Cut and Fill Stopes

Mining will begin at the lower most section of each stope by slashing through the width of the ore body in (5 m) lifts. The openings will then be backfilled with cemented rock fill. During mining of each 5 m high cut, support in the form of rock bolts or split sets will be installed in the back and sidewalls as informed by geotechnical conditions and daily inspections. The backfill is placed in order to provide a new working platform for the level above.

16.5.2   Backfilling Transverse Long Hole Stopes

For transverse stoping, primary stopes will be mined leaving behind an adjacent secondary stope of the same width. Primary stopes will be laid out so that temporary pillars are left in place to allow for the safe extraction of all primary stopes. The primary stopes will then be backfilled using cemented rock fill to allow for the extraction of secondary stopes.  The process may be done in two or more phases of extraction to allow time for the cemented rock fill to set prior to extraction of the adjacent stopes.

16.5.3   Backfilling to Extract Pillars

For the purpose of the PFS, EBA has included the extraction of high grade accessible pillars. In the case of rib pillars, the stopes on either side of the rib pillar to be extracted will be backfilled.  In the case of sill pillars the stope below the sill pillar must be backfilled. This will provide the necessary support to allow the rib and sill pillars to be extracted safely. New access drifts will have to be developed from the decline ramps for rib pillar extraction as the sublevel drifts will have been mined out during the initial extraction of the stopes.

16.6   Overall Mining Schedule

The mining schedule describes the estimates tonnages which will be mined from the underground, open pit and the existing heap leach facility to feed the expansion plan process plant.  Table 16.10 shows the combined schedule for the Santa Elena Expansion Project.

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Table 16.10: Summary of Overall Production Schedule

Aspect of operations	2014	2015	2016	2017	2018	2019	2020	2021	Total life of mine
Total tonnes from underground	127,707	392,412	535,520	646,088	708,359	684,604	580,407	245,225	3,920,323
Total tonnes  from old heap  leach	335,426	615,588	472,480	361,912	299,641	323,396	427,593	8,495	2,844,530
Total tonnes from open pit	544,867	-	-	-	-	-	-	-	544,867
Total tonnes processed	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	253,720	8,191,760
Total AU ounces sold	36,173	39,606	29,361	27,557	27,924	53,357	37,535	11,226	262,739
Total AG ounces sold	1,345,248	1,594,643	1,409,639	1,742,741	1,556,867	1,951,279	1,816,118	702,392	12,118,926

The schedule in table 16.10 excludes 882 kilo tonnes of material to be mined during 2013 and processed on the existing heap leach facility.

16.6.1   Underground Mine Production Schedule

For the purpose of the PFS, a preliminary schedule has been developed so that the stope tonnes and grades can be applied to a mining timeline.  The mining schedule has been developed for the stopes in the reserve model and for the required development to access the stopes throughout the life of mine.

The mining schedule has been developed based on the parameters as shown in Table 16.11 below.

Table 16.11: Scheduling Parameters

Scheduling parameters	Value
Maximum daily production from each longitudinal long hole stope	750 tonnes per day
Maximum daily production from each transverse long hole stope	1,000 tonnes per day
Maximum daily production from each cut and fill stope	350 tonnes per day
Heading/ drift advance rate per blast round	3.8 m
Maximum development advance per jumbo drill with one heading available	1.5 rounds

Based on the above parameters the mine stoping and development schedule was developed by targeting high grade stopes for the earlier in the schedule, where possible and ensuring that at all times at least 3 stopes are available for production, to allow flexibility in the mining operation.  Additionally, mechanised cut and fill stoping areas have been scheduled for the end of the life of mine. The development schedule was established by ensuring that the required development for all stopes operating in a particular time period was completed in advance of production commencing.  Allowance has also been made for a ramp up period.

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After completion of the schedule, verification has been undertaken to ensure that the ore tonnes and grade in the schedule are consistent with the Probable Reserve numbers for underground as shown in section 15.

Mining costs applicable to various types of stoping have been applied to each stope in the Probable Reserves and scheduled.  Development costs are included separately as the timing of development of a stope and mining of a stope may be months apart and thus need separate consideration.

Table 16.12 shows the underground mining schedule developed for the PFS.

Table 16.12: Underground Mining Schedule

Aspects	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	Totals
Ramp development	1,005	742	-	542	1,758	-	584	617	-	-	3,501
Waste development	1,005	1,178	506	1,018	3,331	2,221	2,177	772	429	-	10,454
Ore development	-	-	1,652	2,222	583	1,029	830	-	-	-	6,316
Stope development tonnes	-	-	71,863	66,343	12,016	149,270	183,731	-	-	-	483,223
Stope develop AU grade	-	-	2.08	1.28	0.58	0.93	0.89	-	-	-	6
Stope development AG grade	-	-	93	91	41	91	81	-	-	-	398
Stope mining tonnes	-	-	55,844	326,069	523,504	496,818	524,628	684,604	580,407	245,225	3,437,100
Stope grade AU	-	-	2.58	2.38	1.27	1.07	1.05	2.32	1.70	1.52	14
Stope grade AG	-	-	138	135	92	106	87	116	120	131	924
Total tonnes from underground			127,707	392,412	535,520	646,088	708,359	684,604	580,407	245,225	3,920,323

The mining schedule results in grade and tonnage performance as shown in Figure 16.8. Peak production is reached in year 6, when both access ramps are actively producing.  Mechanised cut and fill mining in the lower portions of the mine have been scheduled later in the life of mine.

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Figure 16.8: Summary of Annual Tonnes by Underground Mining Methods and Grade for the Life of Mine

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16.6.2   Open Pit Mining Schedule

The remaining ore reserves as shown in section 15, within the open pit reserves will be mined from the 30th of April 2013 until completion in July 2014.  In terms of Santa Elena’s current production schedule for 2013, 881,843 tonnes will be mined in 2013 and the remaining 544,867 will be mined in 2014.

16.6.3   Heap Leach Mining Schedule

The heap leach mining schedule has been created in such a way as to increase mill feed to mill capacity for each year of operation. The heap leach resulted in a tonnage of 2.8M tonnes at an average grade of 0.66 g/t Au and 33 g/t Ag for the life of mine. This schedule is not optimized and increases in grade are possible once open pit heap leaching is complete by Q1 2013.

16.7   Ventilation

A preliminary ventilation study was performed for the Santa Elena Mine.  The purpose of the ventilation study was to determine the ventilation requirements in terms of airflow and circuits, as well as provide an estimate of ventilation equipment, infrastructure and the excavations required.

16.7.1   Ventilation Requirements

Ventilation requirements were based on a common industry minimum of 0.06 m3/s per kW of equipment.  Based on the equipment usage at the mine, the total air requirement is 146 m3/s.  Table 16.13 shows the estimated power and utilization of the underground equipment.

Table 16.13: Equipment Power and Utilization

Equipment	Engine	No.	Utilization	Required
	kW	in use	%	m3 /sec
Mine trucks (20 ton)	250	6	100	90
Long hole drill (Sandvik)	130	1	6	0
Scoop trams (Sandvik)	200	3	1 00	36
Jackleg/stoppers (based on 200 cfm air consumption)	37	4	1 00	9
Jumbo drills (cut and fill stopes) DD321-40 (110kW)	110	2	6	1
Scissor lift 92 HP	69	1	25	1
Grader - 110 HP	82	1	20	1
Kubota/s	14	3	33	1
Crew transport	96	2	20	2
Service truck	96	1	20	1
Shotcrete Pump (Airplaco Shotcrete Mixer/Pump)	34	1	50	1
Underground diamond drill/Gopher (HydraCore Gopher)	37	2	50	2
Total	1,270			146

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Operational constraints guided the ventilation design. In particular, hauling will be performed on the lower drift of each stope.  LHD’s are a large proportion of the overall ventilation load, and the ventilation has been planned to avoid ventilation doors on haulage levels.  As a result, air flow has been routed to reach the haulage drifts last.

16.7.2   Ventilation Model

16.7.2.1   Conceptual ventilation design

In order to undertake the ventilation design, conceptual design work was undertaken to create parameters for friction for various types of conduits.  This resulted in conceptual designs for ventilation as shown in Tables 16.14 through 16.17.

16.7.2.2   Intake airway / Escape way Raise

This arrangement may apply to the intake fresh air raise, as is currently being excavated along with the main ramp. Figure 16.10 below shows a partition within the raise, separating the man way from the airway itself. The complete raise has dimensions of 2.13 m by 3.05 m. This raise was sized to provide ventilation, and an escape way route, for the ramp development only (3 trucks, 1 scoop, 1 jumbo, 1 Kabota). It is not meant to be the sole conduit of fresh air for the mine when in production.

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Figure 16.9: Escape Way Raise

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Table 16.14: Design Parameters for Escape Way Raise

Friction Factor (K)	100	x10-10 lb.*min2/ft4
Friction Factor (K)	0.019	kg/m3
Perimeter (P)	7.9	m
Length (L)	152	m
Area (A)	3.9	m2
Air Flow (Q)	61.4	m3/s required
Velocity (V)	15.75	m/s
Static Pressure (Ps)	1426	Pa
Velocity Pressure (Pv)	152	Pa
Total Pressure (Pt)	1578	Pa
Fan Power	97	kW
Fan Power @ 70% efficiency	138	kW

16.7.2.3   Fresh Air Raise Head Without separate man way

This arrangement applies to a fresh air raise without a separate man way. The dimensions used to study the discussed fresh air raise are for a square with 3.4 m sides (Figure 16.11).

Table 16.15: Design Parameters for Fresh Air Raise

Friction Factor (K)	70	x10-10 lb.*min2/ft4
Friction Factor (K)	0.013	kg/m3
Perimeter (P)	13.6	m
Length (L)	455	m
Area (A)	11.6	m2
Air Flow (Q)	146	m3/s required
Velocity (V)	12.63	m/s
Static Pressure (Ps)	1109	Pa
Velocity Pressure (Pv)	98	Pa
Total Pressure (Pt)	1207	Pa
Fan Power	176	kW
Fan Power @ 70% efficiency	252	kW

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16.7.2.4   Raised Bore for main exhaust ventilation airway

Table 16.16 below shows the design parameters and resulting fan power required for an exhaust raise bore with a 3.3 m diameter.

Table 16.16: Design Parameters for Exhaust Raise

Friction Factor (K)	27	x10-10 lb.*min2/ft4
Friction Factor (K)	0.005	kg/m3
Perimeter (P)	10.4	m
Length (L)	455	m
Area (A)	8.55	m2
Air Flow (Q)	146	m3/s required
Velocity (V)	17	m/s
Static Pressure (Ps)	805	Pa
Velocity Pressure (Pv)	178	Pa
Total Pressure (Pt)	983	Pa
Fan Power	144	kW
Fan Power @ 70% efficiency	205	kW

16.7.2.5   Vent Tubing Head For Tunnels and Ore drifts

The design parameters covered in table 16.16 below are calculated for the infrastructure needed to achieve the airflow required at the working faces. The friction factor used is for a smooth plastic fabric, which would be most representative of the material used for the tubing throughout the tunnels and ore drifts. The results take into account both static and velocity pressure head as well as any coupling and exit losses. The size of the tubing was chosen to be 1.07 m in diameter.

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Table 16.17: Design Parameters for Vent Tubing

Friction Factor (K)	20	x10-10 lb.*min2/ft4
Friction Factor (K)	0.004	kg/m3
Perimeter (P)	3.4	m
Length (L)	305	m
Area (A)	0.89	m2
Air Flow (Q)	14.17	m3/s required
Velocity (V)	15.84	m/s
Coupling Loss Equivalent Length	34.76	m
Exit loss Equivalent Length	30.49	m
Number of 90o Bends	0	m
Bend Loss	0	m
Total Theoretical Length of Tubing	370	m
Static Pressure (Ps)	1292	Pa
Velocity Pressure (Pv)	154	Pa
Total Pressure (Pt)	1446	Pa
Fan Power	20.5	kW
Fan Power @ 70% efficiency	29	kW

16.7.2.6   Setup of ventilation model

Ventsim Visual Advanced (Ventsim) was used as the ventilation modeling software.  Ventsim provides a number of features useful for underground ventilation design, including fan specification, and airflow modeling throughout the mine workings.  The software utilizes the Hardy-Cross method to iterate to a ventilation solution.  Minimum inputs to the software are a tunnel system, surface connection, and air moving device, such as a fan.

Initially, solid files were imported from the mine design software as centerlines of the mine workings.  Tunnels were constructed around the centerlines, and then set as their actual type.  Group properties were set for each of the tunnel types.  The tunnel types have specific shapes, as well as friction factors. Tunnel types and their properties are shown in Table 16.18.

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Table 16.18: Dimensions and Friction Factors for Underground Workings

Type	Dimension 1	Dimension 2	Shape	Material	Friction Factor
Drifts	4	4	Square	Average Blasted	0.01
Ramps	4	4	Square	Average Blasted	0.01
Duct	1.07	1.07	Round	Smooth Plastic Duct	0.00371
Vent Shaft	3	3	Round	Bored Raise	0.005
Stope Raises	2	2	Round	Average Blasted	0.01
Stope Access	4	4	Square	Average Blasted	0.01

The ventilation model is sensitive to an accurate friction factor being chosen.  The friction factors ultimately chosen are shown in Table 16.17.  After the initial setup of the model, three phases were created for ventilation planning purposes only.  These phases are representative of early, mid, and late mining stages.  Figures 16.12, 16.13 and 16.14 show the ventilation phases and the ventilation requirements for each phase.

Ventsim supports a forced airflow method that allows a specific required airflow to be placed where a fan would be placed, and the required fan pressure, and then models the airflow throughout the mine.  A forced airflow of 146 cu.m/s was selected from the initial ventilation calculations, and used throughout the modeling as the target flow rate.

A final step was connecting the various workings to the ramps and drifts with attack ramps and small ventilation raises.  These were selected for ease of ventilation modeling and not mining feasibility.  As a result, the actual location of access ramps into the drifts is not reflected in the Ventsim model. Additionally, some of the drifts will require in-place ducting to supply air to the working face.

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Figure 16.10: Phase 1 Ventilation Air Flow Diagram

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Figure 16.11: Phase 2 Ventilation Air Flow Diagram

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Figure 16.12: Phase 3 Ventilation Air Flow Diagram

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16.7.3   Interpretation of Ventilation Conceptual Design

Based on the three sample phases, Santa Elena can be ventilated by one main exhaust fan. During development under the pit, auxiliary fans and ducting will be required, until the under pit developments are attached to the main developments.  Ventilation velocities are reasonable in the ramp and shafts, and airflow is adequate at the most distant drifts.

It may be advantageous to use two main fans, to provide redundancy in the event of mechanical failures.  As well, a stench gas system should be installed as warning to evacuate the mine in case of emergency.

16.8   Open Pit Mining Method

Santa Elena is a conventional open pit. All drilling, blasting and mining is completed by a local Mexican mine contractor utilizing a fleet of:

§   4 – 40 tonnes Caterpillar 740 articulating trucks.

§   3 – 70 tonnes Caterpillar 769 haul trucks.

§   Support equipment including one D8 and two D9 dozers, two excavators, two air track drills (Tamrock and G. Denver), one 140H Caterpillar grader, explosives truck, and maintenance vehicles.

§   Auxiliary equipment includes a water truck and light vehicles.

The current optimized pit (Tinaja Pit) is designed for an estimated mine life of 4 years which was originally 6.5 years. The design was changed to avoid a strip ratio up to 10:1 in year 3 and prevent excavation with a subsequent high cost diversion of the adjacent Tinaja gulch which has high water flows in the rainy season. With the optimization of the original pit, a portion of the original reserves have been transferred to underground. The overall strip ratio for the Tinaja Pit is approximately 4:1.

Standard operating procedures for the open pit includes:

§   Daily meetings between owner (Santa Elena operations) and mine contractor to review safety, production objectives and ore control for the day.

§   Blasthole drilling of ore is on an average 3.5 by 3.0 metre pattern, 6 metre hole depth with a 1 metre sub-drill. Blasthole drilling of waste is on an average of 4.0 by 4.0 metre pattern.

§   Blastholes are loaded with mixture of ANFO or emulsions depending on standing water in the hole. Standing ignition is used for basting and fragmentation.

§   Typically, a blast pattern consists of 50 to 100 blastholes with blasting occurring on average 3 times per week.

§   Blastholes considered to be in or near ore are systematically sampled and sent to the onsite lab for gold, silver and copper analysis. Turnaround time for the lab is estimated at 24 hours. Analytical results are given to the geology department and standard ore control is completed using AutoCAD and Gemcom software.

§   Ore control is flagged and cut lines are established in the pit for extraction of ore and waste.

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§   Daily surveys are completed in the pit to determine volumes extracted and for floor grade control. Truck counts are also completed by the mine contractor and owner personnel. Invoicing for the contractor is based on a cost per tonne with the reconciliation of surveys and truck counts.

§   Mine production is guided by a pre-set mine design that is considered geotechnically adequate to extract materials while operating in a safe manner.

§   Blasted ore from the pit is hauled to the nearby crushing facility and either end dumped into the primary jaw crusher or stockpiled for loader feed crushing. As of April 31, 2103, an estimated 2.84 million tonnes of ore have been extracted from the pit delivered to the crusher and leach pad.

§   Blasted waste rock from the pit is hauled to the waste dump on the northwest of the pit. As of April 31, 2013, an estimated 10.1 million tonnes of waste have been excavated and delivered to the waste facility (see Figure 18.1).

16.9   Spent Ore Mining And Rehandling Methods

Spent ore will be delivered to the mill using of a 3.5 cubic meter loader and a conveyor belt (grasshopper and stacker). A loader will load the spent ore from the leach pad on to a 24 inch wide conveyor belt capable of handling 220 tonnes per hour to be delivered to a stockpile area for blending with underground ore using a reclaim tunnel at the processing facility. Grade control will be performed on spent ore as using a belt sampler. The amount of spent ore required to be processed is flexible. For this report a 50/50 blend is used to maintain a nominal 3,000 tonnes per day feed to the processing facility (1,500 tonnes per day underground ore and 1,500 tonne per day spent ore from the leach pad).  Refer to figure 17.2 for a layout of the plant and 16.15 for the overall site layout which shows the heap leach facility in relation to the plant site and ROM stockpiles.

Figure 16.15 Layout of the open pit and waste dump on April 30 2013

16.10   Tailings Disposal

The tailings produced from the processing facility will be filtered and deposited as dry-stack tailings. The tailings handling circuit will have the following equipment:

§   final CCD thickener underflow pumps,

§   filter feed slurry pumps,

§   filter feed stock tank; 3,200 mm diameter × 7,600 mm,

§   three belt filters with a capacity ,

§   filter press conveyor belts,

§   air compressors,

§   tailings conveyor belt,

§   tailings filtrate pump box.

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The final CCD thickener underflow will be the final plant tailings. The thickened tailings will be pumped from the final CCD thickener at a density of approximately 65% solids. Slurry will enter the pump box which will feed the tailings filter feed stock tank. The plate and frame filters will dewater the tailings slurry to a moisture content of about 20%. The dewatered solids will be conveyed to the lined leach pad for aeration and cyanide destruction. Once the cyanide toxicity in the dry tailings is at acceptable levels, it will be transported by conveyor to the waste facility for deposition and reclamation in the future. The filtrate from the filter presses will be collected in a filtrate pump box and will be pumped to the barren solution tank for reuse within the leach circuit.

17.0   RECOVERY METHODS

The ore from both underground and open pit resources will be processed by conventional cyanide leaching technology, shown in Figures 17.1 and 17.2. In addition partially leached material from the existing heap leach operations will be blended with underground ore at a variable rate and reprocessed through the same plant.

Metallurgical test work has demonstrated that Santa Elena ore is amenable to conventional processing.  Test work has also demonstrated that the existing heap leach residue can be further treated to recover most of the residual gold and silver.

Santa Elena ore contains an estimated average grade (open pit, underground and heap leach) of 1.24 g/t Au and 75 g/t Ag and after crushing and grinding can be leached in cyanide to yield approximately 92% Au Recovery and 67.5% Ag recovery. Because of the relatively high level of silver in the ore (and hence solutions) there are advantages and benefits to using traditional CCD and Merrill-Crowe for metal recovery rather than CIL/CIP. The partially leached heap ore yielded recoveries of approximately 52% Au and 29% Ag when crushed to 10 mm and processed on the heap leach.  Since the leach cycle is being prematurely terminated to start the mill on re-leaching after grinding the balance of the metals are recovered to the level expected from new ore.

The process plant has been designed to treat a nominal 3000 tonne per day (tpd) of ore, a mixture of freshly mined material and partially leached heap residue.  The plant has been designed to treat any proportion of these two types of feed.  Because this may include up to 100% of fine crushed material from the heap leach pads it was considered that this would be unsuitable feed for a SAG milling circuit and so it was elected to go to a conventional crushing and grinding circuit.

ROM ore is delivered to a dump pocket ahead of the crushing plant.  Ore is reclaimed from the dump pocket by vibrating grizzly feeder and fed to a 30``x54`` jaw crusher (the one from the existing heap leach crushing plant).  Product is screened and oversize fed to a XL400 secondary cone crusher.  The cone crushed product is again screened and the oversize to a bin ahead of two  XL 400 tertiary cone crushers, operated on close circuit with the same screen.

Crushed ore is placed on an open stockpile above a reclaim tunnel.  Reclaimed heap leach material is collected and conveyed to a second stockpile located above the same reclaim tunnel.  Each stockpile has two feeders under it (total 4).  There is a weightometer between the two piles on the reclaim conveyor and another after the reclaimed material has been combined (as well as a sampler) to allow for accounting of both the fresh ore and heap leach residue.  The stacking belts feeding the stockpiles also have weightometers and cross belt samplers.

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Ore is reclaimed from the stockpiles and fed to a 15’x21’ (4.62 m x 6.77 m) ball mill fitted with a 2200 kW motor. This mill is designed to produce a 100µm (P80) product and the mill circuit is closed by Hydrocyclones. The ore grinding takes place in barren solution and so significant amounts of gold and silver will dissolve in the grinding circuit.   The cyclone overflow is thickened ahead of the leach circuit, and the supernatant solution advanced as pregnant solution or used elsewhere in the circuit depending on the grade.  Lime and Cyanide are also added to the grinding circuit to maintain desired leaching conditions of pH10.5 and 1000mg/L NaCN strength.

The thickener underflow pulp at 55% solids is fed to the leach circuit consisting of five 13.6mx 12.8m high tanks giving a total of 48 hrs. residence time. These tanks are aerated. In a slight departure from conventional practice after the first leach tank the leach slurry can be diluted and washed with lower grade solution before being re-introduced into the leach circuit (tank #2).  This solution change has been shown to boost the leaching rate, reducing the overall leach time required which provides opportunities to maximize capacity at the plant.

The last leach tank discharges to the CCD (counter current decantation) circuit:  This consists of three thickeners in addition to the grinding circuit thickener and the intermediate thickener, making 5 thickeners in all.  A wash water ratio of 3.2:1 is anticipated.

The washed residue from the final CCD thickener underflow is fed to a surge tank ahead of the two tailings belt filters. Filtration of tailings and dry stacking the cake was selected to minimize water usage, and produce stable tailings.  The dry tailings will be conveyed to an intermediate point adjacent to the disposal area (leach pads), and then distributed over the disposal area (waste dump) after detoxification. Provision has been made for cyanide detoxification of residue, but calculation of the likely concentration in the residue suggests it will not be required.  The intermediate transfer point will be sealed and any drainage from the residue collected and returned to the ore process plant.

The pregnant solution is collected in the existing pregnant solution pond. The various qualities of the possible solution (grinding thickener, intermediate thickener and #1 CCD thickener overflows) will be monitored to maintain a relatively high grade solution, with lower tenor solutions used elsewhere in the circuit.

The pregnant solution is reclaimed from the pond and fed to a conventional Merrill-Crowe (MC) recovery circuit.  Solution is first clarified in a vertical leaf precoat clarifier using DE (diatomaceous earth) as the clarifying medium.  The solutions are then de-aerated under vacuum to remove oxygen, and then zinc dust is added to precipitate gold and silver which is removed from the solution by plate and frame filter press. The existing MC plant will be utilized alongside a new MC plant.

The barren solution from the precipitation press is recycled for re-use in the plant as process water. The filter cake is discharged, dried and mixed with flux for smelting to Dore.

The plant also has reagent handling systems.  Lime is added dry to the mill feed belt.  Cyanide is mixed as a 10% solution and distributed in the plant. Flocculent, zinc, anti-scalent, systems are in place as well as systems for Copper Sulphate and Sodium Metabisulphate if required for cyanide de-toxification.

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Figure 17.1: Santa Elena Expansion 3000 tpd Process Flowsheet

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Figure 17.2: Santa Elena Expansion Plan Plant Mechanical General Arrangement

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18.0   PROJECT INFRASTRUCTURE

The project infrastructure is shown in figure 18.1

The Santa Elena open pit heap leach mine was constructed in late 2009 and 2010, and has been operational since 2010. There are a number of facilities currently in use at the Santa Elena site.

The Santa Elena Expansion Project will be undertaken within the current infrastructure. Broadly speaking the Expansion requires the addition of a processing plant and facilities for the underground mine. The same infrastructure facilities utilized by the open pit mine will continue to be used for the expansion.

Current facilities at the mine consist of (see Figure 18.1 for general site plan):

§   7 km long main access road from paved highway and local community of Banamichi,

§   2,500 tpd ore open pit mine utilizing a mine contractor,

§   Waste dump with the capacity of an estimated 35 million tonnes,

§   3-stage crusher provided by Excel Machinery of Amarillo, Texas,

§   Lined and certified leach pad designed by Vector Engineering of Denver, Colorado,

§   Lined and certified barren and pregnant solution pond designed by Vector Engineering of Denver, Colorado,

§   Lined and certified emergency pond designed for 100 year event,

§   Merrill Crowe plant and refinery,

§   On-site laboratory for production and exploration work,

§   Administration office,

§   Maintenance shop for mine contractor,

§   Diesel generators, and

§   All required piping, power and security.

The material on the existing heap leach facility will be removed, and there is space on the facility for rehandling of the tailings prior to transport to the waste dump as dry stack tailings.

In January of 2012, the expansion of the mine from an open pit heap leach operation to an underground mill operation commenced with ground breaking of the underground portal. As of April 31, 2013, the expansion was approximately 30% complete with all major equipment purchased and the completion of all earthworks for the new processing facility, tank construction underway, and approximately 1,300 metres of underground ramping and development.

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Figure 18.1: Site General Arrangement Plan

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18.1   Surface Infrastructure for the Expansion Plan

18.1.1   Access Roads

The access road to the Santa Elena mine is a 7 km unpaved road, which connects to a nearby paved highway and the local community of Banamichi. This road is currently operated and maintained by SilverCrest.

18.1.2   Offices, Cafeteria, Warehouse and Change Houses

The existing offices, cafeteria, warehouse and change house at Santa Elena are expected to be adequate for the Santa Elena Expansion Project. Currently, the office space consists of a main office located where the access road enters the mine area and a smaller mobile office for the surveying and the underground manager/s.  The process plant being constructed will include office space for the technical staff operating the processing facility.  No additional office space is required for the expansion plan.

18.1.3   Fuel Storage Facility

The diesel storage requirements for the mining machinery for the open pit mine will be sufficient for the underground mining equipment. The consumption of each individual equipment type is highlighted in table 18.1 with the total daily diesel consumption also listed.  Diesel for power generation will be stored at the power generation site/facility.

Table 18.1: Diesel Consumption (Litres) per day

Description	Number	kW	Estimated Hours            Operated/day	Fuel Consumption/Day
Mine trucks (20 ton)	6	250	16	2851.2
Long hole drill (Sandvik)	1	130	1.5	23.2
Scoop trams (Sandvik)	3	200	16	1140.5
Jumbo drills (cut and fill stopes) DD321-40 (110kW)	2	110	1.5	39.2
Scissor lift 92 HP	1	69	2	16.4
Grader - 110 HP	1	82	4	39
Kubota/s	3	14	2	10
Crew transport	2	96	2	45.6
Service truck	1	96	3	34.2
			TOTAL	4199.3

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18.1.4   Water storage facility

A water storage facility is located as shown in Figure 18.1, no additional water storage is required for the expansion plan.

18.1.5   Electrical Distribution

The underground mine will received power from surface generators supplying power to the process  plant and the underground mine.  The overall installed capacity will be in the order of 6.5 MW.  Roughly 1 MW of the installed capacity will be required for underground.  The generators will produce power which will feed a surface substation, from which power will be drawn for the process facility and the underground.  A 5 kV cable will take power underground, either via the fresh air raise or via a borehole.

18.1.6   Explosive Magazine

Santa Elena currently has powder magazines for the open pit mining.  This will be used for surface bulk storage of explosives for the underground operation.  Underground temporary storage of explosives will be undertaken in suitable areas or containers, as directed by Mexican mining health and safety regulations.

18.1.7    Maintenance Shop

A maintenance shop exists for the underground machinery used in the ramp development.  This shop is adjacent to the current underground adit as seen in figure 18.1 above. In addition to surface maintenance shops, and underground maintenance shop will be constructed on the 536 level.  This shop will consist of 3 to 4 bays, which have been slashed to accommodate an underground truck with the load body in the tipping position.  The underground shops will be fitted with overhead trolleys for lifting of heavy equipment and will include tools and equipment as required for maintenance.

18.1.8   On-site Laboratory

The existing laboratory used for the open pit will continue to be used for the underground mining.

18.1.9   Cement Silo

Santa Elena may require a cement storage area. This cement may then be added to waste rockfill and mixed in transport trucks on their way to the open stopes. The rate of backfilling will vary with the production rate and the type of mining undertaken.  The maximum backfill requirements will occur when mostly transverse long hole stoping is being carried out from year 6 to year 8 of mining.  The volumes will average around 360 to 400 m3 of backfill per day. The daily requirement for Portland Cement has been estimated at around 32 tonnes. Storage for cement is sized in a manner such that there will be at minimum a week’s worth of cement on site to feed the backfill requirements. At this rate the weekly storage capacity should be 225 tonnes. EBA has estimated that a 300 tonne silo will be required for the cement storage.  Cement will be delivered to site as required to maintain the storage inventory.

18.2   Underground Infrastructure

Additional facilities will be required to support the mining operations underground. SilverCrest has completed the main ramp from the adit at 780 elevation to the 625 level as of the end of April 2013.  The current ramp includes muck bays, sumps and a fresh air raise to surface.  This arrangement will be continued to the 536 level after which a return airway will be raise bored from surface to create a ventilation circuit. In addition to the ramp, cross cuts and drifts required for the mining of the ore, further infrastructure is required as stated below.

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18.2.1   Underground Maintenance Shop/Warehouse

The underground shop will be situated on the 536 level, in a more central location where equipment from all areas of the mine can easily access the shop. The shop will include 3 bays with dimensions to allow for work on scoop trams, jumbo drills, and underground haul trucks. The bays will accommodate scoop trams with a raised bucket. Next to the maintenance shop will be a bay equipped with shelving to store frequently used parts as well as other parts on consignment, consumables, and tires. A concrete floor will be placed in working areas in the workshop to enable working on level surfaces and for containment of spilled contaminants.

18.2.2   Refuge Station

A prefabricated refuge station will be located in an unused drift or a muck bay to serve as an emergency shelter, as required by health and safety regulations.

18.2.3   Dewatering

All development headings will be sloped at a 2% gradient towards the ramps to allow for water to drain in a controlled direction. Sumps will be developed on each level to collect any inflow water and from there the water will be either recycled and used as drill water, or pumped to surface. The sumps will be installed at a vertical interval of 60 to 70 m.  A conceptual arrangement for the dewatering of the mine is shown in figure 18.2, equipment and flow rates are expected to very considerably over the life of mine.

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Figure 18.2: Conceptual Water Balance

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18.2.4   Underground Electrical Installations

A conceptual layout of the underground electrical installations up until the 536 level is shown in Figure 18.3.

The site currently has no grid power and the use of diesel power generators will continue for the Expansion of the project. The power generation capacity will be increased to supply power for the process plant and the underground mining requirements.  The underground power consumption has been estimated at 800 kW effective power, requiring an estimated 1MW of installed power generation capacity.  The primary use of power underground will be by ventilation fans, drills, lighting plants, sump pumps and the workshop, while the majority of power used on the overall Santa Elena site will be by the crusher and process plant. To extend the services to underground, a 5 kV high tension power cable will be run down the ventilation raises or a dedicated borehole, connecting to a substation on each of the active levels. The underground machinery will be run on 440 V, with step down transformers mounted on skids delivering the power to the levels as required.  Where required, smaller transformers will step down the voltage for underground lighting, shops and lunch rooms. All electrical installations will be done under supervision of a qualified electrician.   The electrical equipment required for underground will include:

§   Underground cable vacuum breaker

§   Main underground cable from substation to adit – 5kV

§   Cable from adit to ventilation shaft collar – 5kV

§   Ventilation fan transformers and switchgear

§   Main underground cable – 5 kV

§   Underground power centres mounted on skids including circuit protection, transformers, 5 kV feed couplings and 480/440 /120 V output couplings

§   480/440 V underground feeder cables

§   Surface cap lamp charger stations

§   Miscellaneous electrical equipment including distribution boards, lighting and low voltage cables

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Figure 18.3: Conceptual U/G Electrical Distribution Diagram

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18.2.5  Water Services

Underground water requirements include both potable and non-potable water. Non potable water for the jumbo and long hole drills will be sourced from sumps on each level if possible, or from surface water storage tanks. A service water pipeline will be provided down the ramp or down the ventilation shafts as required.  Potable water will be taken down in bottles for each shift.

18.2.6   Powder Storage

The main explosives magazine will be on surface. A temporary storage area for a short term supply will be developed on some of the levels underground. The underground explosive magazines will hold enough charge and detonators to be able to supply the underground workings for 1 week.

18.2.7   Compressed Air

A mobile compressor will be used to provide compressed air for jackleg drilling power as well as ANFO loading.  This compressor will be powered by electricity.

18.3   Processing Plant Infrastructure

The major equipment to be installed by January 2014 for the processing plant is shown in table 18.2 (see Figure 17.1, 17.2 and 18.1 for details and layouts for the processing plant);

Table 18.2: Major Equipment Components as part of the Processing Plant

Class	Vendor	Item
Crushing & Screening	FLS	Cone Crushers
	FLS	Vibrating Screens (Secondary & Tertiary)
	BATSA	Conveyor Belts
	FLS	Vibrating Pan Feeders
	PPI Pella	Conveyor Idlers
	THERMOSFISHER	Conveyor Belt Sampler
Grinding & Crushing	FLS	Ball Mill
	FLS	Cyclones
Leaching & CCD	Outotec	Thickeners
	Mixtec	12 Agitators
Recovery (MC Plant)	FLS	MC Plant + Refinery
	FLOWSERVE	Process Water Pumps
Warehouse & Management	FLS	Slurry Pumps
	Kolberg Pioneer, Inc.	Belt Feeders
Reagents	BASF	Flocculant Mixing System
	Process IQ	Wad Analyzer

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Class Vendor Item

Thickening & Tailings Management	FLS	Horizontal Belt Filters
Auxiliary Buildings	LYM (IR)	Compressor Air Equipment
Generators	WWWILLIAMS	Power House Project Gear
	WWWILLIAMS	7-E9000 Evolution LV MCC
	GE	Switchgears
	CONDUMEX	Transformers
Other Equipment	BATSA	Grasshoppers/Stackers
	Nixon Grua	Crane 2004 Link Belt RTC
		Dozer

18.4   Waste Rock and Tailings Storage Facilities

The waste rock storage facility at Santa Elena has been designed for a capacity of 35 million tonnes and fully covers all planned open pit and underground waste and dry stack tailings. Of the 35 million tonne capacity, 20 million tonnes has been used for the open pit waste rock, leaving roughly 15 million tonnes of capacity for underground waste rock and tailings.  Note that not all waste rock from underground will be deposited at the waste dump sites as roughly 14% will be used to backfill some of the underground stopes.  Figure 18.1 above shows the waste dump location and design.  Table 18.3 below shows the estimated volumes of waste rock and tailings which will be placed on the waste dump for the Santa Elena Expansion Project.

Table 18.3: Waste Rock Volumes

Source of waste rock	Volume in m3	Tonnes
Open pit waste stripping 2014	1,625,000	2,600,000
Underground ramp development LOM	91,026	145,641
Underground waste development LOM	281,228	449,965
Open pit ore tailings 2014	417,944	668,710
Underground ore tailings LOM	2,450,202	3,920,323*
Heap leach reprocessing ore tailings LOM	2,251,581	3,602,530
Total capacity required	7,116,981	11,387,169
Less underground backfill capacity	1,002,889**	1,604,622
Required capacity on waste rock storage facility	6,114,092	9,782,547
Remaining capacity on the waste rock dump	8,000,000	12,800,000
Excess capacity	1,885,908	3,017,453

* Includes heap leach reserves up to April 30 2013 plus all open pit ore mined from April 30 2013 to December 30 2013

** Based on cut and fill stopes, transverse stopes and stopes on 597 level (rib pillar removal on 597 level required the adjacent stopes to be backfilled)

Tailings from the processing facility will be washed, filtered to approximately 18% moisture content, drained on an exposed portion of the existing leach pad for at least week and conveyed for dry stacking on top of the ROM waste rock dump.   Removal and cyanide detoxification will be achieved in combination with multiple filtering, a wash cycle and photo-degradation on the leach pad prior to be conveyed to the waste dump.  The estimated tonnage of tailings to be delivered to the waste dump will depend on the plant processing capacity, with a planned nominal throughput of 3,000 tpd.  The tailings are estimated to have a bulk density of 1.61 t/m3.

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The infrastructure required for handling the dry stack tailings that has been included for the PFS include an overland conveyor to transport the material to the waste rock storage dump and a loader to rehandle the material from the heap leach pad onto the over land conveyor.  Note that the existing bull dozer used on the current heap leach facility is of sufficient size to move 3,000 tonnes of dry stack tailings per day on the waste rock dump.

It must be noted that additional capacity will also be available for detoxified tailings material on the current heap leach facility, later in the mine life, when a significant portion of the material on the existing heap leach has been reprocessed.  No further waste rock storage facilities are foreseeably required.

19.0   MARKET STUDIES AND CONTRACT

19.1   Sales of gold and silver doré

EBA is satisfied that SilverCrest currently sells gold and silver from the open pit operations and therefore an assessment of the market for silver and gold is not regarded as material to this PFS.

EBA has included costs for refining, transportation and customs expenses for selling of the mine poured doré bars within the economic analysis presented in Section 22.

20.0   ENVIRONMANTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

20.1   Land Use Change Permit

Nusantara is authorized by SEMARNAT (May 2, 2013) from the environmental impact assessment view to develop the Project Expansion of the Santa Elena Mine which consists on adding 71.12 ha surrounding the already authorized surface of 99.35 ha for a total of 170.47 ha. The authorization is valid for 10 years from the next day of the day of reception of this document. It may be extended by written request within 30 days before the date of expiration.

The permitted new construction includes: portal and ramp, diesel gens, crusher and mill, tanks construction for leaching by CCD for underground mining, Merrill Crowe area, stacking area for ROM material, offices, mill shop and warehouse and internal road network to communicate the area of underground process.

20.1.1   Summary of ARD/ML Work for Santa Elena

Geochemical characterization of mine waste and geochemical modelling has been undertaken at the Santa Elena Mine and includes consideration for acid rock drainage (ARD) and metal leaching (ML) potential. This work was undertaken by Global Resource Engineering Ltd. (GRE, Breckenridge, 2009). Testing was completed in two phases, the first in 2006 and the second in 2009. Phase II of the program was based on preliminary findings in Phase I and includes additional analyses and measured analytes. The results for both programs are typically reported together without distinction for phase.

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Mine waste material was characterized by a series of static tests, including acid base accounting (ABA), whole rock metals analysis and Synthetic Precipitation Leaching Procedure (SPLP). Testing was completed on a variety of lithological units defined as andesite. Apart from the mineralized andesite ore, the mine waste material considered for geochemical characterization include the following:

§   Geo-thermally altered andesite near the mineralized zone - Andesite within the sulfide-rich halo.

§   Unaltered andesite distant from the mineralized zone - Andesite outside the mineralized halo

§   High calcite andesite

ABA testing has been completed on 42 samples. Results indicate that the andesite units inside and outside of the sulfide-rich halo are classified as potentially acid-generating (PAG) and that the andesite in the calcite-rich zones is net neutralizing. Samples from the sulfide-rich halo zone show a strong variance in net neutralization potential, with variable levels of sulfide and calcite present.

Whole Rock Metals analysis was completed on 14 samples. Results indicate that there is a significant difference in calcite concentrations between the various lithological groupings, and no significant difference in sulfide concentrations. Manganese was noted in elevated concentrations in select samples, and has also been noted in elevated levels in groundwater samples from site.

SPLP testing, with variable analytes measured, has been completed on a total of 42 samples. Results indicate that only concentrations of aluminium in the acid leachate are elevated. Key indicators of alkaline rock drainage, including arsenic, selenium and zinc were not noted at high concentrations.

Based on the results of the static testing, kinetic test cells were established to evaluate reaction kinetics and the potential time to onset of acidic conditions. Six samples were composited to form two kinetic test cells.

Geochemical modelling was undertaken to predict ARD/ML potential for the mine waste material over an extended period of storage on surface at the mine site. The geochemical model includes consideration for precipitation, surface water flow, reaction kinetics and geochemical reactions.

Kinetic test result and the experience of storing mine waste on surface at the mine site indicate that geochemical reactions at the site are slow. This is in large part due to the water supply and flow dynamics of the site. It is not expected that the mine waste material will generate acidic leachate or leachate with elevated metal concentrations. The current geochemical risk mitigation plan includes infiltration management to manage any leachate and runoff that may develop. The mitigation plan includes continual testing and monitoring as well as concurrent reclamation.

20.2   Closure Planning

An independent Closure and Mine Reclamation Plan was created for the Santa Elena project in March, 2010, by Global Resource Engineering Ltd. (Breckenridge, 2010).  This plan incorporated study results from baseline environmental impact, water quality and geotechnical stability studies for the original open pit, processing and waste dump.  Closure plan activities included, but were not limited to closure and capping of the waste facility, closure and capping of the spent heap leach facility, disposal and/or treatment of hazardous site chemicals and impacted soils, and re-contouring of impacted grounds.  A closure cost of $USD 2.28M (NPV projected to 2017) was included with a 10% contingency in place.

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EBA has included consideration for mine closure, remediation and ongoing monitoring and stewardship activities within the economic model at an estimated value of $USD 5M.  A revision to this closure plan is recommended.

21.0   CAPITAL AND OPERATING COSTS

EBA has estimated total capital costs remaining of $88 million, which excludes $24.2 million spent on the project as of the 30 April, 2013.  The total remaining capital is made up of $48 million initial, required for construction of the plant, initial underground mining equipment purchases and preproduction underground development, and a further $40 million sustaining capital, which includes ongoing development, ongoing mining equipment purchases and ongoing infrastructure requirements.  EBA has included an estimated of contingency of $9.3 million over the life of mine the Santa Elena Expansion Project.

EBA has estimated total operating costs at an average of $39 per tonne of ore processed.

The sources of information for these costs include:

§   Overall cost and equipment and materials for the processing plant

§   Underground design work for pre-production development costs

§   Quotes from equipment suppliers

§   Equipment operating costs from equipment suppliers

§   Industry cost publications including Costmine

§   SilverCrest financial reporting

§   Santa Elena actual costs including salaries, general and administrative costs (G and A), and consumables costs per unit

The capital costs have been entered into a capital cost schedule, which takes into consideration when the expense occurs, as well as lead times on equipment purchases and payment terms for equipment. As of April 30, 2013, approximately 30% of the new process facility had been constructed with a schedule for completion by January 2014.

The operating costs have been entered into a cost model to determine the cost per tonne of rock material handled to the processing plant or to surface, whether waste or mineralized material.  The operating cost is divided into cost per tonne underground mining, cost per tonne processing and cost per tonne General and Administrative.

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21.1   Capital Costs

EBA has collected information on the budgeted capital costs for the Santa Elena Expansion Project from SilverCrest, and used the underground mine design to estimate the capital required to start and sustain the operation.  A breakdown of the estimated capital expenses per year of the project is included in appendix F. Table 21.1 below shows a summary of the breakdown of the capital estimate, including total capital cost estimated for initial, sustaining and costs already spent on the Expansion Project. Total capital required including sustaining, direct and indirect and contingency but excluding sunk costs is estimated at $US 88 M.

Table 21.1: Capital Cost Summary

Capital expense item	Estimated cost in US $
Total capital – Life of Mine including contingency	87,813,477
Initial capital including contingency	48,067,874
Mining equipment purchases	2,343,810
Development – UG	3,480,550
Process plant including EPCM	41,928,472
Electrical and services – UG	72,042
Indirect	243,000
Sustaining capital including contingency	39,745,603
Mining equipment purchases	5,144,587
Development	31,284,348
Electrical and services	2,283,346
Health and safety	143,442
Construction (incl. labour)	322,880
Indirect	567,000
Contingency included in the above items	9,304,765
Budget spent from January 2012 to  April 30, 2013 (sunk costs)	24,263,480
Underground development completed as of April 30, 2013	5,300,000
Processing plant construction and mill equipment purchases completed as of April 30, 2013	18,963,479

21.1.1   Mining Capital Costs

Mining equipment expenses for the project are summarised in table 21.2, and have been estimated based on production requirements, the mining schedule, ventilation requirements, and general operational requirements including provision for ongoing mining related equipment expenses over the life of mine.

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Table 21.2: Equipment Capital Expenditure

Mining Equipment Type	No.	Estimated Cost Including Contingency And Freight
Long hole drills	1	1,107,750
Scoop trams	3	1,854,090
Jackleg/stopers	4	65,000
Jumbo drills (cut and fill stopes) DD321-40 (110kW)	2	2,085,300
Jumbo accessories	2	304,460
Primary ventilation	5	205,023
Auxiliary fans	10	133,245
Anfo loader	1	30,030
Scissor lift 92 HP	1	115,500
Grader – 110 HP	1	209,000
Kubota/s	3	82,500
Crew transport	2	132,000
Service truck	1	86,900
Shotcrete plant	1	20,790
Underground diamond drill	2	221,463
Compressor	1	60,972
Pipe fusing machine	1	38,375
Ongoing mining related capex		736,000
Cement storage silo for backfill	1	91,125
Total		7,579,523

21.1.2   Underground Development

EBA has estimated the required preproduction and sustaining development required for the underground operations.  This is summarised in table 21.3

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Table 21.3: Development Expenditure

Underground development component	Estimated cost including contingency
Ramp Year -1	1,791,930
Ramp Year 2	1,433,590
Ramp Year 3	4,649,910
Ramp Year 5	1,309,620
Waste development year -1	1,493,620
Waste development year 1	1,338,370
Waste development year 2	1,259,020
Waste development year 3	4,522,375
Waste development year 4	6,385,375
Waste development year 5	3,882,938
Waste development year 6	377,813
Waste development year 7	1,045,688
Ventilation shaft/raise pre-production	652,447
Ventilation shaft/raise year 4	626,251
Ventilation drifts year 1	558,095
Ventilation drifts sustaining	1,646,531
Raise bore underground station 536 level	28,124
Raise bore underground station 325 level	28,124
Pit ramp shaft	707,470
Ventilation cross cuts - pit ramp	28,124
Fresh air raises year 1	133,115
Fresh air raises year 3	97,235
Fresh air raises year 4	315,385
Underground shops	258,750
Underground definition drilling - preproduction	195,000
Total underground development	34,764,898

21.1.3   Underground Electrical Installations

Underground electrical installations have been estimated based on the layout of the underground mine and the planned supply voltage of 5000 V as well as the equipment voltages of 440 V.  Table 21.4 shows the estimated underground electrical equipment and services required and the estimated cost

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Table 21.4: Underground Electrical and Service Cost

Component of underground electrical distribution	Estimated cost including contingency
Surface main underground vacuum breaker	28,750
Main 5kV line to adit area	50,000
Surface to underground high voltage line initial	18,188
Surface to underground high voltage line sustaining	46,875
Cable to surface ventilation fans	12,500
Transformer at surface ventilation	27,500
440v underground line/s	97,500
Underground power centers incl. transformer 500kVA	280,600
Miscellaneous electrical equipment and lighting	67,500
Surface cap lamp battery charger	37,984
Underground water line initial	70,200
Underground water line sustaining	585,000
Underground dewater line - initial	32,500
Underground dewater line - sustaining	65,000
Underground dewater pumps centrifugal initial - 60kW	68,611
Underground dewater pumps centrifugal initial - 45kW	137,223
Submersible pumps 40 HP	57,590
Pump valves/ suction pipes etc.	6,861
Underground dewater pumps centrifugal sustaining	196,726
Trash pumps for slurry	92,508
Compressed air line	195,000
Ventilation tubing	81,900
Leaky feeder communication	98,872
Total	2,355,388

21.1.4   Construction Consumables and Labour

EBA has estimated construction work required to establish the underground systems including ventilation infrastructure, maintenance and safety infrastructure.  Table 21.5 shows a breakdown of the estimated costs.

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Table 21.5: Construction Consumables and Labour

Underground infrastructure construction costs	Estimated costs including contingency
Vent shaft collar concrete	3,837
Vent shaft collar labour	6,750
Vent shaft steel	266
general underground electrical labour	8,100
Fan electrical labour	5,400
Underground ventilation doors	41,885
Ventilation walls materials	9,592
Ventilation walls labour	24,300
Refuge chambers	202,500
Underground machine shop	20,250
Underground pump stations concrete	9,592
Underground pump stations labour	9,281
Ladder ways in fresh air raise year 1	29,925
Ladder ways in fresh air raise year 2	19,007
Ladder ways in fresh air raise year 3	26,982
Ladder ways in fresh air raise year 4	44,876
Ladder ways in fresh air raise year 5	19,007
Ladder ways in fresh air raise year 6	19,130
General underground installations labour hrs.	121,500
Total	622,180

21.1.5   Processing Capital Costs

The estimate as confirmed by John Fox, P.Eng for the cost of construction for the processing plant is as shown in table 21.6 below.  This process plant construction is budgeted to be completed by the end of 2013.  The table illustrates that a portion of the process plant budget is already spent.  The budget includes the equipment purchases, tailings facility, engineering procurement and construction management, as well as power generation.

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Table 21.6: Processing Capital Expenditure

	Opening Balance – Dec 31, 2012	2013 Bod Budget – USD$	Estimated Total – Dec 31, 2013	Paid To Date (Sunk Costs) - In USD$	Paid % Apr 30, 13	Estimate To Complete	Total Projected
Crushing & Screening	409,235	7,420,795	7,830,031	2,189,006	28%	5,641,025	7,830,031
Storage & Handling	-	4,060,339	4,060,339	471,881	12%	3,588,459	4,060,339
Grinding & Classification	630,360	8,470,280	9,100,639	1,340,650	15%	7,759,989	9,100,639
Leaching and CCD	444,606	7,012,128	7,456,735	2,717,322	36%	4,739,413	7,456,735
Recovery (MC Plant)	524,523	5,979,879	6,504,402	2,267,527	35%	4,236,875	6,504,402
Reagents	-	2,499,638	2,499,638	101,571	4%	2,398,067	2,499,638
Water storage	-	347,384	347,384	8,009	2%	339,375	347,384
Tailings	144,431	5,444,546	5,588,977	1,235,010	22%	4,353,967	5,588,977
Auxiliaries	119,000	1,395,271	1,514,271	621,636	41%	892,635	1,514,271
EPCM Group	1,909,259	4,154,638	6,063,897	4,352,395	72%	1,711,502	6,063,897
Gent Sets	-	6,191,048	6,191,048	2,276,429	37%	3,914,619	6,191,048
Stackers	790,869		1,020,869	1,018,479	100%	-	1,018,479
Other Equipment	363,565		363,565	363,565	100%	-	363,565
Total mill and process plant	5,335,849	53,205,946	58,541,795	18,963,479	32%	39,575,926	58,539,404

21.1.6   Indirect Capital Costs

EBA has considered a cost of $900,000 for indirect capital costs this includes:

1.   $200,000 for staff recruiting and expatriation costs

2.   $400,000 for underground design and engineering

21.1.7    Sunk Costs

SilverCrest has advised that the capital spent up to the April 30 2013 includes $5,3 million on the underground development, which has reached the 625 level, and $18.9 million on the processing facility as shown in table 21.6.  These sunk costs have not been included in the economic analysis.

21.2   Operating Costs

The estimation of operating costs has been undertaken by dividing the operation into four areas for costing, these are:

1.   Mining of the underground ore, estimated by EBA

2.   Mining of the reprocessing ore on the heap leach facility , estimated by SilverCrest

3.   Processing the ore estimated by John Fox of Laurion Consulting

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4.   General and Administrative (G and A) costs, estimated by EBA

EBA has summarised the mining costs in Table 21.7 below:

Table 21.7: Operating Costs

Aspects of operations	Cost per tonne ore
Average underground mining cost including ore development	$25 / tonne ore
Rehandling cost of heap leach ore	$0.62 / tonne ore
Processing of fresh ore	$20.50 / tonne ore
Processing of old heap leach ore	$18.50 / tonne ore
Handling of ore to tailings storage facility	$1 / tonne ore
G and A	$ 3.75 / tonne ore processed
G and A per year – based on Santa Elena operating costs for 2013	$3,650,000 /year
Average operating cost per ounce gold equivalent sold*	$569 / oz.
Average operating cost per ounce silver equivalent sold	$11 / oz.

* Gold equivalent ounces calculated at gold to silver post recovery equivalency ratio of 1:52

** Silver equivalent ounces calculated at gold to silver post recovery equivalency ratio of 1:52

21.2.1   Underground Mining Cost Estimate

Assumptions for Underground Mining Costs Estimation

The interim cost estimates for the Santa Elena underground are based on several assumptions. These assumptions include:

1.   Any part of the ore body that has a dip shallower than 50° cannot be mined by long hole stoping, and will be extracted using mechanized cut and fill.

2.   All underground development has been costed at $2,300 / m, as per recent quote by contractor.

3.   Diesel cost is $0.80 / litre of diesel or $3.02 / gal.

4.   Ammonium nitrate explosives is costed at $0.96 / kg of explosive, with emulsion or water gel cost at $3.42 per kg.

5.   Underground haulage of ore has been assumed to be carried out by contractor at a rate of $45/hour for a 20 tonne truck.

6.   The rock engineering work has only be completed at the level required for PFS, further rock engineering work may result in changes to the costs.

Notes

1.   All units are metric

2.   Currency is in US dollars

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3.   Price used for cut-off grade is US $1,450 for gold and $28 for silver, based on projected prices and recovery

4.   Cut-off grades are presented as gold equivalent grades using a gold to silver ratio of 70:1 EBA has undertaken costing for two mining methods, namely long hole stoping and mechanized cut and fill.  The long hole stoping costs are divided into longitudinal long hole stoping and transverse long hole stoping (see section 16 for details of the mining methods).  The costing has further been divided by stoping width, this has been done to provide for the fact that some costs are fixed per length of stope, whereas others are estimable per tonne regardless of the width of the stope.

Summary of mining operating costs for long hole stoping is shown in Table 21.8.

Table 21.8: Long Hole Mining Cost Summary

Summary of estimated costs and cut-off grade calculation - Long Hole
	Long hole stopes < 15m			Long hole stopes > 15m		Units
Mining width	5	10	15	20	25	m
Salary staff*	$1.35	$1.01	$0.68	$0.68	$0.68
Production labour	$3.35	$2.51	$1.68	$1.68	$1.68	$/tonne
Drilling	$1.06	$1.06	$1.06	$1.06	$1.06	$/tonne
Blasting	$0.54	$0.54	$0.54	$0.54	$0.54	$/tonne
Loading	$2.02	$2.02	$2.02	$2.02	$2.02	$/tonne
Service vehicle costs	$1.20	$1.20	$1.20	$1.20	$1.20	$/tonne
Power costs	$5.55	$5.55	$5.55	$5.55	$5.55	$/tonne
Underground haulage ore**	$1.76	$1.76	$1.76	$1.76	$1.76	$/tonne
Long hole backfill / transverse stopes only***				$6.10	$6.10	$/tonne
Cable bolting****	$2.24	$6.76				$/tonne
Bolting of slashed drifts to access hanging wall *****		$0.06				$/tonne
Maintenance labour	$1.67	$1.26	$0.84	$0.84	$0.84	$/tonne
Exploration	$0.80	$0.80	$0.80	$0.80	$0.80	$/tonne
Underground stope development per tonne	$14.78	$7.39	$4.93	$10.10	$9.05	$/tonne
Underground mining cost diluted ore per tonne (including stope development)	$36.33	$31.92	$21.04	$32.32	$31.27	$/tonne
Cut-off grade gold equivalent ******	1.39	1.29	1.04	1.30	1.28	g/tonne

* This includes salary staff over and above that already included in Santa Elena open pit operations which is costed under G and A

** Underground haulage by contractor at an estimated contract rate of $45/hour for trucking which is averaged over the life of mine

*** Backfill applicable only to transverse stopes to enable extraction of primary and secondary stopes (see section 16 for details)

**** Cable bolting is only considered for narrow ore and mining width areas, to reduce dilution, this cost estimate includes an estimate of the cost of slashing the drift to access the hanging wall to place cable bolts

***** Additional bolting required for drifts slashed to access hanging wall for cable bolting

****** Cut of grade includes consideration of processing cost and G and A costs

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Summary of mining operating costs for mechanised cut and fill stoping is shown in table 21.9

Table 21.9: Cut and Fill Cost Summary

Summary of estimated costs and cut-off grade calculation - Cut and Fill
	Stope widths					Units
Mining width	5	10	15	20	25	m
Salary staff on mining*	$1.35	$1.01	$0.68	$0.68	$0.68
Production labour	$3.00	$2.25	$1.50	$1.50	$1.50	$/tonne
Drilling	$2.82	$2.82	$2.82	$2.82	$2.82	$/tonne
Blasting	$1.42	$1.42	$1.42	$1.42	$1.42	$/tonne
Loading	$2.14	$2.14	$2.14	$2.14	$2.14	$/tonne
Service vehicle costs	$1.78	$1.78	$1.78	$1.78	$1.78	$/tonne
Power costs	$10.88	$10.88	$10.88	$10.88	$10.88	$/tonne
Underground haulage ore **	$2.07	$2.07	$2.07	$2.07	$2.07	$/tonne
Backfilling of stopes ***	$7.56	$7.56	$7.56	$7.56	$7.56	$/tonne
Bolting of slashed drifts to access hanging wall	$0.11	$0.11	$0.11	$0.11	$0.11	$/tonne
Maintenance labour	$1.67	$1.26	$0.84	$0.84	$0.84	$/tonne
Exploration	$0.80	$0.80	$0.80	$0.80	$0.80	$/tonne
Underground stope development per tonne	$13.54	$7.65	$5.33	$4.09	$3.32	$/tonne
Underground mining cost	$47.80	$40.74	$37.26	$36.02	$35.24	$/tonne
Cut-off grade gold equivalent****	1.66	1.50	1.42	1.39	1.37	g/tonne

* This includes salary staff over and above that already included in Santa Elena open pit operations which is costed under G and A

** Underground haulage by contractor at an estimated contract rate of $45/hour for trucking which is averaged over the life of mine

*** Backfilling is considered to be done with waste rock and dry stack tailings, with up to 5% cement, hydraulic backfill may also be considered involving pulping underground

**** Cut of grade includes consideration of processing cost and G and A costs

21.2.2   Costs for Loading and Transporting Existing Heap Leach Material to the Mill for Reprocessing

EBA has estimated the cost of loading and transporting the material on the existing heap leach pad to the mill for reprocessing.  This cost is based on using a loader (by contractor, including 25% markup) to load the material onto grasshoppers (portable jump conveyors) which will deliver the material to the reprocess stockpile.  Tables 21.10 to 21.12 show the heap leach reprocessing cost and assumptions.

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Table 21.10: Reprocess Material Rehandling Cost

Loader Cost
Fuel Consumption Rate	0.04	gallons/HP/hour
CAT 950K	211	HP
Fuel Consumption per hour	8.44	gallons/hour
	32.07	Litres/hour
Fuel Cost/Litre	$0.78	$/L
Cost of Fuel per hour	$25.02	$/hour
Labour Cost per hour	$8.00	$/hour
Overhaul		*Costs taken from 2010 CostMine Estimators Guide for 3.8 m3 Wheeled Loader. Converted to 2013 costs
Parts	$1.64
Labour	$1.51
Maintenance
Parts	$3.05
Labour	$2.81
Lube	$2.14
Tires	$11.36
Wear Parts	$0.39
Loader Maintenance Cost/hour	$22.89	$/hour
Total Cost per hour	$55.91	$/hour total
Total Cost per Day	$782.73	$/day
Final Cost per tonne	$0.39	$/tonne
Contractor Cost per tonne	$0.49	$/tonne

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Table 21.11: Grasshopper Conveyor Costs

Grasshopper Conveyor Operating Costs
Consumables/Grasshopper
Replacement Part	Cost/Unit	Replacements/Year	Cost
Motor	$2300/motor	0.5	$1,152
Belt	$31/meter	25	$773
Idlers	$210/Idler	12	$2,520
Bearings	$164/Bearing	4	$656
		Total Consumables Cost	$5,101
Electrical Cost
	7.5 HP conversion to kW	5.63	kW
	Power Draw kWhr	47.25	kWhr
	Power Cost @ $0.22/kWhr	$10.40	Cost/Day
		Total Power Cost	$3,794
		Total Cost/Grasshopper	$8,895
		Cost for 10 Grasshopper	$88,949
		Cost/tonne	$0.12

Table 21.12: Material Handling Cost from Heap Leach to Mill

Aspect	Value	Units
Front end loader to grasshopper	$0.49	$/tonne
Grasshopper operating costs	$0.12	$/tonne
Total Heap Leach to Processing Cost/tonne	$0.61	$/tonne

21.2.3   Processing Plant Operating Costs

The proposed plant for Santa Elena has been designed to process 1million Tonnes/annum of Santa Elena ore. Power will be generated on site from diesel fuel.  Reagent consumption is based on test work for the major consumables (Lime, Cyanide, etc.), but general experience for minor reagents (Zinc, DE etc.).  Steel consumption in comminution is based on power consumption and other relationships established by FC Bond.  Manpower levels and maintenance supplies are based on experience in similar operations. The breakdown of the estimates for processing cost is shown in Table 21.13.

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Table 21.13: Summary of Processing Costs

	US$/year	US$/Tonne	Notes
Manpower	2,097,227	2.10	Revised by Santa Elena
Reagents	4,329,600	4.33	Includes freight allowance
Steel (Comminution)	1,568,684	1.57	Includes freight allowance
Power	8,580,000	8.58	Based on 9.8 Peso/L Diesel  (22c/kWh)
Water	400,000	0.40	New item
Tailings management	1,000,000	1.00	Moving dry tailings by dozer
Maintenance Supplies	1,500,000	1.50	Based on % of Capital Cost  (revised)
TOTAL	19,475,511	19.48

The figures in Table 21.13 are averaged for the whole resource (mined ore and reclaimed heap leach).  Mined ore has to be crushed and thus this material would cost perhaps $1.00/t above average, and the heap leach material doesn’t require crushing and would be about $1.00/t below average ($20.50 for Underground and $18.50 for spent heap leach material)

21.3   Tailings Handling Costs

EBA has estimated the cost of handling of tailings by from end loader, overland conveyor and a bull dozer onto the waste rock dump at $1 / tonne ore processed.

21.4   General and Administrative Operating Costs

In order to assess the general and administrative costs associated with the cost of production, EBA has reviewed the 2012 and 2013 financial statements produced by SilverCrest.  It is the opinion of EBA that the current and historic costs are a reflection of the expected costs for the expanded operation.  EBA has also noted that the production rates achieved during the current operations will be comparable with the initial phases of the Santa Elena underground operation, during a time when the heap leach reprocessing is being undertaken, subsequent to that production rate overall will decrease marginally.  EBA has determined that the current open pit scenario and operational cost framework for G and A is roughly $3,650,000 per year.  EBA expects that G and A to remain relatively similar motivated by the following points:

1.   The current open pit is operated by contractor mining, and therefore much of the skilled staff required for mining are included in the G and A cost and not the mining cost

2.   The production rate for the future underground mining, including the heap leach is similar

3.   Santa Elena currently has staff related expenses under G and A who work on the underground operation, who will continue working on the underground operation

4.   The key changes from the current operational scenario applicable to the Expansion are the construction of a plant and the underground mine.  Both aspects of the Expansion include costs estimates relative to current operations, therefore additional costs are included in the mining and /or processing cost.

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Therefore G and A costs for the Expansion have been estimated at $3.65 Million per year, averaging an estimated $3.60 per tonne over the life of mine.

22.0   ECONOMIC ANALYSIS

22.1   Introduction

Tetra Tech prepared an economic evaluation of the Santa Elena Expansion Project based on a pre-tax financial model and post-tax financial results.  The analysis is based on Q2 2013 US dollars.  No gearing or adjustment for inflation/currency gap is assumed. The complete economic evaluation is provided in Appendix F. This economic analysis includes current Sandstorm agreement terms (see section 4.3).  Note that Sandstorm’s participation in the underground production is pending. This economic analysis does not include sunk costs for all expenses applicable to the Santa Elena expansion up to the 30th of April 2013.

22.2   Technical Assumptions

Technical-Economic assumptions used in the analysis are summarized in Table 22.1.

Currency exchange rates are based on three-year trailing averages.

Refinery costs, as described in Section19.1, reflect actual contract terms and include provision for freight and marketing costs.

Income tax is calculated at a rate of 30% on operating profit less depreciation, applicable loss-carry forwards and employee profit share payments.  Employee profit share is included in the G&A and is structured according to current SilverCrest operations.

Table 22.1: Technical-Economic Modeling Assumptions

Parameter	Value
Markets:
Gold Price1	US$1,450/troy-oz
Silver Price1	US$28.00/troy-oz
Currency Exchange CDN:US$2	1.00:1.00
Currency Exchange US$:MEX$2	1.00:13
Refinery Terms:
Gold	Pay for 99.98%, US$ x/payable oz
Silver	Pay For 99.85%, US$ y/payable oz
Financing:
Analysis Basis	Q2 2013 US dollars
Gearing	None
Income Tax Rate	30%
Depreciation	10-year, Straight Line

i.   EBA’s long-term consensus metal prices as of April 30, 2013.

ii.   Based on three year trailing average (as of April 30, 2013) and rounded –off to two decimals

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22.3   Summary of Financial Results

The pre-tax and post tax economic analysis results for the base case are shown in table 22.2.

Table 22.2: Base Case Economic Analysis Results

Aspects of financial analysis	Units	Value	Per Au oz eq produced	Per Ag oz eq produced
Production
Gold Sold post refiner	k.oz	262
Silver Sold post refiner	k.oz	12,101
Gold Eq ounces	k.oz	496
Silver Eq ounces	k.oz	25,680
Revenue
Gold Sales	US$’000	$346,110
Silver Sales	US$’000	$338,821
Gross Sales	US$’000	$684,931
Operating Expenses
Underground mining	US$’000	$99,443	200.5	3.87
Open pit mining	US$’000	$2,166	4.4	0.08
Heap leach mining	US$’000	$1,764	3.6	0.07
Processing	US$’000	$151,470	305.4	5.90
G&A	US$’000	$27,381	55.2	1.07
Total Operating Costs*	US$’000	$282,223	569	11.0
Freight & Refining	US$’000	$5,579	11.2	0.22
Pre-tax financial results
Operating Margin	US$’000	$397,130	800.7	15.46
Capital Costs	US$’000	-$87,813	-177	-3.42
Mine closure costs	US$’000	-$5,000	-10.1	-0.19
Working Capital	US$’000	-$1,836	-3.7	-0.07
Pre-tax cash flow	US$’000	$302,481	609.8	11.8
Pre-tax NPV 5%	US$’000	$223,670	450.9	8.7
Pre-tax IRR**	%	88%
Post tax financial analysis results
Taxes	US$’000	$100,839	203	3.93
Cash Flow post tax	US$’000	$201,642	406.5	7.85
Post tax NPV 5%	US$’000	$145,556	293.5	5.7
Post tax IRR**	%	60%

* Operating cost per AU EQ oz. sold varies between $496 and $639 over the life of mine

** The IRR presented here do not reflect total project economics but reflect incremental project economics as they do not include sunk costs for the Santa Elena expansion prior to April 30 2013

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The economic analysis does not include 2013 open pit heap leach production, which has an estimated NPV of $20 million based on projected production cash flows at a gold price of $1,250 and a silver price of $19.50.  The economics also excludes the fact that the heap leach reserves will be increased by the open pit production in 2013, and this material will be available for milling.  The heap leach pad reserves will be restated in 2014, first quarter to include the material placed on the pad for 2013.

The post tax financial model was established on a 100% equity basis, excluding debt financing and loan interest charges.  The post tax financial results of the base case are compared to roughly the spot price in the first week of July 2013 in Table 22.3 below.

Table 22.3: Comparison of Base Case Post Tax Results used in the PFS with Spot Price Post Tax Results at mid July 2013

	Item	Base Case	Spot Price
	Gold Price (US$/oz)	$1,450	$1,250
	Silver price (US$/oz	$28	$19.50
Pre tax	Pretax IRR (%)**	88%	49%
	DCF NPV @ 5.0% in millions	$223.7	$108.7
	Payback (production years)	1.1	1.7
Post tax	IRR (%)**	60%	32%
	DCF NPV @ 5.0% in millions	$145.6	$65.1
	DCF NPV @ 8.0%  in millions	$120.4	$50.8
	Payback (production years)	1.5	3.1

* IRR numbers shown here do not reflect full project economics as they do not consider sunk costs

22.4   Cash flows

Figure 22.1 shows the pretax and post tax cash flows.  The peak cash flow occurs in sixth year of operation when higher underground production is achieved and higher grades are being mined in the schedule.

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Figure 22.1: Pre-tax and Post-tax Cash Flows

22.5   Post - tax Results and Sensitivity Analysis

The post tax model has been developed for the PFS by Tetra Tech.  The following assumptions have been used in development of the after tax model:

1.   A tax rate of 30%.

2.   Any employee profit share is included in the G & A and is structured according to current SilverCrest operations.

3.   Depreciation has been deducted from taxable earning as straight-line depreciation over 10 years.

4.   No losses have been carried forward.

5.   Other deductions applicable to pre-tax financial results are applicable to post tax financial results.

The base case as shown Table 22.3 above is used as the base case for the sensitivity analysis.  The sensitivities are determined by initially determining which key variable the project is most sensitive to and then selecting those for further analysis.  In the case of the Santa Elena Expansion Project, these are in order of most sensitive to least sensitive; gold and silver price, operating and capital cost.  The key variables are adjusted from -30% to +30%.  The results for NPV and are IRR shown in Figures 22.2 and 22.3. Tables 22.4 shows the cash flow analysis used to derive the post-tax sensitivities.

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To evaluate the financial performance of the Project, the production schedule was incorporated into the 100% equity pre-tax financial model to develop annual recovered metal production from the relationships of tonnage processed, head grades, and recoveries.

All costs and revenues were assumed to occur at the end of each year in which they were scheduled to occur.

Gold and silver payable values were calculated based on base case metal price and exchange rate.  Unit operating costs for mining, processing and G&A were applied to monthly mined or processed  tonnages, to determine the overall operating cost, which was deducted from the revenues to derive monthly operating cash flow.  Monthly operating cash flows were totaled into annual cash flows, and the financial analysis was run on annual cash flows.

Initial and sustaining capital costs were incorporated on a year-by-year basis over the LOM.  Reclamation bond, working capital and capital equipment salvation value have been included in the cash flow. Capital expenditures were then deducted from the operating cash flow to determine the net cash flow before taxes.

Initial capital expenditures presented in this study include costs accumulated prior to first production of doré for the new process facility but excludes cost for exploration, open pit, and heap leach. Sustaining capital includes expenditures for mining and processing additions, equipment replacement, and process plant expansion.

The undiscounted annual net cash flow (NCF) and cumulative net cash flow (CNCF) are shown in Figure 22.1

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Table 22.4: Discounted Post Tax Cash Flow Model

Discount cash flow model
	Total	Year -1	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
AU ounces sold	262,739		36,173	39,606	29,361	27,557	27,924	53,357	37,535	11,226	0
AG Ounces sold	12,118,926		1,345,248	1,594,643	1,409,639	1,742,741	1,556,867	1,951,279	1,816,118	702,392	0
AU Refinery deductions (oz)	(525)		(72)	(79)	(59)	(55)	(56)	(107)	(75)	(22)	0
AG Refinery deductions (oz)	(18,178)		(2,018)	(2,392)	(2,114)	(2,614)	(2,335)	(2,927)	(2,724)	(1,054)	0
Payable Metal Au(oz) @ Spot	231,214		28,881	31,622	23,442	17,488	27,868	53,250	37,460	11,204	0
Payable Metal Au(oz) Sandstorm	31,000		7,220	7,905	5,861	10,014	0	0	0	0	0
Payable Metal Ag(oz)	12,100,747		1,343,231	1,592,251	1,407,525	1,740,126	1,554,531	1,948,352	1,813,393	701,339	0
Operating revenue	$684,931,142		$82,014,419	$93,201,266	$75,453,077	$77,586,387	$83,935,732	$131,766,045	$105,091,482	$35,882,733	$0
Dore transport and selling costs	($5,578,528)		($625,767)	($739,498)	($649,770)	($797,224)	($714,497)	($893,445)	($836,931)	($321,395)	$0
Operating costs	($282,222,851)		($31,815,916)	($35,254,445)	($35,232,343)	($35,874,143)	($37,792,957)	($45,464,850)	($44,930,788)	($15,857,410)	$0
Operating Profit	$397,129,763		$49,572,736	$57,207,324	$39,570,964	$40,915,021	$45,428,278	$85,407,749	$59,323,763	$19,703,928	$0
Capital	($87,813,477)	($48,067,874)	($6,272,455)	($3,605,930)	($11,554,820)	($10,982,336)	($5,379,848)	($553,384)	($1,221,259)	($175,571)	$0
Working capital	($1,835,534)		($1,835,534)	$0	$0	$0	$0	$0	$0	$0	$0
Closure costs	($5,000,000)										($5,000,000)
					��
Income Tax Payable	($100,838,858)		($13,241,611)	($15,423,809)	($9,786,257)	($9,860,004)	($11,052,586)	($23,029,825)	($15,167,992)	($3,276,774)	$0
Post-tax cash flow	$201,641,895	($48,067,874)	$28,223,136	$38,177,584	$18,229,888	$20,072,681	$28,995,845	$61,824,540	$42,934,512	$16,251,583	($5,000,000)

* Appendix F shows the financial analysis including tax calculations and capital cost schedules

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Figure 22.2: NPV Sensitivity Analysis

Figure 22.3: RR Sensitivity Analysis

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It should be noted that the sensitivity of the Santa Elena Expansion Project to gold and silver price volatility is nearly equal. It can be seen that the NPV and IRR return similar results when either one of the commodity prices is changed.

22.6   Purchase Agreement

On May 14, 2009, SilverCrest entered into a definitive Purchase Agreement with Sandstorm Gold Ltd. under which Nusantara agreed to sell 20% of future gold production from the Santa Elena open pit to Sandstorm, up to 50,000 ounces of gold, in exchange for an Upfront Deposit of $USD 12,000,000. The agreement also provides for ongoing per-ounce payments by Sandstorm equal to the lesser of $350 and the prevailing spot gold market price upon delivery of gold. The per ounce price of $USD 350 is subject to an increase of 1% per annum commencing on the 3rd anniversary of the date that Santa Elena Project began commercial production (i.e. July 2014).

22.7   Smelter Terms

Santa Elena is an operating mine and currently sells gold and silver as part of ongoing operations.  Based on information provided by SilverCrest, EBA has added a cost to account for smelter and refinery costs.

23.0   ADJACENT PROPERTIES

There are no issues related to adjacent properties.

24.0   OTHER RELEVANT DATA AND INFORMATION

The following information has been considered useful to understand the Santa Elena expansion Project.

24.1   Project Execution Plan – Progress Completed

It should be noted that progress towards executing the Santa Elena Expansion Project has already been achieved.  In particular the main ramp, from the 780 level adit, has been advanced to the 625 level (at an elevation of 625 m a.s.l.).  At this elevation a crosscut has been developed into the ore body.  At this elevation inflows of ground water at over 200 gallon per minute, were experienced.  This water will be managed by SilverCrest and may be available to the mine for the process plant.

Progress has also been made in purchasing equipment for the construction of the process plant.  Table 24.1 shows the progress on equipment purchases and construction of the process plant.

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Table 24.1: Process Plant Equipment Purchase Progress

Process plant (Mill) budget as of 30th April 2013
	Opening Balance	2013 Budget	Estimated Total	Paid to Date	Paid %
	Dec 31, 2012	USD$	Dec 31, 2013	USD$	Apr 30, 2013
MILL	US$	US$	US$	US$	US$
Crushing & Screening	409,235	7,077,357	7,486,593	2,189,006	28%
Storage & Handling	-	3,154,074	3,154,074	471,881	12%
Grinding & Classification	630,360	7,206,496	7,836,856	1,340,650	15%
Leaching and CCD	444,606	7,906,602	8,351,209	2,717,322	36%
Recovery (MC Plant)	524,523	5,838,728	6,363,251	2,267,527	35%
Reagents	-	2,109,104	2,109,104	101,571	4%
Water storage	-	337,083	337,083	8,009	2%
Tailings	144,431	5,349,124	5,493,555	1,235,010	22%
Auxiliaries	119,000	1,890,270	2,009,271	621,636	41%
Engineering procurement and construction management	1,683,882	4,153,794	5,837,675	4,352,395	72%
Power generation (diesel generators)	-	7,953,312	7,953,312	2,276,429	37%
Stackers	790,869	230,000	1,020,869	1,018,479	100%
Other Equipment	363,565		363,565	363,565	100%
Total	5,110,472	53,205,945	58,316,417	18,963,479	32%

24.2   Schedule to the Start of Underground Stoping

Based on the progress to date, there is no concern for slippage in the schedule for the Santa Elena Expansion Project. Also, the availability of spent ore can mitigate this risk. In addition, schedule slippage of 2 to 3 months will not cause a significant loss to the project as in that case the mined ore would be heap leached until completion of the process mill.  Though recoveries would be lower in the heap leach than in the mill, the heap leached ore is planned to be reprocessed and thereby resulting in similar overall recoveries. The critical date for the underground mining to commence is after exhaustion of the open pit reserves.  Figure 24.1 shows the schedule for the project as a Gantt chart with the milestone of commencement with stoping as the objective.

Figure 24.1 shows Project schedule up to commencement with stoping (two stopes available)

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Figure 24.1: Santa Elena Expansion Project Development Schedule

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25.0   INTERPRETATION AND CONCLUSION

25.1   Geology and Underground Resource Estimate

The Santa Elena vein is a low sulphidation quartz-calcite epithermal vein striking in  E-NE and dipping between 45-75° to the south. The principal trend of mineralization has a rake of approximately 25° to the southeast and has been delineated to a depth of approximately 475 metres below surface.  Gold and silver grades are constrained within the main quartz vein with some additional mineralization noted in stringers and quartz breccia along the hanging wall and footwall contacts of the vein.  The metal grades are zoned with silver proportions increasing with depth.

The deposit is strongly oxidized at surface and down to depths of at least 150 metres below surface. However, weak to moderate alteration associated with disseminated fine grained pyrite is noted where northwest trending structures crosscut the main vein.  The oxide zone transitions at depth to sulphide bearing, noted by trace amounts of visible argentite noted in the drill core.

25.2   Underground Reserves and Mining Methods

Underground probable reserves of 3.9 million tonnes at an average grade of 3 g/t Au equivalent have potential for economic extraction at current metal prices.  A combination of longitudinal and transverse long hole stoping and mechanised cut and fill methods will be suitable for underground mining.

25.3   Costs and Project Economics

EBA has estimated that approximately $48 million in capital investment is required to commence with production as part of the expansion plan with approximately another $40 million required over the life of mine, this including a contingency of $ 9.3 million over the life of mine.  Operating costs are estimated to be approximately $569 per ounce of gold equivalent and $11 per ounce silver equivalent.  At the base case metal prices the estimated post tax net present value estimate of the operation is roughly $ 145 million at a discount rate of 5%. At the current metal prices the estimated post tax net present value estimate of the operation is roughly $ 65 million at a discount rate of 5%.

26.0   RECOMMENDATIONS

26.1   Geology and Underground Resource Estimate

26.1.1   Risks and Opportunities

Mineralization within the southeasterly plunging trend has been well delineated by the recent 2012-2013 drilling campaign; however, more drilling is required in preparation for underground mining.  Additional drilling on the property should aim to test and confirm the following items:

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§   Delineation drilling along the eastern strike extent of the main vein between the 600 and the 400 metre level elevation to delineate trends with elevated grades and increase resource classification for blocks in this area,

§   Infill drilling on 12 metre centres within the current Indicated resource to upgrade block model confidence for underground mining,

§   accurately delineate the distribution of sulphide bearing mineralization and the transition from near surface oxide to un-oxidized material at depth as this change in mineralogy may have a significant impact of metal recoveries,

§   Dedicated drilling within the Cholugo and Cholugo Dos veins to upgrade current resource and as exploration to depth,

§   Map and delineate the mineralized NW trending structure which cross-cut the main Santa Elena vein as these may support additional resources,

§   Extensively map the Tinaja fault in drilling and from underground development to characterize the vein width and to project the fault into the proposed mine development for planning purposes, and

§   Conduct detailed drill core and underground mapping to assist in determining control and orientation of high grade mineralization trends and ore shoots.

26.2   Geotechnical Recommendation

Geotechnical recommendations are split into two sections for this Pre-Feasibility report:

§   Key recommendations to improve stope size and refine pillar dimensions.

§   Geotechnical considerations in overall mine design.

26.2.1   Stope Size and Pillar Dimensioning

26.2.1.1   Hanging wall  and orebody NGI-Q values

§   Due to the apparent degradation of rock mass quality with depth in the 2012 drilling program it is recommended that further drilling be conducted underground. This degradation could be due to limited availability of geotechnical data at present as footwall rock quality appears fairly consistent at depth. This apparent degradation could create overly conservative stope design parameters and it is recommended that rockmass quality at depth requires further clarification through a drilling program.

26.2.1.2   Joint set orientation

Joint set orientation can play a key role in stope stability. The orientation data at Santa Elena is limited with six orientated holes gleaning a limited number of measurements to gain joint set data for stope layout.

§   Further joint set measurements can be taken through consistent geotechnical mapping of the ore zone and hanging wall during mining activities.

§   Further orientated boreholes can be undertaken during subsurface and surface exploration. During the 2012 surface drilling program a limited number of joint orientation data was able to be retrieved. It is strongly recommended that the use of a Televiewer be incorporated into further data collection.  This allows for in-situ joint orientation data to be collected that will aid in stope definition.

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26.2.1.3   In-situ stress regime

The in-situ stress regime at Santa Elena needs clarification as the stress regime has been assumed at pre-feasibility level.

§   It is strongly advised that in-situ stress analysis be conducted at Santa Elena in order to determine principal stress magnitude and trajectory and horizontal to vertical stress ratio. This should be done in order to maximize ore extraction, minimize pillar dimensions, assist in pillar recovery design and minimize support cost to local rather than regional zones based on the in-situ stress regime. EBA can assist in developing a cost effective stress measurement program in conjunction with Silver Crest for this purpose.

26.2.1.4   Hanging wall  and ore body strength and deformation characteristics

Clarification of hanging wall and ore body strength characteristics is recommended beyond pre-feasibility level as there is concern that the UCS strength results that have been returned so far are relatively low for the general rock types encountered.

§   In addition to simple UCS strengths for the footwall, hanging wall  and orebody the Hoek-Brown parameters and deformation characteristics need to be determined for the rock conditions found at Santa Elena in order to refine stope and pillar design to a higher degree of accuracy beyond pre-feasibility level.

§   It is strongly recommended that tri-axial testing be carried out on hanging wall, footwall and ore samples together with further UCS testing with strain measurements in order to determine site specific strength and deformation characteristics of the rock types at Santa Elena for inclusion into future geotechnical modelling of the mine and for stope and pillar refinement. It is recommended that 80% of future testing be carried out at Sonora University, Mexico and 20% at UBC, Vancouver as a reference check.

26.2.2   Geotechnical Considerations in Overall Mine Design

§   As the orebody varies in dip and dip direction and thickness, both along strike and down dip, it is strongly advised that three dimensional geotechnical analysis be conducted on the orebody once the stress regime and rock deformation characteristics have been determined. This can be advantageous to mine sequencing, backfill addition and pillar recovery in order to further maximise ore extraction.

§   It is recommended to space rib pillars across levels so as to avoid stacking of pillars. This will minimize the creation and effect of tall slender pillars on the overall mine wide excavation.

§   This PFS does not consider the interaction between the pit and the proposed underground stopes. This should be reviewed once stoping outlines have been defined for feasibility or design stages.

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26.3   Recommendations on Mining Methods and Mining Costs

EBA recommend that SilverCrest further investigate the purchase of commercial underground mining trucks as opposed to use of an underground haulage contractor.  EBA suggests that the reliability and productivity of commercial underground mining trucks, may improve the efficiency of haulage of ore and waste rock to surface.

EBA recommend that SilverCrest investigate hydraulic and/or paste fill as a means of disposing of mill tailings, and the use of boreholes drilled from surface to deliver the tailings underground, as opposed to backhaul by trucks.

26.4   Environmental, Closure and Asset Retirement

A previous independent closure and reclamation plan was created in 2010 for the original open pit mining and heap leaching operations.  Numerous elements of the project and impacts to the property have been modified from the original project scope for the current Expansion Project.  It is recommended that SilverCrest update the estimated funding, or bond, requirements that will be necessary for site closure and long term monitoring activities.

27.0   REFERENCES

Aguirre-Díaz, G., and McDowell, F., 1991, The volcanic section at Nazas, Durango, Mexico, and the possibility of widespread Eocene volcanism within the Sierra Madre Occidental: Journal of Geophysical Research, v. 96, p. 13,373–13,388.

Aguirre-Díaz, G., and McDowell, F., 1993, Nature and timing of faulting and synextensional magmatism in the southern Basin and Range, central-eastern Durango, Mexico: Geological Society of America Bulletin, v. 105, p. 1435–1444.

Alaniz-Alvarez and Nieto-Samaniego, A.F., 2007, the Taxco-San Miguel de Allende fault system and the Trans-Mexican Volcanic Belt: Two tectonic boundaries in central Mexico active during the Cenozoic, in Alaniz-Alvarez, S.A and Nieto-Samaniego, A.F., ed., Geology of Mexico: Celebrating the Centenary of the Geological Society of Mexico: Geological society of America special Paper 422, p. 301-316.

Anderson, T. A., and Schmidt,V. A., 1983, The evolution of Middle America and the Gulf of Mexico-Caribbean Sea during Mesozoic time: Geological Society of America Bulletin, v. 9, p. 941–966.

Barr, P.J.F, Chow, J., Allard, G. and Wallis., C.S. 2011.NI 43-101 Technical Report Reserve Update for the Santa Elena Open Pit and Preliminary Assessment for the Santa Elena and Cruz de Mayo Expansion Project Sonora, Mexico. Prepared for SilverCrest Mines Ltd, Effective Date; April 1st, 2011.

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. Global Resource Engineering (2011)

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.

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Campa, M. F., and P., Coney, 1983, Tectonostratigraphic terranes and mineral resources distribution in Mexico: Canadian Journal of Earth Sciences, v. 20, p. 1040–1051.

Carter, 1990) or Carter, T.G. (1992), .A new approach to Surface Crown Pillar Design,. Proc. 16th Can. Rock Mechanics Symposium, Sudbury, pp. 75-83.

Clark, L.M. and Pakalnis, R.C. (1997). An Empirical Design Approach for Estimating Unplanned Dilution from Open Stope Hanging walls and Footwalls. Proceedings of the 99th AGM - CIM. Vancouver, pp. 25.

Mining Cost Service. 2012. Costmine Infomine USA, Inc.

Index Mundi. >http://www.indexmundi.com/<. Accessed May 20th 2013

Nieto-Samaniego, A.F., Alaniz-Alvarez, S.A., and Camprubi, A. (2007), Mesa Central of Mexico: Stratigraphy, structure, and Cenozoic tectonic evolution, p. 41-70; in Geology of Mexico: Celebrating the Centenary of the Geological Society of Mexico, The Geological Society of America, Special Paper 422, 2007, edited by Susana A. Alaniz-Alvarez and Angel F. Nieto-Samaniego; 465pp.

Ferrari, L. Valencia-Moreo, M., Bryan, S., (2007), Magmatism and tectonics of the Sierra Madre Occidental and its relation with the evolution of the western margin of north America, p. 1-29; in Geology of Mexico: Celebrating the Centenary of the Geological Society of Mexico, The Geological Society of America, Special Paper 422, 2007, edited by Susana A. Alaniz-Alvarez and Angel F. Nieto-Samaniego; 465pp.

Fier, E.N, and Wallis, C.S., 2007, Technical Report On The Cruz De Mayo Property, Sonora Mexico, Prepared For SilverCrest Mines Inc. December 10, 2007.

Fier, E. N., 2009, Technical Report On The Santa Elena Property, Sonora Mexico, Prepared For Silvercrest Mines Inc., February 15, 2009.

Johnson, C. M., 1991, Large-scale crust formation and lithosphere modification beneath middle to late Cenozoic calderas and volcanic fields, Western North-America: Journal of Geophysical Research, v. 96, p. 13485–13507.

Lang, B., 1994, Span Design for entry type excavations, MASC Thesis, University of British Columbia, Vancouver, BC.

Mathews, K.E. et al., 1981, Prediction of stable excavation spans for mining at depths below 1000m in hard rock, CAMMET) Report DSS Serial No. OSQ80-00081.

Prosser, B.S and Wallace K.G 1999. Practical Values of Friction Factors. Mine Ventilation Services, Inc. (MVS)

Potvin, Y., M. Hudyma, H.D.S. Miller, 1988, Design guidelines for open stope support, CIM Bulletin, vol. 82, No. 926, June, pp. 53-62.Nickson (1992)

Rogers, John J.W et.al 2004. Continents and Supercontinents. Chapter 6. p85.

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Wark, D. A., Kempter, K. A., and McDowell, F. W., 1990, Evolution of waning subduction-related magmatism, northern Sierra Madre Occidental, Mexico: Geological Society of America Bulletin, v. 102, p. 1555–1564.

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APPENDIX A

QUALIFIED PERSON CERTIFICATES

Certificate of qualified person – Nathan Eric Fier

I, Nathan Eric Fier, do hereby declare that:

1.   I currently reside in Mission, British Columbia, Canada, and am currently the President and Chief Operating Officer for SilverCrest Mines Inc. My office address is Suite 501, 570 Granville Street, Vancouver, B.C., V6C 3P1.

2.   I hold two Bachelors of Science degrees in Geological Engineering (1984) and Mining Engineering (1986) from Montana Tech, Butte, Montana.

3.   I am a member in good standing in the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC), member # 135165, (APEGSK), member # 23381 and a Certified Professional Geologist registered with the American Institute of Professional Geologists (AIPG, USA) member # 10622.

4.   I am an author and Qualified Person (within the meaning of National Instrument 43-101) responsible for the preparation of the Technical Report entitled:

“SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY AND OPEN PIT RESERVE UPDATE” SONORA, MEXICO

NI 43-101 TECHNICAL REPORT

PREPARED FOR SILVERCREST MINE S INC.

July 25, 2013

Effective Date: April 30, 2013

5.   I am responsible for sections 1, 4.2 to 4.4, 10.3, 11.5, 12.6, 13.2, 14.5 to 15.6 excluding 15.4, 16.6.2 to 16.10 excluding 16.7, 18.1, 18.4, 19, 21.1.6, 21.1.7, 21.2.2, 21.3, 21.4, 23, 24, and 26.4.

6.   As a Qualified Person for this report, I have read the National Instrument 43-101 and Companion Policy and confirm that the sections of this report for which I am responsible have been prepared in compliance to National Instrument 43-101.

7.   I have visited the Santa Elena property on numerous occasions between January 2005 to July 2013.

8.   I have worked on numerous related properties over 28 years of worldwide geologic and mining experience including but not limited to;

a.   Various Technical Reports on Santa Elena Property, Mexico – low sulfidation epithermal deposit (LSED).

b.   Technical Report on Cruz de Mayo Property, Mexico – LSED.

c.   Technical Report on San Marcial Property, Mexico – LSED.

d.   Due Diligence on Gold Road Property, Arizona – LSED.

e.   Technical Report on La Joya Property, Mexico – Skarn.

f.   Development manager for the La Colorada and La Trinidad mines, Mexico – LSED.

g.   Technical Report on El Zapote Property, El Salvador – Breccia Pipe.

h.   Technical Report on El Ocote Property, Honduras – Breccia Pipe.

i.   Work on several Carlin-type gold deposits – Gold Quarry, Betze-Post, Leeville.

j.   Work on several Copper-gold-moly porphyry deposits – Batu Hijau, Endako.

9.   I am not independent of SilverCrest Mines as independence is described in Section 1.5 of the National Instrument 43-101. In addition I am currently a shareholder of SilverCrest and am I directly entitled to financially benefit from its success.

10.   Prior to this report, I have had extensive prior involvement on the Santa Elena property including multiple previous Technical Reports, exploration, development, construction and production.

11.   To the best of my knowledge, information and belief, as of the effective date of the Technical Report, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Certificate of qualified person - P. James F. Barr

I, P. James F. Barr, do hereby declare that:

1)   I currently reside in Vancouver, British Columbia, Canada, and am currently employed as Senior Geologist by EBA, Engineering Consultants Ltd., with office address at 9th floor, 1066 W Hastings Street, Vancouver, British Columbia.

2)   I hold a Bachelors of Science with Honours from the University of Waterloo (2003), Ontario, Canada, with a major in Environmental Science, Earth Science and Chemistry and I have practiced as an exploration and resource geologist in Canada and Mexico since 2003.

3)   I am a member in good standing in the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC), member 35150.

4)   I am a co-author and Qualified Person responsible for the preparation of the Technical Report  entitled:

“SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY AND OPEN PIT RESERVE UPDATE”

SONORA, MEXICO

NI 43-101 TECHNICAL REPORT

PREPARED FOR SILVERCREST MINES INC

July 25, 2013

Effective Date: April 30th, 2013

5)   I am responsible for sections 2-10, 11.1-11.4, 11.6, 12.1-12.5, 14.1-14.4, 20, 25.1 and 26.1 of this Technical Report.

6)   As a Qualified Person for this report, I have read the National Instrument 43-101 and Companion Policy and confirm that this report has been prepared in compliance to National Instrument 43-101.

7)   I have visited the Santa Elena property on 4 separate occasions since November 2010, the most recent of which occurred from Oct 13-14, 2012,

8)   I have worked on and visited numerous epithermal, skarn and geologically related properties in this and other regions of Mexico, and have been conducting Mineral Resource Estimates for more than 5 years.

9)   I am independent of SilverCrest Mines Inc. as independence is described in Section 1.5 of the National Instrument 43-101.  In addition, I am currently not a shareholder of SilverCrest nor am I directly entitled to financially benefit from its success.

10)   Prior to this report, I was a co-author of the May 11, 2011, Technical Report titled “NI 43-101 Technical Report, Reserve Update for the Santa Elena Open Pit and Preliminary Assessment for the Santa Elena and Cruz de Mayo Expansion Project, Sonora, Mexico”, with Effective Date April 1st, 2011.

11)   To the best of my knowledge, information and belief, as of the Effective Date of the report, the parts of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not mis-leading.

Certificate of qualified person – C. Michael Tansey

I, C. Michael Tansey, do hereby declare that:

1) I currently reside in St. Albert, Alberta, Canada, and am currently employed as a senior mining engineer by EBA Engineering Consultants Ltd. My address is 32 Windsor Crescent, St. Albert, Alberta.

2) I hold a Bachelors of Science in Geotechnical Sciences from the University of Montreal (Loyola College), graduating in 1970.

3) I am a member in good standing in Northwest Territories and Nunavut Association of Professional Engineers and Geoscientists (NAPEG), member #1368

4) I am an author and Qualified Person (within the meaning of National Instrument 43-101) responsible for the preparation of the Technical Report entitled:

“SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY AND OPEN PIT RESERVE UPDATE”

SONORA, MEXICO

NI 43-101 TECHNICAL REPORT

PREPARED FOR SILVERCREST MINES INC.

July 25, 2013

Effective Date: April 30, 2013

5) I am responsible, for sections  15.4, 16 excluding 16.2, 16.8, 16.9 and 16.10, 18.2, section 21 excluding 21.1.5, 21.1.6, 21.1.7, 21.2.2, 21.2.3 and 21.3, 25.2 and section 26.3.

6) As a Qualified Person for this report, I have read the National Instrument 43-101 and Companion

Policy and confirm that this report has been prepared in compliance to National Instrument 43-101.

7) I have visited the Santa Elena property in August 2011 for 4 days and January 2012 for 4 days.

8) I have practiced my profession continuously from 1971 to 2006, semi-retired from 2007 to present, and have worked in mine engineering positions at underground nickel, copper, zinc and gold operations in Canada, including senior technical and operations supervision; conducted mining project evaluations, project permitting and closure planning for underground gold properties; and conducted mining research and development projects in the potash industry.

9) I am independent of SilverCrest Mines as independence is described in Section 1.5 of the National

Instrument 43-101. In addition I am currently not a shareholder of SilverCrest nor am I directly entitled to financially benefit from its success.

10) Prior to this report, I have had no prior involvement on this property.

11) To the best of my knowledge, information and belief, as of the effective date of the Technical Report, the parts of the Technical Report for which I am responsible contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Certificate of Qualified Person – John Fox, P.Eng

I, John R.W. Fox, do hereby declare that:

1)   I currently reside in Vancouver, British Columbia, Canada, and am currently a consulting metallurgical engineer with Laurion Consulting Inc with offices at 302-304 West Cordova St. Vancouver, B.0 , V6B 1E8

2)   I hold a Bachelors of Science in Applied Mineral Sciences and Extractive Metallurgy from the University of Leed, UK.

3)   I am a member in good standing in the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC), member #12578.

4)   I am an author and Qualified Person (within the meaning of National Instrument 43-101) responsible for part of the preparation of the Technical Report entitled:

“SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY AND OPEN PIT RESERVE UPDATE”

SONORA, MEXICO

NI 43-101 TECHNICAL REPORT

PREPARED FOR SILVERCREST MINES INC.

July 25, 2013

Effective Date: April 30, 2013

5)   I am responsible for sections 13 except 13.2, 17, 1.5, 18.3, 21.1.5 & 21.2.3.

6)   As a Qualified Person for this report, I have read the National Instrument 43-101 and Companion Policy and confirm that this report has been prepared in compliance to National Instrument 43­101.

7)   I have visited the Santa Elena property in June 2013 for one day.

8)   I have been involved in numerous gold and silver properties including, Escalante Silver (Utah USA) Patchway, Brompton & Renco Gold (Zimbabwe) Snip & Bronson Creek (B.C) Ketza River (Yukon) Kremnica (Czeck Republic) Congress (Arizona, USA) La Colorado (Mexico) Equis (Colombia) La Victoria(Venezuela) Pangkut and Way Linggo (Indonesia) and many more.

9)   I am independent of SilverCrest Mines as independence is described in Section 1.5 of the National Instrument 43-101. In addition I am currently not a shareholder of SilverCrest nor am I directly entitled to financially benefit from its success.

10)   Prior to this report, I have had no prior involvement on this property.

11)   To the best of my knowledge, information and belief, as of the Effective Date of the report, the parts of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not mis-leading.

Certificate of Qualified Person – Carlos Chaparro

I, Carlos Chaparro, do hereby declare that:

1)   I currently reside in Vancouver, British Columbia, Canada, and am currently employed as Senior Geotechnical Engineer by EBA, Engineering Consultants Ltd., with office address at 9th floor, 1066 W Hastings Street, Vancouver, British Columbia.

2)   I hold a Bachelors of Science in Civil Engineering from Universidad Javeriana in Bogota, Colombia and a Masters Degree in Civil Geotechnical Engineering from University of Illinois. I have practiced as geotechnical engineer in Colombia, USA, Canada and Mexico since 1995.

3)   I am a member in good standing in the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC), member 148633.

4)   I am a co-author and Qualified Person responsible for the preparation of the Technical Report  entitled:

“SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY AND OPEN PIT RESERVE UPDATE”

SONORA, MEXICO

NI 43-101 TECHNICAL REPORT

PREPARED FOR SILVERCREST MINES INC.

July 25, 2013

Effective Date: April 30, 2013

5)   I am responsible for sections 16.2 and 26.2 of this Technical Report.

6)   As a Qualified Person for this report, I have read the National Instrument 43-101 and Companion Policy and confirm that this report has been prepared in compliance to National Instrument 43-101.

7)   I have visited the Santa Elena property on the 7th, 8th, 11th and 12th of July, 2012.

8)   I have worked on and visited several underground and numerous open pit mines: underground projects in Central BC, Yukon, Nunavut and Mexico; open pit projects in Northwest Territories, Yukon, Nunavut, Ontario, Alaska, Greenland, Mexico and West Africa.

9)   I am independent of SilverCrest Mines Inc. as independence is described in Section 1.5 of the National Instrument 43-101.  In addition, I am currently not a shareholder of SilverCrest nor am I directly entitled to financially benefit from its success.

10)   Prior to this report, I have no previous involvement with the property.

11)   To the best of my knowledge, information and belief, as of the Effective Date of the report, the parts of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not mis-leading.

Certificate of qualified person – Nick Michael

I, Nick Michael, do hereby declare that:

I am a Principal Mineral Economist with Tetra Tech, Inc. with a business address at 350 Indiana Street, Suite 500, Golden, Colorado 80401, USA.

I am a graduate of the Colorado School of Mines in Golden, Colorado USA in mining engineering (1983) and received and received an MBA from Willamette University (1986).  I have practiced my profession continuously since 1987.  Since 1990, I have completed valuations, evaluations (technical-economic models), and have audited a variety of projects including exploration, pre-production (feasibility-level), operating and mine closure projects.  I have also served as expert witness with respect to technical-economic issues.

I have read the definition of “qualified person” set out in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“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 a Qualified Person (within the meaning of National Instrument 43-101) responsible for the preparation of the Technical Report entitled:

“SANTA ELENA EXPANSION PRE-FEASIBILITY STUDY AND OPEN PIT RESERVE UPDATE”

SONORA, MEXICO

NI 43-101 TECHNICAL REPORT

PREPARED FOR SILVERCREST MINES INC.

July 25, 2013

Effective Date: April 30, 2013

I am responsible for sections 22 and 25.3.

As a Qualified Person for this report, I have read the National Instrument 43-101 and Companion Policy and confirm that this report has been prepared in compliance to National Instrument 43-101.

I have never visited the Santa Elena property.

I am independent of SilverCrest Mines as independence is described in Section 1.5 of the National Instrument 43-101.  In addition I am currently not a shareholder of SilverCrest nor am I directly entitled to financially benefit from its success.

Prior to this report, I have had no prior involvement on this property.

To the best of my knowledge, information and belief, as of the effective date of the Technical Report, the part of the Technical Report for which I am responsible contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL 30, 2013 | RELEASED DATE: JULY 25, 2013

APPENDIX B

CERTIFIED REFERENCE MATERIAL DATA SHEETS

CDN Resource Laboratories Ltd.

#2, 20148 – 102nd Ave, Langley, B.C., Canada, V1M 4B4, 604-882-8422, Fax: 604-882-8466 (www.cdnlabs.com)

REFERENCE MATERIAL: CDN-CM-17

Recommended values and the “Between Lab” Two Standard Deviations

Gold	1.37 g/t	±	0.13 g/t	Certified value	30g FA /ICP or AA
Copper	0.791 %	±	0.040 %	Certified value	4-acid /ICP or AA
Copper	0.798 %	±	0.034 %	Certified value	Aqua regia /ICP or AA
Molybdenum	0.075 %	±	0.008 %	Certified value	4-acid / ICP or AA
Molybdenum	0.073 %	±	0.014 %	* *Provisional * *	Aqua regia /ICP or AA
Silver	14.4 g/t	±	1.4 g/t	Certified value	4-acid /ICP or AA
Silver	14.9 g/t	±	1.4 g/t	Certified value	Aqua regia /ICP or AA

Note:   Standards with an RSD of near or less than 5% are certified; RSD’s of between 5% and 15% are Provisional; RSD’s over 15% are Indicated. Provisional and Indicated values cannot be used to monitor accuracy with a high degree of certainty.

PREPARED BY: CERTIFIED BY: DATE OF CERTIFICATION: INDEPENDENT GEOCHEMIST: CDN Resource Laboratories Ltd. Duncan Sanderson, B.Sc., Licensed Assayer of British Columbia Dr. Barry Smee., Ph.D., P. Geo. December 8, 2011

METHOD OF PREPARATION:

Reject ore material was dried, crushed, pulverized and then passed through a 270 mesh screen. The +270 material was discarded. The -270 material was mixed for 5 days in a double-cone mixer. Splits were taken and sent to 15 laboratories for round robin assaying.

ORIGIN OF REFERENCE MATERIAL:

Standard CDN-CM-17 was prepared using 770 kg of a granitic rock blended with 29 kg of a Cu-Au-Mo concentrate and 5 kg of a high grade Ag ore.

Approximate chemical composition (from whole rock analysis) is as follows:

	Percent		Percent
SiO2	68.0	MgO	2.1
Al2O3	11.8	K2O	1.1
Fe2O3	6.8	TiO2	0.6
CaO	3.3	LOI	2.0
Na2O	2.8	S	1.0
C	0.1

Statistical Procedures:

The final limits were calculated after first determining if all data was compatible within a spread normally expected for similar analytical methods done by reputable laboratories. Data from any one laboratory was removed from further calculations when the mean of all analyses from that laboratory failed a t test of the global means of the other laboratories. The means and standard deviations were calculated using all remaining data. Any analysis that fell outside of the mean ±2 standard deviations was removed from the ensuing data base. The mean and standard deviations were again calculated using the remaining data. This method is different from that used by Government agencies in that the actual “between-laboratory” standard deviation is used in the calculations. This produces upper and lower limits that reflect actual individual analyses rather than a grouped set of analyses. The limits can therefore be used to monitor accuracy from individual analyses, unlike the Confidence Limits published on other standards.

REFERENCE MATERIAL: CDN-CM-17

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13	Lab 14	Lab 15
	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t
CM-17-1	1.35	1.29	1.20	1.43	1.38	1.36	1.36	1.32	1.43	1.40	1.33	1.41	1.30	1.41	1.45
CM-17-2	1.40	1.23	1.29	1.39	1.28	1.35	1.41	1.34	1.42	1.39	1.33	1.42	1.37	1.38	1.33
CM-17-3	1.35	1.17	1.21	1.42	1.34	1.26	1.33	1.26	1.45	1.49	1.32	1.42	1.35	1.42	1.31
CM-17-4	1.43	1.45	1.26	1.41	1.33	1.29	1.44	1.23	1.35	1.47	1.32	1.44	1.34	1.42	1.30
CM-17-5	1.42	1.39	1.32	1.36	1.41	1.31	1.32	1.31	1.42	1.40	1.30	1.45	1.43	1.40	1.34
CM-17-6	1.48	1.25	1.41	1.42	1.32	1.24	1.44	1.27	1.53	1.36	1.36	1.45	1.29	1.42	1.37
CM-17-7	1.45	1.34	1.40	1.37	1.42	1.28	1.40	1.28	1.40	1.36	1.33	1.44	1.55	1.40	1.34
CM-17-8	1.33	1.29	1.56	1.39	1.26	1.33	1.35	1.28	1.45	1.46	1.30	1.41	1.45	1.42	1.29
CM-17-9	1.37	1.47	1.24	1.43	1.37	1.27	1.42	1.32	1.48	1.49	1.34	1.45	1.29	1.40	1.39
CM-17-10	1.44	1.34	1.22	1.45	1.31	1.38	1.45	1.31	1.42	1.34	1.30	1.45	1.38	1.40	1.32
Mean	1.40	1.32	1.31	1.41	1.34	1.30	1.39	1.29	1.44	1.42	1.32	1.44	1.38	1.41	1.35
Std. Devn.	0.0498	0.0958	0.1160	0.0292	0.0526	0.0470	0.0483	0.0336	0.0479	0.0568	0.0195	0.0175	0.0825	0.0123	0.0473
% RSD	3.55	7.27	8.84	2.08	3.92	3.61	3.47	2.60	3.34	4.01	1.47	1.22	6.00	0.88	3.51
Total	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t
CM-17-1	14	13.1	14.6	14.4	14.0	15.5	15	15	14.2	13.5	14.3	15.0		13.8	16
CM-17-2	14	14.1	14.6	14.3	14.2	14.9	14	15	14.9	13.4	15.7	14.9		13.9	17
CM-17-3	15	14.1	14.1	14.0	14.0	15.5	14	15	14.7	13.9	14.4	15.3	16.6	13.7	16
CM-17-4	14	13.3	14.4	14.1	13.8	15.0	15	13	15.0	14.2	15.1	15.2	14.7	13.9	16
CM-17-5	15	12.4	14.6	13.9	13.7	14.5	14	13	14.2	14.2	13.4	15.2	16.2	13.8	17
CM-17-6	14	12.7	14.7	14.7	14.2	15.1	14	13	14.7	13.7	14.2	15.2	15.8	14.3	16
CM-17-7	14	12.9	14.6	14.1	14.4	15.2	14	15	15.0	13.3	15.3	15.1	15.1	14.4	15
CM-17-8	14	13.5	14.3	14.7	14.0	14.3	13	14	14.6	14.7	13.7	15.0	16.1	13.7	15
CM-17-9	14	12.9	14.7	13.8	14.1	15.8	15	15	14.8	13.7	14.6	14.8	15.2	14.5	13
CM-17-10	15	12.8	14.0	14.2	13.8	14.8	14	14	14.5	13.9	13.4	15.0	17.4	14.3	14
Mean	14.3	13.2	14.5	14.2	14.0	15.1	14.2	14.2	14.7	13.9	14.4	15.1	15.9	14.0	15.5
Std. Devn.	0.4830	0.5731	0.2503	0.3084	0.2150	0.4648	0.6325	0.9189	0.2914	0.4275	0.7866	0.1567	0.8854	0.3173	1.2693
% RSD	3.38	4.35	1.73	2.17	1.53	3.09	4.45	6.47	1.99	3.09	5.46	1.04	5.57	2.26	8.19
Aqua Regia	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t
CM-17-1	15	14.8	15.0	14.0	13.9	16.2	13	15	14.9	15.6	16	14.8		14.5	16.0
CM-17-2	16	14.9	14.7	14.1	14.3	15.9	15	15	15.6	14.9	17	14.9		13.7	15.8
CM-17-3	15	14.6	14.4	14.9	13.9	15.6	15	15	15.1	14.7	16	14.8	14.0	14.0	14.8
CM-17-4	15	14.8	15.0	15.6	14.2	15.4	14	15	15.2	14.4	16	15.1	14.2	13.9	14.6
CM-17-5	16	14.5	14.5	13.3	13.7	15.3	14	15	14.8	13.7	17	14.7	14.3	13.9	17.2
CM-17-6	15	15.9	14.1	14.4	13.8	15.2	14	15	15.1	15.6	16	15.1	13.4	13.1	16.0
CM-17-7	15	14.8	14.6	15.0	14.5	15.9	14	15	14.8	13.7	17	15.0	13.7	14.0	14.6
CM-17-8	16	14.2	14.5	15.1	14.3	15.7	14	15	14.8	15.7	16	15.0	15.0	13.7	15.8
CM-17-9	16	14.6	14.9	14.3	14.6	16.5	13	15	14.9	15.5	16	14.8	13.6	14.2	14.6
CM-17-10	16	16.0	14.3	14.7	14.0	15.9	13	15	14.9	15.2	16	14.8	13.5	13.1	15.8
Mean	15.5	14.9	14.6	14.5	14.1	15.8	13.9	15.0	15.0	14.9	16.3	14.9	14.0	13.8	15.5
Std. Devn.	0.5270	0.5840	0.3018	0.6586	0.3048	0.4061	0.7379	0.0000	0.2514	0.7630	0.4830	0.1414	0.5165	0.4413	0.8548
% RSD	3.40	3.92	2.07	4.53	2.16	2.58	5.31	0.00	1.68	5.12	2.96	0.95	3.70	3.20	5.51

REFERENCE MATERIAL: CDN-CM-17

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13	Lab 14	Lab 15
Total	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu
CM-17-1	0.766	0.759	0.81	0.789	0.80	0.812	0.780	0.767	0.802	0.756	0.809	0.83		0.785	0.777
CM-17-2	0.780	0.762	0.79	0.791	0.81	0.801	0.788	0.790	0.804	0.754	0.798	0.82		0.786	0.792
CM-17-3	0.780	0.760	0.80	0.795	0.80	0.806	0.766	0.796	0.789	0.755	0.777	0.83	0.859	0.788	0.777
CM-17-4	0.785	0.755	0.83	0.786	0.81	0.805	0.785	0.791	0.801	0.768	0.798	0.81	0.827	0.785	0.789
CM-17-5	0.790	0.763	0.82	0.790	0.82	0.795	0.783	0.769	0.792	0.750	0.804	0.80	0.848	0.783	0.808
CM-17-6	0.785	0.767	0.83	0.796	0.82	0.800	0.786	0.788	0.792	0.744	0.780	0.83	0.854	0.778	0.807
CM-17-7	0.775	0.751	0.78	0.795	0.80	0.796	0.768	0.757	0.798	0.765	0.812	0.81	0.810	0.785	0.803
CM-17-8	0.778	0.779	0.83	0.793	0.81	0.793	0.775	0.770	0.810	0.766	0.775	0.82	0.816	0.788	0.798
CM-17-9	0.803	0.753	0.80	0.794	0.81	0.808	0.762	0.781	0.802	0.759	0.800	0.82	0.791	0.793	0.789
CM-17-10	0.795	0.748	0.80	0.796	0.82	0.809	0.762	0.767	0.807	0.749	0.777	0.82	0.820	0.786	0.786
Mean	0.784	0.760	0.809	0.793	0.810	0.803	0.776	0.778	0.800	0.757	0.793	0.819	0.828	0.786	0.793
Std. Devn.	0.0105	0.0090	0.0179	0.0034	0.0082	0.0065	0.0102	0.0132	0.0069	0.0080	0.0143	0.0099	0.0235	0.0039	0.0113
% RSD	1.34	1.18	2.22	0.43	1.01	0.81	1.32	1.70	0.86	1.06	1.80	1.21	2.84	0.49	1.42
Aqua Regia	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu
CM-17-1	0.816	0.790	0.78	0.820	0.80	0.792	0.745	0.802	0.819	0.836	0.792	0.82		0.654	0.817
CM-17-2	0.810	0.788	0.81	0.773	0.81	0.801	0.740	0.792	0.813	0.809	0.789	0.81		0.626	0.797
CM-17-3	0.791	0.799	0.74	0.787	0.80	0.785	0.775	0.811	0.813	0.820	0.755	0.82	0.799	0.643	0.778
CM-17-4	0.810	0.795	0.74	0.777	0.81	0.813	0.753	0.791	0.818	0.814	0.798	0.81	0.790	0.666	0.780
CM-17-5	0.808	0.792	0.76	0.773	0.80	0.815	0.745	0.789	0.816	0.828	0.778	0.80	0.803	0.670	0.745
CM-17-6	0.775	0.800	0.75	0.774	0.81	0.804	0.750	0.791	0.815	0.814	0.781	0.81	0.798	0.630	0.762
CM-17-7	0.799	0.787	0.78	0.789	0.81	0.821	0.750	0.757	0.810	0.826	0.781	0.81	0.793	0.620	0.758
CM-17-8	0.805	0.782	0.80	0.785	0.81	0.814	0.750	0.801	0.816	0.842	0.789	0.81	0.803	0.635	0.791
CM-17-9	0.795	0.783	0.79	0.773	0.81	0.803	0.754	0.785	0.814	0.814	0.810	0.80	0.799	0.652	0.772
CM-17-10	0.811	0.799	0.80	0.792	0.81	0.826	0.759	0.781	0.815	0.819	0.800	0.80	0.792	0.662	0.786
Mean	0.802	0.792	0.775	0.784	0.807	0.807	0.752	0.790	0.815	0.822	0.787	0.809	0.797	0.646	0.779
Std. Devn.	0.0123	0.0066	0.0259	0.0146	0.0048	0.0128	0.0096	0.0146	0.0026	0.0107	0.0150	0.0074	0.0051	0.0175	0.0208
% RSD	1.53	0.84	3.35	1.86	0.60	1.59	1.28	1.84	0.32	1.30	1.91	0.91	0.63	2.71	2.67
Total	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo
CM-17-1	0.073	0.075	0.07	0.072	0.082	0.074	0.077	0.080	0.074	0.074	0.079	0.07		0.075	0.074
CM-17-2	0.074	0.075	0.07	0.071	0.080	0.073	0.078	0.079	0.073	0.074	0.079	0.07		0.075	0.076
CM-17-3	0.075	0.076	0.07	0.073	0.084	0.072	0.076	0.079	0.074	0.075	0.078	0.07	0.085	0.076	0.074
CM-17-4	0.074	0.074	0.07	0.072	0.082	0.071	0.076	0.077	0.073	0.076	0.080	0.07	0.081	0.076	0.076
CM-17-5	0.074	0.075	0.07	0.072	0.083	0.071	0.074	0.079	0.074	0.075	0.078	0.07	0.084	0.077	0.083
CM-17-6	0.074	0.077	0.07	0.070	0.082	0.071	0.079	0.079	0.074	0.075	0.075	0.07	0.084	0.075	0.076
CM-17-7	0.076	0.073	0.07	0.071	0.084	0.071	0.076	0.081	0.073	0.076	0.077	0.07	0.081	0.076	0.074
CM-17-8	0.076	0.075	0.07	0.072	0.083	0.072	0.077	0.084	0.073	0.076	0.077	0.07	0.081	0.076	0.075
CM-17-9	0.076	0.072	0.07	0.072	0.083	0.073	0.076	0.081	0.074	0.074	0.081	0.07	0.078	0.076	0.076
CM-17-10	0.075	0.071	0.07	0.072	0.082	0.073	0.075	0.080	0.075	0.075	0.078	0.07	0.081	0.076	0.074
Mean	0.075	0.074	0.070	0.072	0.083	0.072	0.076	0.080	0.074	0.075	0.078	0.070	0.082	0.076	0.076
Std. Devn.	0.0011	0.0017	0.0000	0.0008	0.0012	0.0011	0.0014	0.0019	0.0007	0.0009	0.0017	0.0000	0.0023	0.0004	0.0026
% RSD	1.42	2.30	0.00	1.15	1.43	1.53	1.87	2.32	0.92	1.17	2.16	0.00	2.83	0.59	3.42
Aqua Regia	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo	% Mo
CM-17-1	0.076	0.054	0.07	0.072	0.082	0.063	0.078	0.077	0.066	0.085	0.075	0.07		0.050	0.078
CM-17-2	0.075	0.051	0.07	0.071	0.083	0.054	0.080	0.076	0.068	0.084	0.074	0.07		0.046	0.075
CM-17-3	0.075	0.054	0.07	0.071	0.083	0.056	0.079	0.078	0.068	0.082	0.069	0.07	0.074	0.046	0.074
CM-17-4	0.075	0.052	0.07	0.073	0.082	0.055	0.078	0.078	0.065	0.084	0.074	0.07	0.071	0.049	0.073
CM-17-5	0.076	0.053	0.07	0.071	0.082	0.059	0.080	0.075	0.067	0.084	0.072	0.06	0.073	0.049	0.069
CM-17-6	0.073	0.056	0.07	0.071	0.082	0.057	0.079	0.076	0.068	0.083	0.072	0.06	0.077	0.048	0.070
CM-17-7	0.075	0.051	0.07	0.070	0.083	0.059	0.080	0.075	0.067	0.083	0.073	0.06	0.074	0.050	0.071
CM-17-8	0.074	0.053	0.07	0.071	0.081	0.058	0.078	0.076	0.067	0.086	0.071	0.07	0.076	0.046	0.078
CM-17-9	0.075	0.054	0.07	0.071	0.082	0.063	0.081	0.076	0.069	0.084	0.075	0.07	0.076	0.046	0.074
CM-17-10	0.076	0.054	0.08	0.072	0.081	0.061	0.079	0.076	0.067	0.083	0.072	0.06	0.073	0.048	0.071
Mean	0.075	0.053	0.071	0.071	0.082	0.059	0.079	0.076	0.067	0.084	0.073	0.066	0.074	0.048	0.073
Std. Devn.	0.0009	0.0015	0.0032	0.0008	0.0007	0.0031	0.0010	0.0011	0.0011	0.0010	0.0019	0.0052	0.0017	0.0015	0.0031
% RSD	1.26	2.78	4.45	1.15	0.90	5.36	1.30	1.39	1.69	1.19	2.60	7.82	2.33	3.17	4.17

Note: Both Cu and Mo aqua regia data from Lab 14 was excluded for failing the t test.

REFERENCE MATERIAL: CDN-CM-17

Participating Laboratories:
  (not in same order as listed in table of results)

Acme Analytical Laboratories Ltd., Vancouver, B.C., Canada

Actlabs, Ancaster, Ontario, Canada

Actlabs, Stewart, B.C., Canada

Actlabs, Thunder Bay, Ontario, Canada

ALS Chemex Laboratories, North Vancouver, B.C., Canada

AGAT, Mississauga, Ontario, Canada

AHK, Alaska, USA

Alex Stewart Argentina SA

Stewart Group, Kamloops, B.C., Canada

CIMM, Lima, Peru

Inspectorate, Richmond, B.C., Canada
Omac, Ireland
SGS, Lima, Peru

Skyline Assayers & Laboratories, Arizona, USA
TSL Laboratories, Saskatoon, Canada

Legal Notice:

This certificate and the reference material described in it have been prepared with due care and attention. However CDN Resource Laboratories Ltd. or Barry Smee accept no liability for any decisions or actions taken following the use of the reference material. Our liability is limited solely to the cost of the reference material.

Certified by

 

Duncan Sanderson, Certified Assayer of B.C.

Geochemist

 

Dr. Barry Smee, Ph.D., P. Geo.

CDN Resource Laboratories Ltd.

#2, 20148 – 102nd Ave, Langley, B.C., Canada, V1M 4B4, 604-882-8422, Fax: 604-882-8466 (www.cdnlabs.com)

REFERENCE MATERIAL: CDN-ME-5

Recommended values and the “Between Lab” Two Standard Deviations

Gold	1.07 g/t Au	±	0.14 g/t Au	** provisional value only **
Silver	206.1 g/t Ag	±	13.1 g/t Ag	(FA / Grav)
Silver	205.6 g/t Ag	±	9.3 g/t Ag	(Digestion, ICP)
Copper	0.840 % Cu	±	0.048 % Cu
Lead	2.13 % Pb	±	0.12 % Pb
Zinc	0.579 % Zn	±	0.020 % Zn

** Note:

Standards with an RSD of near or less than 5% are certified, RSD’s of between 5% and 15% are Provisional, and RSD’s over 15% are Indicated. Provisional and Indicated values cannot be used to monitor accuracy with a high degree of certainty.

PREPARED BY: CERTIFIED BY: DATE OF CERTIFICATION: INDEPENDENT GEOCHEMIST: CDN Resource Laboratories Ltd. Duncan Sanderson, B.Sc., Licensed Assayer of British Columbia Dr. Barry Smee., Ph.D., P. Geo. September 3, 2009

METHOD OF PREPARATION:
Reject ore material was dried, crushed, pulverized and then passed through a 270 mesh screen. The +270 material was discarded. The -270 material was mixed for 5 days in a double-cone mixer. Splits were taken and sent to twelve laboratories for round robin assaying.

ORIGIN OF REFERENCE MATERIAL:
This standard is made primarily from ore supplied by US Silver from the Coeur d’Alene mining district in northern Idaho. The mineralization occurs as veins hosted by weakly metamorphosed, siliceous sediments. Ag-Cu ore occurs as tetrahedrite, and variable amounts of pyrite and chalcopyrite. Minor Pb is associated with Ag-Cu veins. Other portions of the mineralized areas include Pb-Ag veins primarily consisting of galena and quartz. The standard was made by mixing 200 kg of US Silver ore with 40 kg of a high grade Au ore, 22 kg of a Cu, Zn concentrate and 535 kg of a blank granitic material.

Approximate chemical composition is as follows:

	Percent		Percent
SiO2	65.7	MgO	1.5
Al2O3	9.2	K2O	1.1
Fe2O3	10.0	TiO2	0.4
CaO	2.2	LOI	4.5
Na2O	2.2	S	1.8

Statistical Procedures:

The final limits were calculated after first determining if all data was compatible within a spread normally expected for similar analytical methods done by reputable laboratories. Data from any one laboratory was removed from further calculations when the mean of all analyses from that laboratory failed a t test of the global means of the other laboratories. The means and standard deviations were calculated using all remaining data. Any analysis that fell outside of the mean ±2 standard deviations was removed from the ensuing data base. The mean and standard deviations were again calculated using the remaining data. This method is different from that used by Government agencies in that the actual “between-laboratory” standard deviation is used in the calculations. This produces upper and lower limits that reflect actual individual analyses rather than a grouped set of analyses. The limits can therefore be used to monitor accuracy from individual analyses, unlike the Confidence Limits published on other standards.

Assay Procedures:
Au: Fire assay pre-concentration, AA or ICP finish (30g sub-sample).

Ag: Fire assay pre-concentration, gravimetric finish (30g sub-sample).
Ag, Cu, Pb, Zn: 4-acid digestion, AA or ICP finish.

REFERENCE MATERIAL: CDN-ME-5

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13
	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t
CDN-ME-5-1	0.82	0.92	1.06	0.99	1.08	1.19	1.10	0.91	1.17	1.10	0.99	1.08	1.02
CDN-ME-5-2	1.32	1.12	1.02	0.96	1.17	1.07	1.20	1.08	0.84	1.26	1.21	1.28	1.01
CDN-ME-5-3	1.28	1.16	1.05	1.02	1.18	1.13	1.17	1.02	1.12	1.02	1.45	0.99	1.03
CDN-ME-5-4	0.91	1.26	1.10	1.11	1.08	1.02	1.18	1.06	1.16	1.19	1.60	1.07	1.01
CDN-ME-5-5	1.40	0.89	1.12	1.03	1.15	1.09	1.27	1.01	1.01	1.02	1.26	1.09	1.03
CDN-ME-5-6	0.91	0.97	1.11	1.16	1.14	1.05	1.26	1.11	1.04	1.00	1.10	1.12	1.01
CDN-ME-5-7	1.24	1.08	1.10	1.11	1.17	1.04	1.13	1.17	1.06	0.91	0.88	0.98	1.03
CDN-ME-5-8	1.08	0.99	1.08	1.07	1.01	1.04	1.19	1.15	1.20	1.04	1.47	0.94	1.02
CDN-ME-5-9	1.01	1.07	1.12	0.95	1.12	1.14	1.10	0.98	1.12	1.01	1.00	0.99	1.03
CDN-ME-5-10	0.99	1.06	1.02	1.13	1.25	1.09	1.10	1.13	1.05	0.97	1.18	0.99	1.01
Mean	1.10	1.05	1.08	1.05	1.14	1.09	1.17	1.06	1.08	1.05	1.21	1.05	1.02
Std. Devn.	0.2004	0.1109	0.0385	0.0738	0.0665	0.0536	0.0622	0.0823	0.1040	0.1044	0.2348	0.0993	0.0094
% RSD	18.28	10.54	3.57	7.01	5.86	4.93	5.33	7.75	9.66	9.93	19.35	9.44	0.92
(FA / Grav.)	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t
CDN-ME-5-1	204	206	206.4	211	205	197	216.2	212	200	210	197
CDN-ME-5-2	207	205	214.6	206	192	197	210.2	186	203	208	199
CDN-ME-5-3	206	207	212.5	210	209	200	212.5	197	204	214	197
CDN-ME-5-4	203	206	210.3	219	207	203	218.2	202	202	217	194
CDN-ME-5-5	210	204	209.7	219	201	200	209.5	198	201	212	194
CDN-ME-5-6	212	203	204.9	224	202	202	215.3	198	201	208	196
CDN-ME-5-7	208	204	205.9	218	205	203	217.1	203	205	212	191
CDN-ME-5-8	201	210	211.3	215	199	200	212.2	204	206	216	200
CDN-ME-5-9	201	205	213.5	213	205	201	215.4	202	202	208	213
CDN-ME-5-10	204	203	210.9	217	206	199	215.8	198	202	212	195
Mean	205.6	205.3	210.0	215.2	203.1	200.2	214.2	200.0	202.6	211.7	197.6
Std. Devn.	3.688	2.111	3.302	5.287	4.886	2.150	2.955	6.616	1.897	3.268	6.004
% RSD	1.79	1.03	1.57	2.46	2.41	1.07	1.38	3.31	0.94	1.54	3.04
Digestion / ICP	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t
CDN-ME-5-1	201	200	206.1	202	211	209	215.0	205	211	201	213	194	212
CDN-ME-5-2	203	204	208.7	202	211	213	209.5	193	211	205	212	202	211
CDN-ME-5-3	203	206	208.2	204	216	215	209.0	200	211	201	217	194	207
CDN-ME-5-4	205	205	205.2	206	210	208	207.6	197	211	203	217	198	205
CDN-ME-5-5	208	209	202.3	204	211	208	210.5	191	213	200	216	200	216
CDN-ME-5-6	202	210	210.2	202	215	210	210.5	196	207	203	219	196	209
CDN-ME-5-7	204	208	210.7	202	208	206	211.1	201	207	198	218	200	210
CDN-ME-5-8	205	205	205.5	206	206	201	208.7	202	205	197	215	200	207
CDN-ME-5-9	207	204	210.1	204	203	211	206.2	197	205	199	211	198	212
CDN-ME-5-10	200	202	207.2	200	209	212	211.6	197	204	201	215	204	210
Mean	203.8	205.3	207.4	203.2	210.0	209.3	210.0	197.9	208.5	200.8	215.3	198.6	209.9
Std. Devn.	2.530	3.093	2.678	1.932	3.859	3.945	2.413	4.202	3.240	2.440	2.627	3.273	3.143
% RSD	1.24	1.51	1.29	0.95	1.84	1.89	1.15	2.12	1.55	1.22	1.22	1.65	1.50

NOTE:  Au data from Labs 1 and 11 was excluded due to poor precision.
Ag data (digestion / ICP) from Lab 11 was excluded for failing the “t” test

REFERENCE MATERIAL: CDN-ME-5

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13
	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu
CDN-ME-5-1	0.846	0.842	0.871	0.816	0.860	0.811	0.858	0.80	0.827	0.859	0.839	0.888	0.809
CDN-ME-5-2	0.847	0.846	0.857	0.825	0.941	0.845	0.872	0.83	0.798	0.872	0.834	0.900	0.815
CDN-ME-5-3	0.847	0.835	0.846	0.821	0.889	0.838	0.853	0.86	0.812	0.870	0.840	0.866	0.802
CDN-ME-5-4	0.858	0.872	0.867	0.826	0.870	0.800	0.839	0.85	0.824	0.867	0.844	0.862	0.813
CDN-ME-5-5	0.843	0.847	0.866	0.820	0.899	0.799	0.854	0.81	0.806	0.870	0.840	0.899	0.815
CDN-ME-5-6	0.848	0.828	0.860	0.833	0.911	0.811	0.853	0.88	0.801	0.873	0.836	0.850	0.801
CDN-ME-5-7	0.852	0.839	0.862	0.836	0.871	0.779	0.839	0.85	0.801	0.872	0.844	0.860	0.810
CDN-ME-5-8	0.857	0.844	0.862	0.830	0.891	0.794	0.849	0.85	0.797	0.874	0.837	0.885	0.802
CDN-ME-5-9	0.871	0.837	0.865	0.827	0.878	0.824	0.870	0.84	0.789	0.866	0.841	0.854	0.805
CDN-ME-5-10	0.857	0.843	0.854	0.823	0.924	0.830	0.845	0.82	0.795	0.870	0.838	0.888	0.809
Mean	0.853	0.843	0.861	0.826	0.893	0.813	0.853	0.839	0.805	0.869	0.839	0.875	0.808
Std. Devn.	0.0083	0.0116	0.0072	0.0060	0.0258	0.0209	0.0113	0.0242	0.0125	0.0044	0.0032	0.0188	0.0052
% RSD	0.98	1.37	0.84	0.73	2.88	2.58	1.32	2.89	1.55	0.51	0.39	2.15	0.64
	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb
CDN-ME-5-1	2.16	2.03	2.13	2.25	2.11	2.11	2.25	2.00	2.10	2.07	2.169	2.17	2.08
CDN-ME-5-2	2.19	2.07	2.12	2.26	2.18	2.16	2.20	2.06	2.06	2.09	2.150	2.17	2.10
CDN-ME-5-3	2.20	2.07	2.12	2.31	2.20	2.22	2.20	2.11	2.06	2.08	2.164	2.17	2.06
CDN-ME-5-4	2.19	2.05	2.18	2.29	2.16	2.03	2.27	2.11	2.10	2.09	2.192	2.14	2.06
CDN-ME-5-5	2.20	2.07	2.17	2.28	2.18	2.17	2.21	2.00	2.06	2.08	2.173	2.18	2.11
CDN-ME-5-6	2.17	2.07	2.17	2.32	2.20	2.03	2.20	2.14	2.03	2.09	2.159	2.16	2.06
CDN-ME-5-7	2.19	2.09	2.15	2.34	2.13	2.11	2.23	2.07	2.02	2.06	2.188	2.14	2.09
CDN-ME-5-8	2.20	2.06	2.13	2.35	2.12	2.07	2.20	2.10	1.99	2.04	2.165	2.22	2.11
CDN-ME-5-9	2.20	2.04	2.16	2.33	2.09	2.21	2.23	2.09	1.98	2.04	2.184	2.20	2.08
CDN-ME-5-10	2.17	2.07	2.10	2.22	2.18	2.17	2.28	2.02	1.99	2.04	2.164	2.22	2.02
Mean	2.19	2.06	2.14	2.30	2.16	2.13	2.23	2.07	2.04	2.07	2.17	2.18	2.08
Std. Devn.	0.0149	0.0175	0.0267	0.0425	0.0395	0.0691	0.0302	0.0492	0.0441	0.0215	0.0134	0.0287	0.0268
% RSD	0.68	0.85	1.25	1.85	1.83	3.25	1.35	2.38	2.16	1.04	0.62	1.32	1.29
	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn
CDN-ME-5-1	0.59	0.583	0.58	0.602	0.574	0.555	0.577	0.54	0.594	0.562	0.585	0.585	0.588
CDN-ME-5-2	0.58	0.584	0.58	0.596	0.590	0.568	0.583	0.56	0.587	0.567	0.582	0.585	0.577
CDN-ME-5-3	0.58	0.582	0.57	0.610	0.595	0.579	0.561	0.57	0.599	0.560	0.588	0.585	0.585
CDN-ME-5-4	0.59	0.607	0.58	0.608	0.581	0.557	0.555	0.57	0.607	0.566	0.584	0.582	0.591
CDN-ME-5-5	0.58	0.596	0.59	0.604	0.580	0.558	0.582	0.55	0.595	0.555	0.597	0.586	0.588
CDN-ME-5-6	0.58	0.578	0.57	0.623	0.589	0.563	0.592	0.58	0.596	0.567	0.581	0.588	0.579
CDN-ME-5-7	0.59	0.587	0.57	0.622	0.570	0.552	0.584	0.57	0.588	0.548	0.589	0.584	0.585
CDN-ME-5-8	0.59	0.583	0.56	0.602	0.564	0.536	0.587	0.58	0.608	0.551	0.585	0.582	0.588
CDN-ME-5-9	0.60	0.583	0.57	0.602	0.565	0.563	0.573	0.57	0.599	0.551	0.583	0.577	0.590
CDN-ME-5-10	0.59	0.588	0.57	0.600	0.578	0.572	0.581	0.56	0.602	0.559	0.582	0.597	0.579
Mean	0.587	0.587	0.574	0.607	0.579	0.560	0.578	0.565	0.598	0.559	0.586	0.585	0.585
Std. Devn.	0.0067	0.0085	0.0084	0.0091	0.0106	0.0119	0.0116	0.0127	0.007	0.0071	0.0048	0.0051	0.005
% RSD	1.15	1.44	1.47	1.50	1.83	2.12	2.00	2.25	1.18	1.27	0.81	0.88	0.85

NOTE: Cu data from Lab 5 was excluded for failing the “t” test.
Pb data from Lab 4 was excluded for failing the “t”test.
Zn data from Lab 4 was excluded for failing the “t”test.

REFERENCE MATERIAL: CDN-ME-5

Participating Laboratories:
   (not in same order as listed in table of results)

 Acme Analytical Laboratories Ltd., Vancouver

 Actlabs-Ancaster, Ontario, Canada

 Actlabs-Thunder Bay, Ontario, Canada

 ALS Chemex Laboratories, North Vancouver

 Alaska Assay Laboratories, Alaska, USA

 Assayers Canada Ltd., Vancouver

 Eco Tech, B.C., Canada

 Genalysis Laboratory, Australia

 Labtium Laboratory, Finland

 SGS Toronto, Ontario, Canada

 Skyline Laboratories, Arizona, USA

 TSL Laboratories Ltd., Saskatoon

 Ultra Trace Analytical Laboratories, Australia

Legal Notice:

This certificate and the reference material described in it have been prepared with due care and attention. However CDN Resource Laboratories Ltd. or Barry Smee accept no liability for any decisions or actions taken following the use of the reference material. Our liability is limited solely to the cost of the reference material.

Certified by

 

Duncan Sanderson, Certified Assayer of B.C.

Geochemist

 

Dr. Barry Smee, Ph.D., P. Geo

CDN Resource Laboratories Ltd.

#2, 20148 – 102nd Ave, Langley, B.C., Canada, V1M 4B4, 604-882-8422, Fax: 604-882-8466 (www.cdnlabs.com)

REFERENCE MATERIAL: CDN-ME-16

   Recommended values and the “Between Lab” Two Standard Deviations

Gold	1.48 g/t	±	0.14 g/t
Silver	30.8 g/t	±	2.2 g/t
Copper	0.671 %	±	0.036 %
Lead	0.879 %	±	0.040 %
Zinc	0.807 %	±	0.040 %

PREPARED BY: CERTIFIED BY: DATE OF CERTIFICATION: INDEPENDENT GEOCHEMIST: CDN Resource Laboratories Ltd. Duncan Sanderson, B.Sc., Licensed Assayer of British Columbia Dr. Barry Smee., Ph.D., P. Geo. September 3, 2009

METHOD OF PREPARATION:
Reject ore material was dried, crushed, pulverized and then passed through a 270 mesh screen. The +270 material was discarded. The -270 material was mixed for 5 days in a double-cone mixer. Splits were taken and sent to 14 laboratories for round robin assaying.

ORIGIN OF REFERENCE MATERIAL:

This standard is made from a mixture of ores.

Approximate chemical composition (from whole rock analysis) is as follows:

	Percent		Percent
SiO2	47.8	MgO	3.9
Al2O3	14.0	K2O	1.2
Fe2O3	14.6	TiO2	0
CaO	9.0	LOI	5.3
Na2O	1.2	S	6.5
C	0.8

Statistical Procedures:

The final limits were calculated after first determining if all data was compatible within a spread normally expected for similar analytical methods done by reputable laboratories. Data from any one laboratory was removed from further calculations when the mean of all analyses from that laboratory failed a t test of the global means of the other laboratories. The means and standard deviations were calculated using all remaining data. Any analysis that fell outside of the mean ±2 standard deviations was removed from the ensuing data base. The mean and standard deviations were again calculated using the remaining data. This method is different from that used by Government agencies in that the actual “between-laboratory” standard deviation is used in the calculations. This produces upper and lower limits that reflect actual individual analyses rather than a grouped set of analyses. The limits can therefore be used to monitor accuracy from individual analyses, unlike the Confidence Limits published on other standards.

Assay Procedures:

Au: Fire assay pre-concentration, AA or ICP finish (30g sub-sample).
Ag, Cu, Pb, Zn: 4-acid digestion, AA or ICP finish.

REFERENCE MATERIAL: CDN-ME-16

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13	Lab 14	Lab 15
	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t
ME-16-1	1.56	1.50	1.51	1.55	1.55	1.53	1.47	1.16	1.27	1.50	1.29	1.46	1.48	1.45	1.48
ME-16-2	1.47	1.42	1.51	1.52	1.50	1.62	1.46	1.46	1.23	1.55	1.52	1.48	1.45	1.43	1.50
ME-16-3	1.51	1.42	1.46	1.56	1.52	1.46	1.43	1.28	1.26	1.55	1.29	1.36	1.55	1.37	1.45
ME-16-4	1.58	1.46	1.50	1.52	1.54	1.69	1.54	1.25	1.18	1.51	1.50	1.34	1.54	1.40	1.59
ME-16-5	1.40	1.46	1.51	1.51	1.48	1.58	1.25	1.32	1.35	1.59	1.46	1.47	1.57	1.45	1.53
ME-16-6	1.43	1.39	1.42	1.45	1.58	1.49	1.47	1.22	1.18	1.35	1.29	1.32	1.51	1.46	1.48
ME-16-7	1.42	1.49	1.51	1.65	1.51	1.51	1.50	1.31	1.10	1.46	1.52	1.37	1.46	1.51	1.52
ME-16-8	1.56	1.42	1.49	1.37	1.45	1.64	1.48	1.18	1.31	1.55	1.35	1.45	1.55	1.49	1.54
ME-16-9	1.56	1.47	1.50	1.58	1.57	1.52	1.69	1.51	1.15	1.68	1.24	1.46	1.55	1.42	1.45
ME-16-10	1.53	1.45	1.48	1.43	1.54	1.52	1.53	1.45	1.29	1.42	1.50	1.38	1.52	1.45	1.48
Mean	1.50	1.45	1.49	1.51	1.52	1.56	1.48	1.31	1.23	1.52	1.40	1.41	1.52	1.44	1.50
Std. Devn.	0.0673	0.0364	0.0292	0.0804	0.0409	0.0725	0.1089	0.1208	0.0790	0.0912	0.1138	0.0606	0.0418	0.0408	0.0437
% RSD	4.48	2.52	1.96	5.31	2.69	4.66	7.34	9.20	6.41	6.02	8.15	4.30	2.76	2.83	2.91
	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t
ME-16-1	32.9	32	30.0	31	30.0	30.4	32	30.9	26.2	30.6	31.2	31	30	29.4	30.8
ME-16-2	33.0	29	31.0	31	31.2	30.0	32	30.6	25.7	30.5	34.7	29	30	30.2	31.1
ME-16-3	32.5	29	29.5	30	31.9	30.3	32	31.2	24.8	30.0	29.1	31	30	29.7	31.6
ME-16-4	32.0	30	31.5	30	30.3	29.9	32	31.6	25.4	30.4	38.7	32	29	31.1	31.3
ME-16-5	32.8	30	30.5	30	31.1	30.0	33	30.0	27.4	29.8	29.2	31	29	30.1	31.1
ME-16-6	32.7	30	32.0	30	30.4	28.8	32	30.0	27.5	30.0	29.3	30	30	28.9	30.8
ME-16-7	32.4	33	30.5	30	30.2	33.0	32	33.2	28.1	30.4	29.5	32	31	31.6	30.8
ME-16-8	32.5	33	30.0	30	31.0	31.4	32	32.4	28.4	32.0	39.0	31	30	29.5	30.5
ME-16-9	33.9	30	30.0	32	30.8	30.0	32	31.6	27.0	31.7	23.3	32	30	30.4	32.0
ME-16-10	32.6	31	30.5	31	30.7	30.2	31	31.9	26.8	30.7	29.7	30	30	29.3	30.4
Mean	32.7	30.7	30.6	30.5	30.8	30.4	32.0	31.3	26.7	30.6	31.4	30.9	29.9	30.0	31.0
Std. Devn.	0.4990	1.4944	0.7619	0.7071	0.5728	1.1106	0.4714	1.0309	1.1842	0.7138	4.8146	0.9944	0.5676	0.8417	0.4926
% RSD	1.52	4.87	2.49	2.32	1.86	3.65	1.47	3.29	4.43	2.33	15.35	3.22	1.90	2.80	1.59

Note: Both Au and Ag results from Laboratory 9 were removed for failing the “t” test.

REFERENCE MATERIAL: CDN-ME-16

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13	Lab 14	Lab 15
	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu
ME-16-1	0.70	0.688	0.676	0.684	0.674	0.68	0.681	0.69	0.687	0.680	0.654	0.656	0.652	0.670	0.645
ME-16-2	0.71	0.680	0.651	0.664	0.675	0.68	0.679	0.69	0.673	0.657	0.661	0.605	0.653	0.664	0.648
ME-16-3	0.68	0.675	0.648	0.666	0.677	0.69	0.698	0.70	0.656	0.655	0.691	0.636	0.662	0.685	0.647
ME-16-4	0.70	0.680	0.678	0.664	0.669	0.70	0.696	0.71	0.649	0.667	0.692	0.631	0.661	0.675	0.640
ME-16-5	0.70	0.684	0.652	0.657	0.675	0.69	0.694	0.68	0.718	0.657	0.666	0.634	0.660	0.671	0.648
ME-16-6	0.68	0.672	0.655	0.665	0.660	0.68	0.699	0.70	0.725	0.660	0.639	0.627	0.652	0.669	0.645
ME-16-7	0.69	0.679	0.657	0.675	0.670	0.69	0.691	0.67	0.711	0.665	0.653	0.655	0.657	0.674	0.647
ME-16-8	0.69	0.683	0.646	0.659	0.672	0.68	0.690	0.69	0.726	0.666	0.653	0.651	0.639	0.679	0.651
ME-16-9	0.69	0.680	0.663	0.662	0.675	0.68	0.692	0.70	0.698	0.679	0.658	0.666	0.643	0.683	0.647
ME-16-10	0.69	0.692	0.667	0.663	0.674	0.68	0.690	0.70	0.684	0.666	0.642	0.632	0.645	0.672	0.651
Mean	0.693	0.681	0.659	0.666	0.672	0.685	0.691	0.693	0.693	0.665	0.661	0.639	0.652	0.674	0.647
Std. Devn.	0.0095	0.0058	0.0113	0.0078	0.0048	0.0071	0.0066	0.0116	0.0277	0.0085	0.0180	0.0178	0.0079	0.0065	0.0032
% RSD	1.37	0.86	1.72	1.17	0.72	1.03	0.96	1.67	4.00	1.28	2.72	2.79	1.22	0.97	0.49
	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb
ME-16-1	0.94	0.861	0.870	0.882	0.884	0.88	0.890	0.88	0.948	0.902	0.898	0.906	0.889	0.87	0.883
ME-16-2	0.93	0.849	0.867	0.852	0.902	0.85	0.880	0.88	0.936	0.906	0.903	0.826	0.897	0.87	0.877
ME-16-3	0.94	0.836	0.879	0.847	0.901	0.87	0.890	0.87	0.912	0.904	0.943	0.858	0.901	0.89	0.889
ME-16-4	0.95	0.853	0.906	0.860	0.883	0.87	0.890	0.84	0.901	0.914	0.962	0.859	0.891	0.86	0.872
ME-16-5	0.95	0.852	0.890	0.846	0.888	0.86	0.890	0.89	1.008	0.893	0.903	0.854	0.895	0.87	0.888
ME-16-6	0.96	0.837	0.892	0.857	0.896	0.85	0.890	0.87	1.011	0.910	0.881	0.845	0.903	0.86	0.880
ME-16-7	0.95	0.852	0.885	0.878	0.890	0.86	0.880	0.91	0.990	0.904	0.876	0.881	0.901	0.87	0.893
ME-16-8	0.95	0.847	0.873	0.858	0.899	0.86	0.880	0.90	1.040	0.907	0.860	0.886	0.879	0.85	0.892
ME-16-9	0.96	0.841	0.867	0.848	0.901	0.86	0.880	0.89	0.973	0.923	0.880	0.862	0.904	0.88	0.888
ME-16-10	0.95	0.860	0.880	0.852	0.901	0.86	0.880	0.90	0.949	0.905	0.892	0.870	0.898	0.85	0.877
Mean	0.948	0.849	0.881	0.858	0.894	0.862	0.885	0.883	0.967	0.907	0.900	0.844	0.896	0.867	0.884
Std. Devn.	0.0092	0.0087	0.0126	0.0125	0.0075	0.0092	0.0053	0.0200	0.0454	0.0081	0.0311	0.0459	0.0077	0.0125	0.0072
% RSD	0.97	1.02	1.43	1.46	0.84	1.07	0.60	2.27	4.69	0.89	3.46	5.44	0.86	1.44	0.81
	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn
ME-16-1	0.78	0.830	0.78	0.840	0.800	0.83	0.850	0.81	0.750	0.819	0.871	0.820	0.801	0.81	0.797
ME-16-2	0.77	0.816	0.80	0.828	0.801	0.82	0.850	0.81	0.737	0.798	0.865	0.761	0.816	0.81	0.792
ME-16-3	0.78	0.807	0.80	0.825	0.800	0.83	0.850	0.79	0.724	0.808	0.885	0.776	0.804	0.82	0.786
ME-16-4	0.79	0.824	0.81	0.827	0.802	0.83	0.870	0.77	0.719	0.821	0.933	0.776	0.796	0.81	0.793
ME-16-5	0.80	0.827	0.79	0.816	0.802	0.82	0.870	0.81	0.792	0.803	0.835	0.780	0.804	0.81	0.783
ME-16-6	0.80	0.809	0.80	0.827	0.801	0.81	0.870	0.80	0.799	0.803	0.842	0.772	0.819	0.80	0.799
ME-16-7	0.79	0.825	0.80	0.848	0.800	0.83	0.850	0.83	0.786	0.812	0.810	0.799	0.803	0.83	0.793
ME-16-8	0.79	0.821	0.78	0.823	0.803	0.82	0.850	0.82	0.793	0.811	0.781	0.803	0.807	0.82	0.796
ME-16-9	0.79	0.822	0.78	0.825	0.807	0.82	0.850	0.82	0.778	0.819	0.791	0.781	0.800	0.82	0.784
ME-16-10	0.79	0.835	0.79	0.829	0.802	0.82	0.850	0.81	0.760	0.810	0.831	0.784	0.792	0.81	0.791
Mean	0.788	0.822	0.793	0.829	0.802	0.823	0.856	0.807	0.764	0.810	0.844	0.785	0.804	0.814	0.791
Std. Devn.	0.0092	0.0088	0.0106	0.0090	0.0021	0.0067	0.0097	0.0170	0.0300	0.0076	0.0459	0.0173	0.0082	0.0084	0.0055
% RSD	1.17	1.07	1.34	1.08	0.26	0.82	1.13	2.11	3.93	0.93	5.44	2.20	1.03	1.04	0.69

Note: Pb data from Laboratory 9 was removed for failing the “t” test.

REFERENCE MATERIAL: CDN-ME-16

Participating Laboratories:

(not in same order as listed in table of results)

Acme Analytical Laboratories Ltd., Vancouver

Actlabs-Ancaster, Ontario, Canada

Actlabs-Thunder Bay, Ontario, Canada

AGAT Laboratories, Ontario, Canada

AHK Geochem, Alaska, USA

ALS Chemex Laboratories, North Vancouver

Genalysis Laboratory, Australia

Inspectorate, Richmond, B.C. Canada

Omac Laboratories Ltd., Ireland

Skyline Assayers and Laboratories, Arizona, USA

SGS - Vancouver, B.C., Canada

Stewart Group, Kamloops, B.C., Canada

Alex Stewart Argentina SA

TSL Laboratories Ltd., Saskatoon

Ultra Trace Analytical Laboratories, Australia

Legal Notice:

This certificate and the reference material described in it have been prepared with due care and attention. However CDN Resource Laboratories Ltd. or Barry Smee accept no liability for any decisions or actions taken following the use of the reference material. Our liability is limited solely to the cost of the reference material.

Certified by

Duncan Sanderson, Certified Assayer of B.C.

Geochemist

Dr. Barry Smee, Ph.D., P. Geo

CDN Resource Laboratories Ltd.

#2, 20148 – 102nd Ave, Langley, B.C., Canada, V1M 4B4, 604-882-8422, Fax: 604-882-8466 (www.cdnlabs.com)

REFERENCE MATERIAL: CDN-ME-19

Recommended values and the “Between Lab” Two Standard Deviations

Gold	0.620 g/t	±	0.062 g/t	Certified value
Silver	103 g/t	±	7 g/t	Certified value
Copper	0.474 %	±	0.018 %	Certified value
Lead	0.98 %	±	0.06 %	Certified value
Zinc	0.75 %	±	0.04 %	Certified value

Note:   Standards with an RSD of near or less than 5% are certified; RSD’s of between 5% and 15% are Provisional; RSD’s over 15% are Indicated. Provisional and Indicated values cannot be used to monitor accuracy with a high degree of certainty.

PREPARED BY:
CERTIFIED BY:
INDEPENDENT GEOCHEMIST:
DATE OF CERTIFICATION: CDN Resource Laboratories Ltd.
Duncan Sanderson, B.Sc., Licensed Assayer of British Columbia
Dr. Barry Smee., Ph.D., P. Geo.
December 09, 2011

METHOD OF PREPARATION:
Reject ore material was dried, crushed, pulverized and then passed through a 270 mesh screen. The +270 material was discarded. The -270 material was mixed for 5 days in a double-cone mixer. Splits were taken and sent to 15 laboratories for round robin assaying.

ORIGIN OF REFERENCE MATERIAL:
The ore was supplied by Capstone Mining Corp. from the Minto Mine in Yukon, Canada. Mineralization is primary chalcopyrite and bornite pervasively disseminated and as stringers within foliated granodiorite units rich in secondary biotite. Sulphide mineralization is typically accompanied by magnetite. Gold is associated with the sulphide mineralization, typically intimately associated with bornite and rarely observed as free gold. 733 kg of the Minto ore was combined with 67 kg of a Au, Ag, Cu, Pb, Zn concentrate.

Approximate chemical composition (from whole rock analysis) is as follows:

	Percent		Percent
SiO2	61.8	MgO	1.3
Al2O3	15.1	K2O	2.9
Fe2O3	5.7	TiO2	0.4
CaO	3.4	LOI	2.7
Na2O	3.6	S	1.5
C	0.3

Statistical Procedures:

The final limits were calculated after first determining if all data was compatible within a spread normally expected for similar analytical methods done by reputable laboratories. Data from any one laboratory was removed from further calculations when the mean of all analyses from that laboratory failed a t test of the global means of the other laboratories. The means and standard deviations were calculated using all remaining data. Any analysis that fell outside of the mean ±2 standard deviations was removed from the ensuing data base. The mean and standard deviations were again calculated using the remaining data. This method is different from that used by Government agencies in that the actual “between-laboratory” standard deviation is used in the calculations. This produces upper and lower limits that reflect actual individual analyses rather than a grouped set of analyses. The limits can therefore be used to monitor accuracy from individual analyses, unlike the Confidence Limits published on other standards.

Assay Procedures:

Au: Fire assay pre-concentration, AA or ICP finish (30g sub-sample).
Ag, Cu, Pb, Zn: 4-acid digestion, AA or ICP finish.

REFERENCE MATERIAL: CDN-ME-19

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13	Lab 14	Lab 15
	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t	Au g/t
ME-19-1	0.597	0.593	0.580	0.612	0.581	0.581	0.571	0.646	0.690	0.614	0.578	0.629	0.639	0.620	0.685
ME-19-2	0.602	0.657	0.601	0.596	0.578	0.628	0.607	0.604	0.634	0.574	0.580	0.628	0.644	0.590	0.720
ME-19-3	0.593	0.613	0.620	0.601	0.631	0.610	0.723	0.651	0.603	0.598	0.595	0.614	0.626	0.610	0.695
ME-19-4	0.603	0.615	0.573	0.616	0.485	0.583	0.600	0.625	0.722	0.633	0.671	0.617	0.672	0.600	0.675
ME-19-5	0.598	0.598	0.592	0.602	0.596	0.558	0.640	0.642	0.586	0.637	0.600	0.599	0.635	0.600	0.685
ME-19-6	0.661	0.657	0.638	0.603	0.623	0.642	0.618	0.623	0.602	0.589	0.617	0.638	0.624	0.600	0.705
ME-19-7	0.655	0.66	0.630	0.610	0.597	0.652	0.603	0.600	0.612	0.603	0.542	0.624	0.667	0.630	0.680
ME-19-8	0.603	0.682	0.610	0.596	0.637	0.593	0.631	0.624	0.592	0.576	0.599	0.614	0.679	0.570	0.715
ME-19-9	0.585	0.699	0.627	0.606	0.629	0.587	0.662	0.685	0.654	0.603	0.680	0.606	0.639	0.570	0.685
ME-19-10	0.663	0.621	0.601	0.603	0.596	0.607	0.669	0.627	0.664	0.622	0.678	0.632	0.666	0.580	0.710
Mean	0.616	0.640	0.607	0.605	0.595	0.604	0.632	0.633	0.636	0.605	0.614	0.620	0.649	0.597	0.696
Std. Devn.	0.0307	0.0364	0.0217	0.0065	0.0442	0.0296	0.0435	0.0247	0.0456	0.0218	0.0473	0.0122	0.0200	0.0200	0.0159
% RSD	4.98	5.69	3.58	1.08	7.43	4.89	6.88	3.90	7.16	3.61	7.70	1.97	3.09	3.35	2.28
	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t	Ag g/t
ME-19-1	99.3	103	104	95.0	99.2	102	96	101	103.2	108	102	103	106	88	102
ME-19-2	100.2	106	110	100.6	99.1	99	100	99	107.9	107	102	100	105	87	105
ME-19-3	99.5	104	109	100.6	99.7	102	103	103	105.2	110	99	106	107	87	100
ME-19-4	97.3	102	106	100.9	100.7	102	92	102	108.2	108	106	107	107	90	101
ME-19-5	98.4	103	106	101.1	99.4	102	93	101	106.9	109	98	107	108	98	103
ME-19-6	100.6	105	108	100.5	100.4	104	96	102	102.2	110	101	104	105	98	104
ME-19-7	103.0	106	109	100.1	100.5	105	96	98	105.2	110	100	106	104	89	100
ME-19-8	98.9	103	107	100.0	100.8	104	93	98	111.0	113	103	105	107	90	103
ME-19-9	106.8	108	106	99.6	101.2	103	95	98	108.3	112	101	103	105	85	103
ME-19-10	101.6	104	106	99.7	101.3	103	91	102	107.0	108	101	103	103	93	102
Mean	101	104	107	100	100	103	96	100	107	110	101	104	106	91	102
Std. Devn.	2.7257	1.8379	1.8529	1.7514	0.8193	1.6465	3.6893	1.9551	2.6269	1.9003	2.2136	2.2211	1.5670	4.5031	1.6364
% RSD	2.71	1.76	1.73	1.75	0.82	1.60	3.86	1.95	2.47	1.74	2.19	2.13	1.48	4.98	1.60

Note: Ag data from Lab 14 was removed for failing the t test.

REFERENCE MATERIAL: CDN-ME-19

Results from round-robin assaying:

	Lab 1	Lab 2	Lab 3	Lab 4	Lab 5	Lab 6	Lab 7	Lab 8	Lab 9	Lab 10	Lab 11	Lab 12	Lab 13	Lab 14	Lab 15
	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu	% Cu
ME-19-1	0.474	0.454	0.47	0.470	0.47	0.473	0.47	0.47	0.488	0.486	0.445	0.458	0.47	0.44	0.48
ME-19-2	0.479	0.482	0.47	0.466	0.46	0.471	0.49	0.46	0.487	0.479	0.438	0.465	0.49	0.44	0.47
ME-19-3	0.473	0.456	0.47	0.471	0.47	0.470	0.50	0.47	0.481	0.480	0.432	0.460	0.47	0.44	0.47
ME-19-4	0.472	0.460	0.48	0.470	0.46	0.473	0.48	0.47	0.486	0.474	0.436	0.476	0.47	0.43	0.48
ME-19-5	0.486	0.457	0.47	0.469	0.46	0.473	0.48	0.47	0.487	0.479	0.439	0.466	0.46	0.44	0.49
ME-19-6	0.483	0.474	0.48	0.468	0.47	0.478	0.47	0.48	0.480	0.488	0.449	0.467	0.48	0.44	0.48
ME-19-7	0.487	0.463	0.48	0.460	0.46	0.479	0.47	0.47	0.482	0.485	0.418	0.482	0.48	0.44	0.48
ME-19-8	0.462	0.459	0.48	0.469	0.46	0.478	0.47	0.47	0.485	0.489	0.432	0.472	0.48	0.44	0.48
ME-19-9	0.483	0.477	0.47	0.469	0.46	0.477	0.47	0.48	0.485	0.496	0.442	0.464	0.48	0.44	0.48
ME-19-10	0.483	0.470	0.48	0.468	0.46	0.479	0.48	0.47	0.491	0.483	0.436	0.463	0.45	0.43	0.47
Mean	0.478	0.465	0.475	0.468	0.463	0.475	0.478	0.471	0.485	0.484	0.437	0.467	0.473	0.438	0.478
Std. Devn.	0.0078	0.0098	0.0053	0.0031	0.0048	0.0034	0.0103	0.0057	0.0034	0.0063	0.0085	0.0074	0.0116	0.0042	0.0063
% RSD	1.64	2.11	1.11	0.66	1.04	0.73	2.16	1.21	0.70	1.30	1.95	1.58	2.45	0.96	1.32
	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb	% Pb
ME-19-1	1.03	0.971	1.01	0.994	0.97	0.96	0.94	0.95	1.022	0.913	0.953	0.979	0.94	0.78	0.96
ME-19-2	1.05	0.971	1.01	1.007	0.94	0.97	0.97	0.97	1.020	0.931	0.935	0.960	0.96	0.77	0.98
ME-19-3	1.03	0.983	1.01	1.018	0.97	0.98	0.97	0.97	1.009	0.944	0.928	0.971	0.96	0.75	0.99
ME-19-4	1.01	0.983	1.03	1.033	0.97	0.97	0.96	0.97	1.017	0.923	0.934	0.975	0.95	0.76	0.97
ME-19-5	1.03	0.983	1.01	1.026	0.97	0.96	0.96	0.97	1.032	0.947	0.912	0.987	0.96	0.78	1.00
ME-19-6	1.05	0.954	1.03	1.002	0.96	0.96	0.95	0.98	1.023	0.941	0.921	0.967	0.94	0.76	0.98
ME-19-7	1.05	0.950	1.03	1.003	0.97	0.98	0.95	0.96	1.025	0.942	0.923	0.974	0.90	0.75	0.96
ME-19-8	1.00	0.983	1.03	0.999	0.98	0.98	0.94	0.96	1.029	0.951	0.956	0.961	0.98	0.76	0.98
ME-19-9	1.02	0.966	1.02	1.006	0.97	0.97	0.95	0.97	1.037	0.963	0.943	0.953	0.94	0.77	0.98
ME-19-10	1.03	0.959	1.03	0.997	1.00	0.97	0.97	0.95	1.038	0.951	0.923	0.957	0.92	0.76	0.97
Mean	1.03	0.97	1.02	1.01	0.97	0.97	0.96	0.97	1.03	0.94	0.93	0.97	0.95	0.76	0.98
Std. Devn.	0.0181	0.0128	0.0099	0.0129	0.0149	0.0082	0.0117	0.0097	0.0090	0.0147	0.0143	0.0107	0.0227	0.0107	0.0125
% RSD	1.75	1.32	0.97	1.28	1.54	0.84	1.23	1.01	0.88	1.56	1.53	1.11	2.41	1.41	1.28
	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn	% Zn
ME-19-1	0.753	0.767	0.73	0.767	0.76	0.77	0.75	0.75	0.816	0.712	0.764	0.764	0.77	0.58	0.71
ME-19-2	0.726	0.759	0.75	0.764	0.73	0.76	0.77	0.75	0.802	0.710	0.756	0.749	0.78	0.57	0.70
ME-19-3	0.749	0.761	0.74	0.778	0.74	0.77	0.77	0.75	0.790	0.728	0.752	0.762	0.77	0.57	0.70
ME-19-4	0.741	0.764	0.74	0.768	0.71	0.77	0.77	0.74	0.805	0.702	0.770	0.762	0.77	0.56	0.70
ME-19-5	0.765	0.756	0.73	0.765	0.71	0.77	0.77	0.75	0.823	0.722	0.731	0.768	0.77	0.57	0.68
ME-19-6	0.764	0.743	0.76	0.764	0.72	0.78	0.75	0.76	0.816	0.721	0.746	0.756	0.76	0.56	0.72
ME-19-7	0.772	0.731	0.75	0.764	0.70	0.78	0.76	0.75	0.817	0.721	0.747	0.763	0.75	0.56	0.70
ME-19-8	0.724	0.759	0.76	0.762	0.72	0.78	0.75	0.75	0.834	0.732	0.768	0.757	0.80	0.56	0.69
ME-19-9	0.744	0.755	0.75	0.768	0.72	0.77	0.76	0.76	0.849	0.740	0.761	0.757	0.76	0.57	0.71
ME-19-10	0.759	0.753	0.76	0.770	0.75	0.78	0.78	0.75	0.835	0.728	0.752	0.750	0.75	0.56	0.71
Mean	0.75	0.75	0.75	0.77	0.73	0.77	0.76	0.75	0.82	0.72	0.75	0.76	0.77	0.57	0.70
Std. Devn.	0.0162	0.0106	0.0116	0.0047	0.0190	0.0067	0.0106	0.0057	0.0175	0.0113	0.0118	0.0061	0.0148	0.0070	0.0114
% RSD	2.16	1.41	1.55	0.61	2.61	0.87	1.39	0.76	2.14	1.56	1.56	0.81	1.92	1.24	1.62

Note:  Cu data from Lab 11 was removed for failing the t test.

Pb and Zn data from Lab 14 was removed for failing the t test.

REFERENCE MATERIAL: CDN-ME-19

Participating Laboratories:
 (not in same order as listed in table of results)

Acme Analytical Laboratories Ltd., Vancouver

Actlabs-Ancaster, Ontario, Canada

Actlabs-Thunder Bay, Ontario, Canada

AGAT Laboratories, Ontario, Canada

AHK Geochem, Alaska, USA

ALS Chemex Laboratories, North Vancouver

Genalysis Laboratory, Australia

Inspectorate, Richmond, B.C. Canada

Omac Laboratories Ltd., Ireland

Skyline Assayers and Laboratories, Arizona, USA

SGS - Vancouver, B.C., Canada

Stewart Group, Kamloops, B.C., Canada

Alex Stewart Argentina SA

TSL Laboratories Ltd., Saskatoon

Ultra Trace Analytical Laboratories, Australia

Legal Notice:

This certificate and the reference material described in it have been prepared with due care and attention. However CDN Resource Laboratories Ltd. or Barry Smee accept no liability for any decisions or actions taken following the use of the reference material. Our liability is limited solely to the cost of the reference material.

Certified by

 

Duncan Sanderson, Certified Assayer of B.C.

Geochemist

Dr. Barry Smee, Ph.D., P. Geo

P.O Box 18-142, Glen Innes 1743
Auckland, New Zealand. P 64 9 634 7696
F 64 9 634 6896E sales@rocklabs.comwww.rooklabs.com

Certificate of Analysis

Reference Material OxD87

Recommended Gold Concentration: 0.417 µ/g
95% Confidence Interval: +/- 0.004 µ/g

The above values apply only to product in jars or sachets which have an identification number

within the following range: 224 699 - 227 965.

Prepared and Certified By:	Malcolm Smith BSc, FNZIC Rocklabs Reference Materials 40 Oakford Park Crescent, Greenhithe Auckland 0632 NEW ZEALAND Email: Malcolm@MSRML.co.nz Telephone: +64 9 444 3534
Date of Certification:	29 November 2010
Certificate Status:	Original
Available Packaging:	This reference material has been packed in wide­mouthed jars that contain 2.5 kg of product. The contents of some jars may be subsequently repacked into sealed polyethylene sachets.
Origin of Reference Material:	Basalt and feldspar minerals with minor quantities of finely divided gold-containing minerals that have been screened to ensure there is no gold nugget effect.
Supplier of Reference Material:	ROCKLABS Ltd P O Box 18 142 Glen Innes Auckland 1743 NEW ZEALAND Email: sales@rocklabs.com Website: www.rocklabs.com
Certificate of Analysis, ROCKLABS Reference Material OxD87, 29 November, 2010. Page 1 of 6.

Description: The reference material is a light grey powder that has been well mixed and a homogeneity test carried out after the entire batch was packaged into wide-mouthed jars. There is no soil component. The product contains crystalline quartz and therefore dust from it should not be inhaled.

The approximate chemical composition is:
(Uncertified Values)

	%
SiO2	58.85
Al2O3	16.51
Na2O	3.16
k2o	8.24
CaO	3.21
MgO	3.08
TiO2	0.91
MnO	0.06
P2O5	0.24
Fe2O3	4.72
L O I	0.40

Intended Use:	This reference material is designed to be included with every batch of samples analysed and the results plotted for quality monitoring and assessment purposes.
Stability:	The container (jar or sachet) and its contents should not be heated to temperatures higher than 50 °C. The reference material is stable, with weight changes of less than 0.5 % at extremes of naturally occurring temperature and humidity conditions.
Method of Preparation:	Pulverized basalt rock and feldspar minerals were blended with finely pulverized and screened, gold-containing minerals. Once the powders were uniformly mixed the composite was placed into 3267 wide-mouthed jars, each bearing a unique number. 60 jars were randomly selected from the packaging run and material from these jars was used for both homogeneity and consensus testing.

Certificate of Analysis, ROCKLABS Reference Material OxD87, 29 November, 2010. Page 2 of 6.

   Homogeneity Assessment:

An independent laboratory carried out gold analysis by fire assay of 30 g portions, using an AAS finish. Steps were taken to minimize laboratory method variation in order to better detect any variation in the candidate reference material.

The contents of six randomly selected jars were compacted by vibration (to simulate the effect of freighting) and five samples removed successively from top to bottom from each jar. In addition, five samples were removed from the last jar in the series. A sample was also removed from the top of each of the 60 jars randomly selected from the 3267 jars in the batch. The results of analysis of the 95 samples (randomly ordered and then consecutively numbered before being sent to the laboratory) produced a relative standard deviation of 1.2 %.

Analytical Methodology:

Once homogeneity had been established, two sub-samples were submitted to a number of well-recognized laboratories in order to assign a gold value by consensus testing. The sub-samples were drawn from the 60 randomly selected jars and each laboratory received samples from two different jars. Indicative concentration ranges were given. All laboratories used fire assay for the gold analysis, with most using an instrument finish and four using a gravimetric finish.

Calculation of Certified Value:

Results for gold were returned from 47 laboratories. Statistical analysis to identify outliers was carried out using the principles detailed in sections 7.3.2 - 7.3.4, ISO 5725-2: 1994. Assessment of each laboratory’s performance was carried out on the basis of z-scores, partly based on the concept described in ISO/IEC Guide 43-1. Details of the criteria used in these examinations are available on request. As a result of these statistical analyses, eight sets of results were excluded for the purpose of assigning a gold concentration value to this reference material. A recommended value was thus calculated from the average of the remaining n = 39 sets of replicate results. The 95 % confidence interval was estimated using the formula:-

 (where X is the estimated average, s is the estimated standard deviation of the laboratory averages, and t is the 0.025 tail-value from Student’s t-distribution with n- 1 degrees of freedom). The recommended value is provided at the beginning of the certificate in ^g/g (ppm) units. A summary of the results used to calculate the recommended value is listed on page 4 and the names of the laboratories that submitted results are listed on page 5. The results are listed in increasing order of the individual laboratory averages.

Statistical analysis of the consensus test results has been carried out by independent statistician, Tim Ball.

Certificate of Analysis, ROCKLABS Reference Material OxD87, 29 November, 2010. Page 3 of 6.

Summary of Results Used to Calculate Gold Value

(Listed in increasing order of individual laboratory averages)

Gold (ppm)
Sample 1	Sample 2	Average
0.40	0.39	0.395
0.4055	0.3890	0.3973
0.395	0.400	0.3975
0.41	0.39	0.400
0.40	0.40	0.400
0.40	0.40	0.400
0.404	0.399	0.4015
0.41	0.40	0.405
0.408	0.404	0.4060
0.406	0.409	0.4075
0.408	0.411	0.4095
0.41	0.41	0.410
0.409	0.413	0.4110
0.411	0.411	0.4110
0.4100	0.4150	0.4125
0.420	0.405	0.4125
0.422	0.403	0.4125
0.4155	0.4100	0.4128
0.408	0.418	0.4130
0.415	0.415	0.4150
0.421	0.411	0.4160
0.413	0.422	0.4175
0.423	0.415	0.4190
0.42	0.42	0.420
0.42	0.42	0.420
0.421	0.423	0.4220
0.4265	0.4210	0.4238
0.430	0.420	0.4250
0.43	0.42	0.425
0.4270	0.4265	0.4268
0.431	0.425	0.428
0.427	0.432	0.4295
0.43	0.43	0.430
0.4325	0.4275	0.4300
0.4305	0.4345	0.4325
0.44	0.43	0.435
0.4420	0.4360	0.4390
0.44	0.44	0.440
0.440	0.447	0.4435
Average of 39 sets = 0.417 ppm Standard deviation of 39 sets = 0.013 ppm Note: this standard deviation should not be used as a basis to set control limits when plotting results from an individual laboratory. Relative standard deviation = 3.1 % 95% Confidence interval for average = 0.004 ppm

Certificate of Analysis, ROCKLABS Reference Material OxD87, 29 November, 2010. Page 4 of 6.

Participating Laboratories

Australia	ALS Minerals, Kalgoorlie ALS Minerals, Orange ALS Minerals, Perth ALS Minerals, Townsville Amdel - Bureau Veritas, Adelaide Amdel - Bureau Veritas, Kalgoorlie Genalysis Laboratory Services, Perth Independent Assay Laboratories, Perth SGS Minerals Services, Perth Ultra Trace Pty Ltd, Perth
Burkina Faso	ALS Minerals, Burkina Faso
Canada	Acme Analytical Laboratories (Vancouver) Ltd ALS Minerals, Val d’Or ALS Minerals, Vancouver Bourlamaque, Quebec Inspectorate Exploration & Mining Services Ltd, Richmond Loring Laboratories (Alberta) Ltd, Calgary SGS Canada Inc, Vancouver SGS Minerals Services, Lakefield Techni-Lab S.G.B. Abitibi Inc, Quebec TSL Laboratories Inc, Saskatoon
Chile	Acme Analytical Laboratories Ltd, Santiago ALS Minerals, La Serena
Cote d’Ivoire	Bureau Veritas Mineral Laboratories
Guyana	Loring Laboratories (Guyana) Ltd
Ireland	OMAC Laboratories
Kyrgyz Republic Mali	Stewart Assay and Environmental Laboratories LLC, Kara-Balta
Mali	ALS Minerals, Bamako
New Zealand	Amdel - Bureau Veritas, Otago Amdel - Bureau Veritas, Reefton SGS Minerals Services, Waihi
Peru	ALS Minerals, Lima Inspectorate Services Peru S.A.C., Callao Minera Yanacocha SRL - Newmont, Lima
Russia	Irgiredmet, Irkutsk
South Africa	AB Analytical Laboratory Services, Boksburg ALS Minerals, Johannesburg Anglo Research, Johannesburg Goldfields West Wits Analytical Laboratory Performance Laboratories, Allanridge Performance Laboratories, Randfontein SGS South Africa (Pty) Ltd, Johannesburg Vaal River Chemical Laboratory, AngloGold Ashanti
USA	ALS Minerals, Reno Barrick Goldstrike - Met Services Newmont Mining Corporation, Carlin Laboratory

Certificate of Analysis, ROCKLABS Reference Material OxD87, 29 November, 2010. Page 5 of 6.

Instructions and Recommendations for Use:

Weigh out quantity usually used for analysis and analyze for total gold by normal procedure. Homogeneity testing has shown that consistent results are obtainable for gold when 30g portions are taken for analysis.

We quote a 95% confidence interval for our estimate of the declared value. This confidence interval reflects our uncertainty in estimating the true value for the gold content of the reference material. The interval is chosen such that, if the same procedure as used here to estimate the declared value were used again and again, then 95% of the trials would give intervals that contained the true value. It is a reflection of how precise the trial has been in estimating the declared value. It does not reflect the variability any particular laboratory will experience in its own repetitive testing.

Some users in the past have misinterpreted this confidence interval as a guide as to how different an individual test result should be from the declared value. Some mistakenly use this interval, or the standard deviation from the consensus test, to set limits for control charts on their own routine test results using the reference material. Such use inevitably leads to many apparent out-of-control points, leading to doubts about the laboratory’s testing, or of the reference material itself.

A much better way of determining the laboratory performance when analysing the reference material is to accumulate a history of the test results obtained, and plot them on a control chart. The appropriate centre line and control limits for this chart should be based on the average level and variability exhibited in the laboratory’s own data. This chart will provide a clear picture of the long-term stability or otherwise of the laboratory testing process, providing good clues as to the causes of any problems. To help our customers do this more simply for themselves, we can provide a free Excel template that will produce sensible graphs, with intelligently chosen limits, from the customer’s own data.

Legal Notice:

This certificate and the reference material described in it have been prepared with due care and attention. However ROCKLABS Ltd, Scott Technology Ltd and Tim Ball Ltd accept no liability for any decisions or actions taken following the use of the reference material.

References:

For further information on the preparation and validation of this reference material please contact Malcolm Smith.

Certifying Officer	Independent Statistician
M G Smith BSc, FNZIC	Tim Ball BSc (Hons)

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL 30, 2013 | RELEASED DATE: JULY 25, 2013

APPENDIX C

METALLURGICAL INVESTIGATION, SUMMARY REPORTS

METALLURGICAL TESTING ON
SAMPLES FROM SANTA ELENA &
CRUZE DE MAYO GOLD-SILVER
DEPOSITS - MEXICO

Prepared for:	SILVERCREST MINES INC.
	Suite 501 - 570 Granville Street
	Vancouver, BC
	Canada V6C 3P1
Attention:	Mr. Brent McFarlane, VP Operations
	Cc: Mr. Salvador A. Salinas, Country Manager
	Cc: Mr. John Fox, P.Eng., Laurion Consulting Inc.
Prepared by:	Inspectorate Exploration & Mining Services Ltd
	Metallurgical Division
	11620 Horseshoe Way
	Richmond, B.C.
	V7A 4V5
Project No.:	1103507

Reported by: Alice Shi, Ph.D. General Manager	Reviewed by: Michael Redfearn, P.Eng. Director Metallurgical Services

Date: December 31, 2012

Santa Elena Gold-Silver Project  1

	TABLE OF CONTENT
		Page No.
1	SUMMARY	3
2	INTRODUCTION	5
3	PROCEDURES	6
3.1	SAMPLE PREPARATION AND CHARACTERIZATION	6
3.2	ASSAY PROCEDURE	7
3.3	GRINDING AND SCREENING	7
3.4	CYANIDATION	8
3.5	BOND BALL MILL WORK INDEX DETERMINATION	9
4	RESULTS AND DISCUSSION	10
4.1	SAMPLE CHARACTERIZATION	10
4.1.1	Chemical Analysis	10
4.1.2	Comminution Studies	10
4.1.3	Mineralogical Examination	11
4.2	CYANIDATION OF SAMPLES FROM SANTA ELENA DEPOSIT	12
4.2.1	Baseline Bottle Roll Tests on Ground Whole Ore	12
4.2.2	Leach Optimization	14
4.2.3	Confirmatory Leach on New Santa Elena Feed	16
4.2.4	Leach Tailing Production	18
4.2.5	Mineralogical Investigation of Leach Residue	19
4.3	CYANIDATION OF CRUZ DE MAYO SAMPLE AND A HI-MN SAMPLE	19
4.4	SUPPORT STUDIES FOR TAILINGS EVALUATION	21
4.4.1	Cyanide Detoxification	21
4.4.2	Characterization of Detoxified Sample	21
4.4.3	Settling Test	22
4.4.4	Vacuum Filtration Test	22
5	CONCLUSIONS AND RECOMMENDATIONS	24

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project 2

APPENDIXES

Appendix I - Sample Receiving Log
Appendix II - Head Characterization
Appendix III - Metallurgical Test Results
Appendix IV - Mineralogical Report

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  3

1 SUMMARY

Inspectorate Exploration & Mining Services Ltd. - Metallurgical Division has carried out a scoping metallurgical testing program on samples originating from SilverCrest’s Santa Elena and Cruz de Mayo properties in Mexico to evaluate the samples’ leaching characteristics.

As was originally requested, the study covers the following major topics: ore sample characterization, bottle roll cyanidation as well as preliminary tailings evaluation.

Four samples including two from Santa Elena deposit, one from Cruze de Mayo deposit and one high manganese sample (Hi-Mn) were tested. The two samples from the Santa Elena deposit assayed 1.46- and 3.31 g/t Au, and 38.5- and 119.1g/t Ag, respectively. Sulfur content in the Santa Elena ore was <0.03%. The sample from the Cruze de Mayo deposit assayed 0.5g/t Au and 148.5g/t Ag, while the Hi-Mn sample assayed 5.04g/t Au and 89.5g/t Ag. All four samples tested contained significant amount of Mn, ranging from 559ppm to 4.4%. Arsenic was in a range of 13- to 55 ppm.

Comminution studies showed that the Santa Elena ore is of moderate hardness with 14.0 kWh/t Bond rod-mill work index and 14.8 kWh/t Bond ball-mill work index. The Cruze de Mayo composite sample showed a harder character of a 19.3kWh/t Bond ball mill work index. Ore microscopy studies indicated that the Cruze de Mayo sample contains trace of gold in native form of less than 5 mm. Argentite and pyrargyrite-proustite are the main silver carriers embedded in silica gangue, which also carries isolated inclusions or flakes of native silver and/or electrum. The Hi-Mn sample consists basically of pyrolusite. Silver as electrum occurs as <20 mm inclusions together with native gold in association with limonite.

Baseline leaching of the Santa Elena ore for 72-hours in 0.5- to 1.5 g/L NaCN recovered 95.6% Au and 56.6% Ag on average at a grind size of less than 120 mm. Finer grinding to ~74 mm improved the baseline Au recovery to 97.4%, but Ag remained refractory throughout with residue grade averaging ~52 g/t Ag. Increasing the cyanide concentration in the range of 0.5- to 1.5 g/L cyanide was of no significant benefit to gold extraction, but was to silver.

Santa Elena Gold-Silver Project 4

Cyanidation of Santa Elena ore in 1.0g/L NaCN with single solution replacement at the target grind of 100 mm improved silver extraction from the baseline 56.6% to ~70%. Solution replacement also benefited both gold and silver leach kinetics. By replacing leach liquor, residence time could be shortened to 30 hours for gold. Increasing the number of solution exchanges and cyanide dosages as well as the addition of oxidants resulted in negligible gains in gold and silver extraction.

Leaching of the Cruz de Mayo sample at 40% solids in 3g/L NaCN with single solution replacement extracted 51% Au and 54% Ag. A reasonable Au extraction of 92% was achieved on the Hi-Mn sample following identical leach conditions, but silver extraction remained at a low 60%.

Settling test on Santa Elena leach tailing with 40 g/t of Percol Magnafloc 368 yielded a 24-hour raked underflow density of 79.2% solids and a unit thickener area of 0.07 m2/h/tpd of dry solids. Vacuum filtration with Polypropylene type Filter cloth PO-808HF at 70% solids had capacities of 32.6 kg/m2/h solids and 64 L/m2/h of filtrate at standard conditions.

Scoping detoxification studies using the INCO SO2/air procedure suggested that less than 1mg/L of CN in the effluent is achievable

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project 5

2INTRODUCTION

Inspectorate - Metallurgical Division was retained by Mr. John Fox, P.Eng., from Laurion Consulting Inc. on behalf of SilverCrest Mines Inc. to perform metallurgical testing on gold and silver bearing samples collected from SilverCrest's Santa Elena Mine located 150km northeast of Hermosillo, near Banamichi in the State of Sonora, Mexico. The optimum process conditions developed from the Santa Elena ore was further confirmed on samples representing the Cruze de Mayo deposit and a high manganese sample (Hi-Mn).

The stated objectives of this metallurgical study program are to conduct baseline investigations to determine leaching characteristics of the samples, and to provide adequate data to optimize gold & silver recoveries.

The test work consisted of head characterization, bottle roll cyanide leaching and preliminary tailings evaluation. Head characterization including assaying and hardness testing were conducted on feed samples from Santa Elena and Cruz de Mayo deposits. Optical mineralogy examination was carried out on Cruz de Mayo feed and Santa Elena leach residue. Metallurgical testing consisting of cyanide leaching and detoxification of cyanide leach tailings as well as detoxified tailings characterization was mainly conducted on the Santa Elena samples. Confirmatory bottle roll cyanide leach following preferred leach conditions was conducted on a composite sample from Cruz de Mayo deposit and a Hi-Mn sample.

This report summarizes the laboratory test procedures and results of the scoping study.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  6

3PROCEDURES

Based on a detailed proposal, specific formats in the overall approach and methods to be used were agreed upon between Inspectorate and the client’s representative (Mr. John Fox from Laurion Consulting).

3.1SAMPLE PREPARATION AND CHARACTERIZATION

Two shipments were received at Inspectorate on July 26th of 2011 and June 6th of 2012 respectively. Sample identification and compositing information are presented in Table 1. Detailed Sample Receiving Logs are provided in Appendix 1.

Table 1. Sample and Compositing List

Deposit	Composite ID	Sample Label	Top Size	Weight (kg)
Santa Elena	Comp. NIVEL 4	4NIVEL 622928	3"	3.8
		COMPOSITO-NIVEL 4	6M	15.7
		COMPOSITO-NVEL 4	6M	14.1
		COMPOSITO-NIVEL 4	6M	30.0
Santa Elena	Comp.4 ± 0 Nivel	Composito 4 + 0 Nivel - 4"	6"	20.7
		Composito 4 + 0 Nivel - 4"	6"	29.9
Cruz de Mayo	Comp. CM	CM2-02	3"	22.0
		CM2-07	3"	17.4
		CM2-15	3"	16.9
		CM2-24	3"	17.6
		CM2.5-61	3"	17.5
		CM2.5-66	3"	20.0
		CM2.5-70	3"	18.2
		CM2.5-88	3"	19.6
Hi Manganese	Comp. Hi- Mn	VOL15Q HIGH Mn	6M	28.0

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  7

Each composite sample was prepared and assayed individually. For Comp. CM and Comp. 4±0 Nivel, the samples were first crushed to 1.5 inches. After removal of sub-samples for comminution testing, the remain samples were further crushed to 6-Tyler mesh and then blended three times prior to rotary split into 2 kg test charges for the metallurgical test program. Head aliquots were also split off from each composite for head characterization.

3.2ASSAY PROCEDURE

The Au and Ag analyses were done by standard fire assay procedures. A weighed sample was mixed with litharge and the required soda ash and borax fluxes to produce a fluid slag at 1000oC. After cooling, the lead button was separated from the slag and re-melted in a cupel, which absorbed the oxidized base metals. The remaining precious metals bead was weighed, dissolved in acid and analyzed by Atomic Absorption (AA) Spectrophotometry. A multi­element ICP scan using standard multi-acid digestion procedure was also employed. Carbon and sulfur speciation used microbalance measurements and gas detections in a Leco furnace, while As was determined by standard wet assays.

3.3GRINDING AND SCREENING

Wet grinding was performed in a dedicated stainless steel laboratory rod mill on 2kg charges of composite sample, at a 65% solids pulp density. Test grinds determined the time required to achieve reliable target grind size distributions.

Particle size analysis was performed using a Rotap™, equipped with 20 cm (8”) diameter test sieves, and stacked in ascending mesh sizes. The sample was initially wet screened at 37 mm (400 Tyler™ mesh). The +37 mm fraction was then dry screened through stacked sieves. Each sieved fraction was collected and weighed for calculating the individual and cumulative percentage retained.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  8

3.4CYANIDATION

The bottle roll cyanide leaching tests were carried out at a pulp density of 40 wt.% solids. Prior to adding sodium cyanide (NaCN), the alkalinity was adjusted with hydrated lime to pH 10 and further maintained at this level. NaCN levels were established and maintained at specific targets and the dissolved oxygen concentration was monitored for the duration of the test.

Six baseline leach tests (C1-C6) were carried out on the NIVEA 4 Comp. in 0.5-, 1.0- and 1.5 g/L NaCN at target grinds of 120 mm and 74 mm, respectively, to evaluate the effect of cyanide concentration and grind size on metals’ recoveries. Intermediate solution samples were removed and analyzed to determine gold and silver dissolutions at 2, 4, 7, 24, 30, 48 and 54 hours of retention time. These tests were terminated after 72 hours with filtration of pregnant leachate solution (PLS). The solid residues were displacement-washed with cyanide solution, followed by two water rinses. The PLS and the final residue were analyzed for gold and silver content for metallurgical balance.

Leach optimization tests (C7-C12) were performed on NIVEA 4 Comp. at a target P80 of 100 mm grind with solution replacement at various retention times. Three cyanide concentrations of 1.0-, 1.5- and 2 g/L cyanide were tested. The addition of lead nitrite and hydrogen peroxide was also tested to assist metals’ dissolution.

Confirmatory leach tests (C13-C17) were conducted on the CM Comp., the Hi- Mn Comp. as well as the new Santa Elena feed sample - Comp.4 ± 0 Nivel to evaluate the samples’ response to preferred leach conditions.

At the end of the test program, two bulk leach tests (C18, C19) were performed to generate tailings for thickening studies at Pocock Industrial.

Reagent concentrations were determined using standard titration methods. The sodium cyanide concentration was determined by titrating with silver nitrate and using para-dimethylamino rhodanine as an indicator.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  9

A detailed test report including a metallurgical balance and the test procedure was prepared for each test, as provided in Appendix III.

3.5BOND BALL MILL WORK INDEX DETERMINATION

The Bond ball-mill work index was measured in a Bico-Braun® laboratory mill, with a standard ball charge medium. The feed size was 100% passing 6-Tyler mesh, and the specified closing screen of 150-Tyler mesh for the product. The bulk density was determined and the feed size distribution was measured to yield undersize, at a specified number of mill revolutions. The product was discharged to sift out the media and perform a screen analysis. The oversize was retained for the next cycle, replacing the weight of the undersize with an equivalent weight of the original feed. The undersize weight minus the feed undersize amount yielded the net product weight. The procedures were repeated for each cycle until a constant circulating load was achieved after 6-7 cycles. This allowed the calculation of a stable amount of net product per revolution, and the calculation of the ball mill work index from the empirical Bond formula.

Other comminution testing was performed at Hazen Research.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  10

4RESULTS AND DISCUSSION

Various sub-topics are to be addressed, as organized in the following sections. Details of the individual test results are presented in the Appendices, while only major findings are highlighted in the text.

4.1SAMPLE CHARACTERIZATION

Preparation and compositing procedures were presented above. Chemical analysis, hardness and mineralogy examination are discussed below.

4.1.1Chemical Analysis

Partial assays for the samples tested are presented in Table 2. Multi-element ICP results are provided in Appendix II.

Table 2. Main Assays on Composites

Items	Units	Comp. NIVEL 4	Comp.4 ± 0 Nivel	Comp. CM	Comp. Hi-Mn
Au	ppm	3.31	1.46	0.50	5.04
Ag	ppm	119.1	38.5	148.5	89.5
S (T)	%	0.03	0.02
Sulfide	%	0.02	<0.01
C	%	1.88	0.06
C-Org	%	0.11	0.1
As	ppm	13	31	55	47
Mn	ppm	2107	1698	559	44246

4.1.2 Comminution Studies

The standard Bond ball mill work index of 14.8(9) kWh/t at a closing screen size of 150-Tyler mesh (105 mm) indicated a medium hard character of the Santa Elena ore. The 19.3 kWh/t work index on the CM Composite showed a harder character of the Cruz de Mayo ore.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  11

Table 3. Comminution Test Summary

Sample ID	Bond Ball Mill Work Index (Bwi) W (kWh/tonne)	Bond Rod Mill Worl Index (Rwi) W (kWh/tonne)	Bond Crusher Impact Work Index (Cwi) kWh/tonne	Bond Abrasion Index (Ai) g	Bulk Density* g/cm3
Comp. NIVEL 4	14.8				1.81
Comp.4 ± 0 Nivel	14.9	14.0			1.66
Comp. CM	19.3		13.6	1.1267	1.69
* at the nominal 6-Tyler mesh crush size

Bond rod mill hardness testing on the new Santa Elena feed at Hazen Research returned a work index value of 14.0 kWh/t, which is in agreement with the Bond ball mill work index results. The comminution test reports are provided in Appendix II.

4.1.3Mineralogical Examination

Ore microscopic investigations were carried out on representative 6-Tyler mesh crushed CM composite and the Hi-Mn composite for gold and silver mineralogy. The mineralogy study was performed by Dr. Lehne at Lehne & Associates Applied Mineralogy in Germany.

The mineralogy examination revealed that the CM composite consists of silica gangue that possesses a pronounced iron staining due to finely disseminated limonite. Silver carriers are argentite and pyrargyrite-proustite embedded in silica gangue. The gangue also carries isolated inclusions or flakes of native silver and/or electrum. Trace of gold occurs in native form in less than 5-micron size.

The Hi-Mn composite is characterized by a distinct manganese mineralization consisting essentially of pyrolusite. Specific silver minerals have not been observed. Silver is contained in electrum which occurs as <20 micron inclusions together with native gold in association with limonite. The small particle sizes of the noble metal phases imply rather resistant locking textures. Detailed findings are provided in Appendix IV.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  12

4.2CYANIDATION OF SAMPLES FROM SANTA ELENA DEPOSIT

Bottle roll cyanide leaching was mainly performed on Santa Elena NIVEL 4 whole ore at various grind sizes and different cyanide dosages to assess gold and silver recovery potentials. A series of test grinds were conducted to establish reliable laboratory grind time requirements prior to metallurgical testing.

4.2.1Baseline Bottle Roll Tests on Ground Whole Ore

Baseline bottle roll tests (C1 to C6) were conducted on ground whole ore to assess the grind- and cyanide-sensitivity of gold and silver recoveries as well as kinetics and reagent requirements. The leach tests were carried out for 72 hours at 40% solids at target P80 grinds of 120 mm and 74 mm. Three cyanide dosages of 0.5-, 1.0- and 1.5 g/L cyanide were tested at each grind, and pH >10.0 was maintained with hydrated lime. The baseline test results are summarized in Table 4, whilst Au and Ag leach profiles are plotted in Figures 1 and 2.

Table 4. Baseline Cyanidation Performance

	Test Conditions		Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
Test No	P80 Size, µm	NaCN, g/L	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C1	121	0.5	3.31	119.1	2.66	113.0	94.2	42.4	0.16	65.10	0.78	0.08
C2	119	1.0	3.31	119.1	2.98	128.9	96.1	59.2	0.12	52.60	1.30	0.11
C3	119	1.5	3.31	119.1	2.78	125.9	95.5	64.5	0.12	44.70	1.55	0.10
C4	75	0.5	3.31	119.1	2.73	115.0	96.5	47.3	0.10	60.60	0.90	0.10
C5	73	1.0	3.31	119.1	3.01	125.7	94.2	61.3	0.17	48.70	1.29	0.07
C6	74	1.5	3.31	119.1	2.91	128.2	97.4	65.2	0.08	44.60	1.82	0.05

As shown in Table 4, the gold responded much better to cyanidation than the silver in these exploratory baseline tests, averaging 95.6% Au versus 56.6% Ag.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  13

In general, residue gold and silver grades decreased with finer grinding and higher cyanide concentration. In the range of 0.5-1.5 g/L cyanide, higher cyanide strength benefited silver extraction, but not gold.

Cyanide consumption was in the neighborhood of 1.3 kg/t at a level of 1 g/L NaCN. In general, less than 0.1 kg/t lime were required to maintain slurry pH >10. Increasing the cyanide strength from 0.5 g/L to 1.5 g/L doubled the NaCN consumption.

Leach kinetics showed that both gold and silver dissolution was fast in the first 8 hours of retention and subsequently slowed down. Finer grind and higher cyanidation strength resulted in faster metal dissolution and better extraction. Gold and silver extractions might improve by extending the residence time to over 72 hours without replacing leach liquor.

Figure 1. Gold Leach Kinetics on Milled NIVEL 4 Comp.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  14

Figure 2. Silver Leach Kinetics on Milled NIVEL 4 Comp.

4.2.2Leach Optimization

The optimization tests (C7 to C12) were conducted on the NIVEL 4 Comp. at a moderate grind of 100 mm at various cyanide dosages with leach solution replacement at various leach intervals. Test conditions and leach results are summarized in Table 5, while leach profiles are plotted in Figures 3 and 4.

In comparison to the baseline test results, leaching with solution replacement resulted in better silver extraction, and faster gold and silver leach kinetics. By replacing the leach solution, up to 70% silver recovery is achievable, and residence time can be shortened to 30 hours for gold. Double the number of solution exchanges and NaCN dosages as well as the addition of oxidants resulted in negligible gains in gold and silver extractions. In general, approximately 96% Au and 70% Ag extraction can be achieved at 40% solids in 1.0g/L NaCN solution with single solution replacement at the target grind of 100 µm.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  15

Table 5. Leach Optimization Test Summary

Test No		Test Conditions		Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
	NaCN, g/L	Oxidant	Sol'n Replacement	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C7	1.0		@ 24 hrs	3.31	119.1	2.58	131.8	95.6	67.6	0.11	42.7	1.23	0.17
C8	1.0	0.25kg/t Pb(NO3)2	@ 24 hrs	3.31	119.1	2.63	126.9	95.8	70.3	0.11	37.7	1.31	0.16
C9	1.0	0.1kg/t H2O2	@ 24 hrs	3.31	119.1	2.80	133.7	95.5	67.9	0.13	42.9	1.31	0.15
C10	1.5		@ 24 hrs	3.31	119.1	2.87	135.7	96.0	70.2	0.11	40.4	1.24	0.16
C11	1.5		@ 14- & 24 hrs	3.31	119.1	3.00	125.9	96.3	67.0	0.11	41.6	1.97	0.12
C12	2.0		@ 14- & 24 hrs	3.31	119.1	2.85	127.6	96.5	69.2	0.10	39.3	2.68	0.16

Figure 3. Optimization Gold Extraction Kinetics

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  16

Figure 4. Optimization Silver Extraction Kinetics

4.2.3Confirmatory Leach on New Santa Elena Feed

Confirmatory leach tests (C15 to C17) were carried out on another Santa Elena NIVEL composite (Comp. 4 ± 0 Nivel) with mine water (also referred as site water in the test spreadsheets) provided by the client. Leach pulp density increased from 40 wt.% for the baseline tests to 45%. During the leach tests, the pregnant leach solution was replaced with fresh mine water after 10 hours leach. Leach conditions and results are summarized in Table 6.

Results showed that over 97% gold extraction is achievable by leaching the Comp. 4 ± 0 Nivel in 1.0g/L cyanide at 45% solids by replacing partial or full leach liquor. Increasing cyanide dosage to 1.5g/L did not result in significant improvement of gold recovery. Silver extraction was ~43% in all three cases.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  17

Table 6. Confirmatory Test on Comp. 4 ± 0 Nivel

		Test Conditions		Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
Test No	NaCN, g/L	Oxidant	Sol'n Replacement	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C7	1.0		@ 24 hrs	3.31	119.1	2.58	131.8	95.6	67.6	0.11	42.7	1.23	0.17
C8	1.0	0.25kg/t Pb(NO3)2	@ 24 hrs	3.31	119.1	2.63	126.9	95.8	70.3	0.11	37.7	1.31	0.16
C9	1.0	0.1 kg/t H2O2	@ 24 hrs	3.31	119.1	2.80	133.7	95.5	67.9	0.13	42.9	1.31	0.15
C10	1.5		@ 24 hrs	3.31	119.1	2.87	135.7	96.0	70.2	0.11	40.4	1.24	0.16
C11	1.5		@ 14- & 24 hrs	3.31	119.1	3.00	125.9	96.3	67.0	0.11	41.6	1.97	0.12
C12	2.0		@ 14- & 24 hrs	3.31	119.1	2.85	127.6	96.5	69.2	0.10	39.3	2.68	0.16

Figure 5. Confirmatory Gold Extraction Kinetics

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  18

 

Figure 6. Confirmatory Silver Extraction Kinetics

4.2.4Leach Tailing Production

All remaining Santa Elena feed samples, including 20kg of Comp. Nivel 4 and 5.5kg of Comp. 4 ± 0 Nivel, were leached to produce sufficient tailing samples for thickening studies at Pocock Industrial. The two feed samples were processed separately due to their difference in color. The leach tests (C18-C19) were carried out at 48% solids with synthetic mine water in 1.5g/L NaCN at a target P80 grind of 100 mm. The leach results are summarized in Table 7.

Table 7. Bulk Leach Summary

Test No	Sample ID		Test Conditions			Extraction		Residue Grade		Consumption (kg/t)
		P80 Size, µm	% Solids	NaCN, g/L	Retention	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C18	Comp. 4 ± 0 Nivel	97	88	1.5	20 hours	94.4	42.4	0.06	21.2	0.52	0.13
C19	Nivel 4 Comp.	111	44	1.5	8 hours	84.9	51.8	0.29	54.9	0.35	0.12

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  19

4.2.5Mineralogical Investigation of Leach Residue

The Santa Elena leach residue (from test C10) was investigated by ore microscope and Scanning Electron Microscopy (SEM) in conjunction with X-Ray Diffraction (XRD) Analysis to determine silver deportment and the nature of silver causing the lower silver extraction in cyanidation.

Once again, the mineralogical investigation was performed by Dr. Lehne at Lehne & Associates Applied Mineralogy in Germany. According to Dr. Lehne, the Santa Elena leach residue consisted mainly of silica and carbonate gangue (quartz and calcite). Specific silver minerals have not been detected. Microanalysis proved a faint and erratic distribution of silver that could not be attributed to specific silver minerals or other silver carriers. "It is assumed that the silver retained in the sample originates from reprecipitation of silver during the leaching process”. (Dr. Lehne, Nov.22, 2011). Reprecipitation can be one of the main reasons for poor silver recovery in leaching operations. Detailed mineralogical findings are presened in Appendix IV.

4.3CYANIDATION OF CRUZ DE MAYO SAMPLE AND A HI-MN SAMPLE

Following the optimum leach conditions established on the Santa Elena samples, confirmatory bottle roll leach tests (C13-C14) were conducted on the composite sample from the Cruz de Mayo deposit (also referred as CM Comp. in the test spreadsheets) and the high manganese composite (Hi-Mn Comp.). Cyanide strength increased to 3g/L to prevent silver reprecipitation. The test conditions and results are summarized in Table 8, and gold and silver leach profiles are plotted in Figures 7 and 8, respectively.

As shown in Table 8, the Cruz de Mayo composite responded poorly to the cyanidation process with 51% Au and 54% Ag recoveries. A reasonable Au extraction of 92% was achieved on the Hi-Mn sample, but not Ag.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

 

Santa Elena Gold-Silver Project  20

Table 8. Cyanidation of CM Comp. and Hi-Mn Comp.

Test No	Sample ID	Test Conditions					Measured Head		Calculated Head		72-h Extraction		Residue Grade		Consumption (kg/t)
		P80 Size, µm	% Solids	NaCN, g/L	Retention	Sol'n Exch.	Au (g/t)	Ag (g/t)	Au (g/t)	Ag (g/t)	Au (%)	Ag (%)	Au (g/t)	Ag (g/t)	NaCN	Lime
C13	CM Comp.	101	40	3.0	72 hours	@ 10hrs	0.50	153.4	0.55	181.6	51.2	54.1	0.27	83.4	2.76	0.30
C14	Hi-Mn Comp.	103	40	3.0	72 hours	@10 hrs	5.04	93.4	5.07	119.9	92.1	60.2	0.40	47.7	2.11	0.19

Figure 7. CM and Hi-Mn Gold Extraction Kinetics

Figure 8. CM and Hi-Mn Silver Extraction Kinetics

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  21

According to mineralogical examination, the silver mineral (native silver and/or electrum, argentite and pyrargyrite-proustite) embedded in silica gangue might have resulted in the underprivileged silver recoveries on the Cruz de Mayo composite and the Hi-Mn composite.

In comparison to the tests done on the Santa Elena samples, cyanide consumption was higher in tests C13 and 14 due to the higher cyanide start level.

4.4SUPPORT STUDIES FOR TAILINGS EVALUATION

4.4.1Cyanide Detoxification

The tailing pulp from leach test C11 was used for scoping detoxification studies. The INCO air/SO2 cyanide destruction was simulated in a batch mode and an overall target of <1mg/L total CN in the effluent was arbitrarily set. Standard Free-CN titrations were performed to follow progress at regular intervals. Detailed analysis of the final products suggests that <1mg/L of total CN in the effluent is achievable.

4.4.2Characterization of Detoxified Sample

Full suites of environmental analyses were done on the detoxified effluent. Assay results as summarized in Table 9 indicate that cyanide is efficiently reduced to <0.5 mg/L. Table 9 also compares the characteristics of the detoxified feed and effluent. Further test details are provided in Appendix III.

Table 9. Solution Assay Summary

Sample	Selected Solution Analyses, mg/L
ID	TCN	CNS	CNO	As	Ag	Cu	Fe	Zn
C11 PLS	620	600	11.4	<0.2	5.8	1.46	8.43	2.54
CD1 Effluent	0.37	0.35	381.1	<0.2	0.59	0.88	0.17	0.45

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  22

4.4.3Settling Test

Settling tests were conducted on Santa Elena leach pulps to generate plant design data. The tests were carried out on leach tests C12 and C17 pulps with additions of 30g/t and 40g/t Percol Magnafloc 368, respectively, on diluted pulp based on flocculant screening tests. Settling test results on final leach residues are summarized in Table 10. Further test details are provided in Appendix III.

Results indicated that settling of Santa Elena leach tailing with 30 g/t of Percol Magnafloc 368 yielded a 24-hour raked underflow density of 74.8% solids and a unit thickener area of 0.1 m2/h/tpd of dry solids. An underflow density of 79.2% solids and a unit thickener area of 0.07 m2/h/tpd of dry solids are achievable by increasing the Percol Magnafloc 368 to 40g/t.

Table 10. Settling Tests on Leach Tailings

Test ID	Sample ID	Flocculant Dosage	Initial Density	24-h Max. u/f Density	Unit Thickener Area	Initial Rate
		g/tonne	% solids	% solids	m2/tpd	m/h
ST-1	Test C12 Leach Tailing	Percol Magnafloc 368, 30	30.0	74.8	0.10	1.6
ST-2	Test C17 Leach Tailing	Percol Magnafloc 368, 40	20.0	79.2	0.07	2.5

4.4.4 Vacuum Filtration Test

Underflows from the settling tests were adjusted to different pulp densities varying from 50% to 70% solids for vacuum filtration studies. The standard 0.1 ft2 laboratory vacuum disc filtration procedure was applied. Polypropylene type Filter cloth POPR 901F and P0-808HF were tested. The filtration test results are presented in Table 11 below with detailed test results provided in Appendix III.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  23

As can be seen in Table 11, no cake or thin cake formed during the filtration tests, and no clear filtrate was achieved in all tests. In most cases, the filter cloth was blinded by fine particles and no filtrate flowed after 5 seconds.

Table 11. Vacuum Filtration Results

Test ID	Sample ID	Pulp Density %	Cloth Type	Filtrate Capacity L/m2/h	Filtrate Clarity*	Solids Capacity kg/m2/h	Solids Moisture %	Note
VF-1	ST-1 Thickened tailing	40	Envirotech POPR-901F	134	3-4	10.3	29.3	No cake formed
VF-2	ST-1 Thickened tailing	40	Envirotech (PO- 808HF)	413	5	34.8	31.0	Thin cake formed
VF-3	ST-2 Thickened tailing	50	Envirotech POPR-901 F	76	2-3	6.2	23.8	No cake formed
VF-4	ST-2 Thickened tailing	60	Envirotech POPR-901 F	51	2-3	20.8	25.3	No cake formed
VF-5	ST-2 Thickened tailing	70	Envirotech POPR-901 F	64	2-3	32.6	14.9	Thin cake formed
VF-6	ST-2 Thickened tailing	50	Envirotech (PO- 808HF)	115	5	5.9	30.3	no cake formed
* Filtrate Clarity Scale 0 Crystal Clear, zero suspended solids 2Somewhat transparent solution 3Less cloudy, non-transparent solution 4Very cloudy discernible solid/liquid interface 5Opaque, no solid/liquid interface visible

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

Santa Elena Gold-Silver Project  24

5CONCLUSIONS AND RECOMMENDATIONS

Beneficiation of the Santa Elena samples by direct cyanide leaching alone will recover most of the gold but leave much of the silver in the residues. Grinding to 74 mm and leaching for 72-hours at 40 wt.% solids in 1.5 g/L NaCN recovered 97% Au and up to 65% Ag, with residue grades of 0.08 g/t Au and 44 g/t Ag. The silver extraction can be improved to 70% with single solution replacement after 10 hours leach. A 48-hour residence is adequate in leaching the Santa Elena ore with solution exchange. The use of lead nitrite and hydrogen peroxide, higher cyanide concentration or finer grinding had limited benefits on leaching. More detailed testing will be needed to optimize silver recovery.

The Cruz de Mayo sample responded poorly to the cyanidation process with 51% Au and 54% Ag recoveries. The gold in the Hi-Mn sample was more favorable to cyanidation than the silver.

Preliminary support studies on the leach pulp also showed encouraging response to detoxification, but poor flocculation, settling and filtration characterizations.

Inspectorate Exploration & Mining Services Ltd. – Metallurgical Division

APPENDIX 1

SAMPLE RECEIVING

SAMPLE RECEIVING LOG SHEET

Receiving Date: 26-Jul-11	Project No: 1103507
Carrier: DHL	Client: SilverCrest - Santa Elena project
Receiver: Victor	Page: 1 of 1

Count	Sample Label	Container Type	Sample Type (C, R, P, Sl, S)	Wet /Dry	Top Size	Weight (kg)
1	CM2-02	Pail	R	Dry	3"	22.00
2	CM2-07	Pail	R	Dry	3"	17.40
3	CM2-15	Pail	R	Dry	3"	16.90
4	CM2-24	Pail	R	Dry	3"	17.60
5	CM2.5-61	Pail	R	Dry	3"	17.50
6	CM2.5-66	Pail	R	Dry	3"	20.00
7	CM2.5-70	Pail	R	Dry	3"	18.20
8	CM2.5-88	Pail	R	Dry	3"	19.60
9	4NIVEL 622928	Pail	R	Dry	3"	3.80
10	COMPOSITO-NIVEL 4	Pail	R	Dry	6M	15.70
11	COMPOSITO-NIVEL 4	Pail	R	Dry	6M	14.10
12	COMPOSITO-NIVEL 4*	Pail	R	Dry	6M	30.00
13	VOL150 HIGH Mn	Pail	R	Dry	6M	28.00
14	* received on Aug.3, 2011
15
16
17
18
19
20
Note :						240.80
Core, Rock, Pulp, Slurry, Solution

SAMPLE RECEIVING LOG SHEET

Receiving Date: 6-Jun-12	Project No: 1103507
Carrier:	Client: SilverCrest - Santa Elena project
Receiver: Ed	Page: 1 of 1

Count	Sample Label	Container Type	Sample Type (C, R, P, Sl, S)	Wet /Dry	Top Size	Weight (kg)
1	Composito 4 + 0 Nivel - 4"#	Plastic Bag	R	Dry	6"	20.66
2	Composito 4 + 0 Nivel - 4"#	Plastic Bag	R	Dry	6"	29.88
3
4
5
6
7
8
9
10
11
12
13
14
15
16
17
18
19
20
Note :
-received in 2 orange plastic boxes
Core, Rock, Pulp, Slurry, Solution

APPENDIX 2

HEAD CHARACTERIZATION

HEAD ASSAY REPORT

Client: SilverCrest Mines Inc.	Date: 14-Jun-12
Sample: as specified	Project: 1103507

Item	Units	Sample ID Comp. Nivel 4	Analytical Method
Au	ppm	3.306	Au-1AT-AA
Ag	ppm	119.1	Ag-4A-OR
As	%	<0.01	As-1A-OR
S	%	0.03	S-LECO
Sulfide	%	0.02	S-SD-OR
C	%	1.88	C-LECO
C-Org	%	0.11	C-OG-OR
Ag	ppm	>100.0	30-4A-TR
Al	%	1.7	30-4A-TR
As	ppm	13	30-4A-TR
Ba	ppm	449	30-4A-TR
Bi	ppm	5	30-4A-TR
Ca	%	7.32	30-4A-TR
Cd	ppm	1.5	30-4A-TR
Co	ppm	3	30-4A-TR
Cr	ppm	247	30-4A-TR
Cu	ppm	85	30-4A-TR
Fe	%	0.98	30-4A-TR
K	%	1.67	30-4A-TR
La	ppm	<10	30-4A-TR
Mg	%	0.12	30-4A-TR
Mn	ppm	2107	30-4A-TR
Mo	ppm	7	30-4A-TR
Na	%	0.1	30-4A-TR
Ni	ppm	5	30-4A-TR
P	ppm	<10	30-4A-TR
Pb	ppm	183	30-4A-TR
Sb	ppm	49	30-4A-TR
Sc	ppm	2	30-4A-TR
Sr	ppm	158	30-4A-TR
Ti	%	0.04	30-4A-TR
Tl	ppm	<10	30-4A-TR
V	ppm	17	30-4A-TR
W	ppm	12	30-4A-TR
Zn	ppm	273	30-4A-TR
Zr	ppm	18	30-4A-TR

HEAD ASSAY REPORT

Client: SilverCrest Mines Inc.	Date: 14-Jun-12
Sample: as specified	Project: 1103507

Item	Units	Sample ID		Analytical Method
		Composito 4 ± 0 Nivel	Composito 4 ±  0 Nivel, DUP
Au	g/mt	1.36	1.55	FA
As	ppm	29	33	1A ICP
S(tot)	%	0.01	0.02	LECO
S(-2)	%	<0.01	<0.01	AsyWet
C(tot)	%	0.06	0.05	LECO
C(Org)	%	0.10	0.10	LECO
Ag	ppm	38.7	38.3	ICPM
Al	%	3.49	3.53	ICPM
As	ppm	67	42	ICPM
Ba	ppm	532	543	ICPM
Bi	ppm	<2	<2	ICPM
Ca	%	1.68	1.67	ICPM
Cd	ppm	1.0	1.0	ICPM
Co	ppm	6	6	ICPM
Cr	ppm	148	158	ICPM
Cu	ppm	70	56	ICPM
Fe	%	1.84	1.81	ICPM
K	%	3.03	3.03	ICPM
La	ppm	<10	<10	ICPM
Mg	%	0.29	0.30	ICPM
Mn	ppm	1684	1711	ICPM
Mo	ppm	3	2	ICPM
Na	%	0.04	0.04	ICPM
Ni	ppm	48	48	ICPM
P	ppm	287	321	ICPM
Pb	ppm	98	100	ICPM
Sb	ppm	34	32	ICPM
Sc	ppm	6	7	ICPM
Sr	ppm	57	57	ICPM
Ti	%	0.11	0.12	ICPM
Tl	ppm	<10	<10	ICPM
V	ppm	53	53	ICPM
W	ppm	49	47	ICPM
Zn	ppm	314	267	ICPM
Zr	ppm	28	31	ICPM

HEAD ASSAY REPORT

Client: SilverCrest Mines Inc.	Date: 14-Jun-12
Sample: as specified	Project: 1103507

Item	Units	Sample ID		Analytical Method
		Comp. CM	Comp. Hi-Mn
Au	ppm	0.504	5.043	Au-1AT-AA
Ag	ppm	148.5	89.5	Ag-4A-OR
Ag	ppm	158.28	97.38	50-4A-UT
Al	%	2.11	0.81	50-4A-UT
As	ppm	55.2	46.7	50-4A-UT
Ba	ppm	243	5793	50-4A-UT
Be	ppm	6.37	78.46	50-4A-UT
Bi	ppm	0.16	0.3	50-4A-UT
Ca	%	0.75	8.15	50-4A-UT
Cd	ppm	3.08	0.58	50-4A-UT
Ce	ppm	21.08	4.97	50-4A-UT
Co	ppm	2	3.4	50-4A-UT
Cr	ppm	212	202	50-4A-UT
Cs	ppm	6.85	3.7	50-4A-UT
Cu	ppm	221.9	83.5	50-4A-UT
Fe	%	0.77	0.59	50-4A-UT
Ga	ppm	9.45	9.32	50-4A-UT
Ge	ppm	1.42	1.31	50-4A-UT
Hf	ppm	4.6	<0.1	50-4A-UT
In	ppm	0.02	<0.01	50-4A-UT
K	%	1.97	0.34	50-4A-UT
La	ppm	10.2	2.8	50-4A-UT
Li	ppm	96.2	37.8	50-4A-UT
Mg	%	0.2	0.06	50-4A-UT
Mn	ppm	559	44246	50-4A-UT
Mo	ppm	20.13	2.18	50-4A-UT
Na	%	0.07	0.02	50-4A-UT
Nb	ppm	5.9	0.7	50-4A-UT
Ni	ppm	5.6	3.9	50-4A-UT
P	ppm	<10	<10	50-4A-UT
Pb	ppm	842.6	139.9	50-4A-UT
Rb	ppm	125.4	22.6	50-4A-UT
Re	ppm	<0.002	<0.002	50-4A-UT
S	%	0.049	0.148	50-4A-UT
Sb	ppm	387.06	265.07	50-4A-UT
Sc	ppm	1.5	0.8	50-4A-UT
Se	ppm	<1.0	<1.0	50-4A-UT
Sn	ppm	0.7	0.7	50-4A-UT
Sr	ppm	90	118.7	50-4A-UT
Ta	ppm	1.31	0.06	50-4A-UT
Te	ppm	<0.05	0.3	50-4A-UT
Th	ppm	6.1	0.6	50-4A-UT
Ti	%	0.045	0.012	50-4A-UT
Tl	ppm	<0.02	269.35	50-4A-UT
U	ppm	2.5	1.1	50-4A-UT
V	ppm	136	12	50-4A-UT
W	ppm	2	123.7	50-4A-UT
Y	ppm	5.9	4.8	50-4A-UT
Zn	ppm	825	142	50-4A-UT
Zr	ppm	19.2	4.3	50-4A-UT

ASSAY REPORT

Client: SilverCrest Mines Inc.	Date: 14-Jun-12
Sample: as specified	Project: 1103507

Item	Units	Sample ID		Analytical
		Site Water	Site Water dupl.	Method
pH	--	8.10	8.08	Meter
Ag	mg/L	<0.02	<0.02	ICPH2O
Al	mg/L	<0.2	<0.2	ICPH2O
As	mg/L	4.5	3.5	ICPH2O
Ba	mg/L	0.05	0.05	ICPH2O
Bi	mg/L	<0.1	<0.1	ICPH2O
Ca	mg/L	18.4	18.4	ICPH2O
Cd	mg/L	<0.01	<0.01	ICPH2O
Co	mg/L	0.01	0.01	ICPH2O
Cr	mg/L	<0.01	<0.01	ICPH2O
Cu	mg/L	0.11	0.10	ICPH2O
Fe	mg/L	3.14	2.43	ICPH2O
Hg	mg/L	<0.05	<0.05	ICPH2O
K	mg/L	<2	<2	ICPH2O
La	mg/L	<0.05	<0.05	ICPH2O
Mg	mg/L	2.5	2.5	ICPH2O
Mn	mg/L	<0.01	<0.01	ICPH2O
Mo	mg/L	<0.02	<0.02	ICPH2O
Na	mg/L	16	16	ICPH2O
Ni	mg/L	<0.02	<0.02	ICPH2O
P	mg/L	0.4	0.5	ICPH2O
Pb	mg/L	<0.05	<0.05	ICPH2O
Sb	mg/L	<0.1	<0.1	ICPH2O
Sc	mg/L	<0.01	<0.01	ICPH2O
Sr	mg/L	0.11	0.11	ICPH2O
Ti	mg/L	<0.1	<0.1	ICPH2O
Tl	mg/L	<0.2	<0.2	ICPH2O
V	mg/L	<0.01	<0.01	ICPH2O
W	mg/L	<0.1	0.1	ICPH2O
Zn	mg/L	1.07	0.89	ICPH2O
Zr	mg/L	<0.01	<0.01	ICPH2O

BOND MILL GRINDABILITY TEST REPORT

Client: SilverCrest - Santa Elena	Date: 26-Aug-11
Test: BI-1	Project: 1103507
Sample: NIVEL-4

TEST CONDITIONS

Cycle	Oversize Wt. grams	Product Wt. grams	Feed Undersize grams	Net Product grams	Product per Rev. grams/rev.	Required Rev. rev.
1	928.30	340.38	227.14	113.24	1.1324	100
2	892.70	375.98	60.94	315.04	1.1844	266
3	891.20	377.48	67.31	310.17	1.2456	249
4	892.30	376.38	67.58	308.80	1.3029	237
5	893.30	375.38	67.39	307.99	1.3628	226
6	901.80	366.88	67.21	299.67	1.3810	217
7	904.90	363.78	65.69	298.09	1.3865	215

SIZE ANALYSIS				TEST RESULTS
Sieve Size		% Passing		Material Charge Wt.-700 mL(g) = 1,268.
Tyler mesh	µm	Feed	Product	Test Screen (^m) = 105
8	2,380	80.8		Undersize in Feed (%) = 17.90
10	1,680	65.8		Circulating Load (%) = 249
14	1,190	54.9		Gbp (ave.) = 1.38
20	841	45.6		Product P80 (^m) = 86
28	595	39.2		Feed F80 (^m) = 2,340
35	420	33.4		W (kWh/ton) = 13.4
48	297	28.3		W (kWh/tonne) = 14.8
65	210	24.2
100	149	21.2
150	105	17.9	100.0
200	74	15.1	68.1
270	53	12.7	49.1
325	44	11.5	40.8
400	37	10.6	30.9

BOND MILL GRINDABILITY TEST REPORT

Client: SilverCrest - Santa Elena	Date: 6-Jan-12
Test: BI-2	Project: 1103507
Sample: CM Comp

TEST CONDITIONS

Cycle	Oversize Wt. grams	Product Wt. grams	Feed Undersize grams	Net Product grams	Product per Rev. grams/rev.	Required Rev. rev.
1	988.4	197.8	153.7	44.0	0.88	50
2	871.0	315.2	25.6	289.5	0.81	356
3	818.5	367.7	40.8	326.8	0.89	366
4	830.8	355.4	47.7	307.7	0.94	326
5	845.2	341.0	46.1	294.9	0.95	311
6	847.8	338.4	44.2	294.2	0.95	310

 SIZE ANALYSIS TEST RESULTS

Sieve Size		% Passing		Material Charge Wt.-700 mL(g) = 1,186.
Tyler mesh	µm	Feed	Product	Test Screen (^m) = 105
8	2,380	77.1		Undersize in Feed (%) = 12.96
10	1,680	61.6		Circulating Load (%) = 250
14	1,190	50.6		Gbp (ave.) = 0.95
20	841	40.9		Product P80 (^m) = 81
28	595	33.5		Feed F80 (^m) = 2,498
35	420	27.6		W (kWh/ton) = 17.6
48	297	22.7		W (kWh/tonne) = 19.3
65	210	18.7
100	149	15.5
150	105	13.0	100.0
200	74	10.6	73.7
270	53	8.9	57.6
325	44	8.1	49.6
400	37	7.3	42.8

Hazen Research, Inc.

4601 Indiana Street • Golden, Colorado 80403 USA

Phone: (303) 279-4501 • Fax: (303) 278-1528

www. hazenresearch. com

November 9, 2011

E-mail and Mail Delivery

Ms. Alice Shi

Inspectorate America Corporation

11620 Horseshoe Way

Richmond, BC V7A 4V5

Canada

E-mail:alice.shi@inspectorate. com

Subject:      Comminution Testing

Hazen Project 010-271

Client Project 1103507

Dear Ms. Shi:

As requested, the sample received at Hazen Research, Inc. on October 14, 2011, was subjected to Bond crusher impact work index (CWi) and Bond abrasion index (Ai) testing. Upon receipt, the sample was assigned a Hazen number for internal identification and future reference. Table 1 summarizes the CWi and Ai results for this sample. The data sheets are enclosed.

Table 1. Sample Identification and Test Results

HRI	Client ID	CWi, kWh/t	Ai, g
52901	1103507 CM Composite	13.59	1.1267

This letter report completes the work authorized for this sample. Hazen looks forward to assisting you with comminution testing in the future. Please contact me if you have questions concerning this project.

Regards,

Daniel W. Gillespie

Project Engineer

DWG/gmr

Enclosures

Abrasion Test	Date:	Oct-11
	Project:	010-271
	Sample:	08753

Purpose:	To determine the abrasion index which can be used to determine steel
	media and liner wear in crushers, rod mills, and ball mills.
Procedure:	The equipment and procedure duplicate the Pennsylvania Crusher
	method for determining an abrasion index.
Sample:	Client Identified:	CM Composite
	Sample Number:	52901

Results:	Original Coupon Weight	=	94.7168 g
	Final Coupon Weight	=	93.5901 g
	Abrasion Index (Ai)	=	1.1267 g

Equipment		Equationsa (Ai > 0.021)	Wear Rate (lb/kWh)
Wet Rod Mill	Rods =	0.35(Ai - 0.020)0.2	= 0.3572
	Liner =	0.035(Ai - 0.015)0.3	= 0.0361
Wet Ball Mill	Balls =	0.35(Ai - 0.015)0.33	= 0.3624
(overflow and grate discharge) Liner	Liner =	0.026(Ai - 0.015)0.3	= 0.0268
Dry Ball Mill	Balls =	0.05Ai0.5	= 0.0531
(grate discharge, Ai < 0.22)	Liner =	0.005Ai0.5	= 0.0053
Crushers	Liner =	(Ai + 0.22)/11	= 0.1224
(gyratory, jaw, and cone)
Roller Crushers	Roll Shell =	(0.1Ai)0.67	= 0.2331

aBond, FC, "Metal Wear in Crushing and Grinding," Allis-Chalmers Publication 07P1701, Dec. 1963.

Hazen Research, Inc. • 1 of 2

Abrasion Test	Date:	Oct-11
	Project:	010-271
	Sample:	08753

Tyler Mesh			Direct Wt.				Cumulative %
i. or mesh	pm			g	%		Passing		Retained
3/4 in.		19,000		0.0		0.0		100.0		0.0
1/2		12,700		234.0		14.6		85.4		14.6
3/8		9,510		398.2		24.8		60.6		39.4
3 mesh		6,700		174.2		10.9		49.7		50.3
4		4,760		76.8		4.8		44.9		55.1
6		3,360		62.4		3.9		41.0		59.0
8		2,380		43.2		2.7		38.3		61.7
10		1,680		39.5		2.5		35.9		64.1
14		1,190		41.4		2.6		33.3		66.7
20		841		33.3		2.1		31.2		68.8
28		595		41.7		2.6		28.6		71.4
35		420		43.7		2.7		25.9		74.1
48		297		53.3		3.3		22.6		77.4
65		210		57.3		3.6		19.0		81.0
100		149		48.7		3.0		15.9		84.1
150		105		56.2		3.5		12.4		87.6
200		74		43.5		2.7		9.7		90.3
-200		-74		156.0		9.7		0.0		100.0
		Total		1,603.4		100.0

Hazen Research, Inc. • 2 of 2

	Laboratory and Shipping:
Mailing Address: 2501	5946 McIntyre Street
Braun Dr. Golden, CO	Golden, CO 80403
80401 (303) 279-0443	Phone: (303) 854-2037
	Fax: (303) 216-0258
	E mail: phillips81@att.net

November 7, 2011

Mr. Dan Gillespie

Hazen Research Inc.

4601 Indiana St.

Golden, CO 80403

Re: Bond Low-Energy Impact Tests PE

Project No: 114010 Client project

No. 010-271 Client Purchase

Order: 99207

Dear Mr. Gillespie:

Phillips Enterprises, LLC has completed 1 low-energy Bond impact test on your ore sample. The following table summarizes the test results. Results are reported in metric units.

Test No.	Sample Designation	Work Index (CWi)
1	HRI 52901	13.59

Comments:

The tests were performed by the standard procedure. There were no abnormal features of the natural rock sample.

A photograph of the sample is available upon request. Please contact Phillips Enterprises if you have any questions. We look forward to being of service to Hazen Research in the future.

Sincerely,

Robert J. Phillips Owner/Manager

PHILLIPS ENTERPRISES LLC  BOND IMPACT TEST

Test #	1	Project No.	114010
		Date:	Nov 2, 2011

Purpose:	To determine the low-energy Work Index for crushing as a basis for determining crushing power requirements.
Sample:	20 pieces of sample "HRI 52901”
Procedure:	Bond procedure for determining work index using a twin-pendulum impact testing machine using natural 75 x 50 mm pieces.

Piece	Sample Thickness			Impact Energy for Breakage
	64ths of inch	Inches	Millimeters	Ft-lb	Joules	Joules/mm
1	121	1.89	48.02	16	21.6936	0.452
2	112	1.75	44.45	28	37.9638	0.854
3	121	1.89	48.02	16	21.6936	0.452
4	112	1.75	44.45	16	21.6936	0.488
5	102	1.59	40.48	24	32.5404	0.804
6	126	1.97	50.01	16	21.6936	0.434
7	151	2.36	59.93	36	48.8106	0.814
8	132	2.06	52.39	44	59.6574	1.139
9	169	2.64	67.07	36	48.8106	0.728
10	132	2.06	52.39	24	32.5404	0.621
11	116	1.81	46.04	8	10.8468	0.236
12	146	2.28	57.94	8	10.8468	0.187
13	150	2.34	59.53	28	37.9638	0.638
14	141	2.20	55.96	28	37.9638	0.678
15	106	1.66	42.07	44	59.6574	1.418
16	117	1.83	46.43	16	21.6936	0.467
17	142	2.22	56.36	16	21.6936	0.385
18	136	2.13	53.98	24	32.5404	0.603
19	108	1.69	42.86	20	27.117	0.633
20	116	1.81	46.04	16	21.6936	0.471
				Average:		0.625
				Specific Gravity:		2.46
(1) Formula =	53.49 X Joules/mm		Work Index(1), kW-hr/m.ton:			13.59
	Specific Gravity		Work Index, kW-hr/short ton:			12.33

Hazen Research, Inc.

4601 Indiana Street • Golden, Colorado 80403 USA

Phone: (303) 279-4501 • Fax: (303) 278-1528

www. hazenresearch. com

July 13, 2012

E-mail and Mail Delivery

Ms. Alice Shi

Inspectorate Exploration & Mining Services Ltd.

Metallurgical Division 11620 Horseshoe Way Richmond, BC V7A 4V5 Canada

E-mail: alice.shi@inspectorate. com

Subject: Comminution Testing Hazen Project 010-325 Report and Appendices Client Project 1103507

Dear Ms. Shi:

As requested, the sample received at Hazen Research, Inc. on June 18, 2012, was subjected to Bond ball mill work index (BWi) and Bond rod mill work index (RWj) testing. Upon receipt, the sample identified as 1103507 Composito 4±0 Nivel was assigned Hazen number 53188 for internal identification and future reference. Table 1 summarizes the BWi and RWi results for this sample. All test data sheets are included in the appendices.

Table 1. Sample Identification, BWi and RWi Results

HRI	Client ID	BWi, kWh/t	RWi, kWh/t
53188	1103507 4±0 Nivel	14.9	14.0

This letter report completes the work authorized for this sample. Hazen looks forward to assisting you with comminution testing in the future. Please contact me if you have questions concerning this project.

Regards,

Daniel W. Gillespie

Project Engineer

DWG/lmr

APPENDIX A

BWi Results

  Hazen Research, Inc.

tandard Bond Ball Mill Grindability Test Date: Jun 2012

Project: 010-325

Sample: 53188

Purpose: To determine the ball mill grindability of the test sample in terms of a Bond ball mill

work index, BWi.

Procedure: The equipment and procedure duplicate the Bond method for determining ball mill

work indices. The sample was stage-crushed to minus 6 mesh. This material

was used as feed for the grindability test.

Sample: Client Identified: 1103507 Composito 4±0 Nivel

Sample No: 53188

Results: P1 = 100% passing size of product 149 μm

P80 = 80% passing size of product 121 μm

F80 = 80% passing size of feed 2,590 μm

Gpr = Grams per revolution 1.58 g

BWi = 13.6 kWh/st

BWi = 14.9 kWh/t

Standard Bond Ball Mill Grindability Test Date: Jun 2012

Project: 010-325

Sample: 53188

Test Conditions and Notes Test Ball Charge

100% passing size of product: 100 mesh Ball Size, Number Weight,

Test feed weight (700 cm3): 1,160.3 g in. of Balls g

Amount of minus 100 mesh in feed: 12.2 % 1.45 43 8,806.4

Target product weights: + 100 mesh 828.8 g 1.17 67 7,209.2

- 100 mesh 331.5 g 1.00 10 672.0

0.75 71 2,009.3

0.61 94 1,428.0

Total: 285 20,125.0

To Be U'Size In Per Mill Circ.

New Feed, In Feed, Ground, Product, Total, Revolution, Load,

Stage No. Revolutions g g g g g g %

1 100 1,160.3 142.1 189.4 283.4 141.3 1.41 309.4

2 210 283.4 34.7 296.8 323.2 288.5 1.37 259.0

3 213 323.2 39.6 291.9 344.8 305.2 1.43 236.5

4 202 344.8 42.2 289.3 344.5 302.3 1.50 236.8

5 193 344.5 42.2 289.3 338.9 296.7 1.54 242.4

6 189 338.9 41.5 290.0 336.3 294.8 1.56 245.0

7 186 336.3 41.2 290.3 331.9 290.7 1.56 249.6

8 186 331.9 40.6 290.9 333.3 292.7 1.57 248.1

9 185 333.3 40.8 290.7 333.3 292.5 1.58 248.1

10 184 333.3 40.8 290.7 331.7 290.9 1.58 249.8

Average of Last Three Stages: 1.58 248.7

Undersize Undersize Product

Hazen Research, Inc. • 2 of 4

Standard Bond Ball Mill Grindability Test Date: Jun 2012

Project: 010-325

Sample: 53188

Test Conditions and Notes Test Ball Charge

100% passing size of product: 100 mesh Ball Size, Number Weight,

Test feed weight (700 cm3): 1,160.3 g in. of Balls g

Amount of minus 100 mesh in feed: 12.2 % 1.45 43 8,806.4

Target product weights: + 100 mesh 828.8 g 1.17 67 7,209.2

- 100 mesh 331.5 g 1.00 10 672.0

0.75 71 2,009.3

0.61 94 1,428.0

Total: 285 20,125.0

To Be U'Size In Per Mill Circ.

New Feed, In Feed, Ground, Product, Total, Revolution, Load,

Stage No. Revolutions g g g g g g %

1 100 1,160.3 142.1 189.4 283.4 141.3 1.41 309.4

2 210 283.4 34.7 296.8 323.2 288.5 1.37 259.0

3 213 323.2 39.6 291.9 344.8 305.2 1.43 236.5

4 202 344.8 42.2 289.3 344.5 302.3 1.50 236.8

5 193 344.5 42.2 289.3 338.9 296.7 1.54 242.4

6 189 338.9 41.5 290.0 336.3 294.8 1.56 245.0

7 186 336.3 41.2 290.3 331.9 290.7 1.56 249.6

8 186 331.9 40.6 290.9 333.3 292.7 1.57 248.1

9 185 333.3 40.8 290.7 333.3 292.5 1.58 248.1

10 184 333.3 40.8 290.7 331.7 290.9 1.58 249.8

Average of Last Three Stages: 1.58 248.7

Undersize Undersize Product

Hazen Research, Inc. • 2 of 4

Standard Bond Ball Mill Grindability Test Date: Jun 2012

Project: 010-325

Sample: 53188

0

10

20

30

40

50

60

70

80

90

100

10 100 1,000 10,000

Cumulative Passing, %

Particle Size, μm

appendix b

rwi results

Client: SilverCrest - Santa Elena       Date: 06/28/12 Test: ST-2

Project: 1103507 Sample: C17 slurry ( 20% solids with rake )    Page: 1 of  Time

(min.)  Height

(cm)    Sludge Density

(w/w % solids)0 39.7    19.8 7  10.2    56.7 8  9.4     59.7 10 8.2     64.9

13      8.1     65.4

18      7.8     66.8

30      7.4     68.9

60      6.9     71.6

129     6.5     74.0

312     6.2     75.9

461     6.0     77.2

1440    5.7     79.2 Slurry pH: 10.5 Coagulant:      n/a

Flocculant: Magnafloc 368, 40g/t Dry Solids Density:     2.68 g/cm3 Liquid Density: 1.00 g/cm3 Weight of Dry Solids:   452.0 g Initial Slurry Weight:  2264 g Initial Slurry Volume:  2000 mL Initial Slurry Height:  39.7 cm Initial Weight Percent Solids:  20.0 w/w % solids Initial Settling Rate:  2.5 m/h Final Sediment Volume:  290 mL Final Sediment Height:  5.7 cm Supernatant Clarity:    3.5

Floc Size:      1 Supernatant Clarity Scale       Floc Size Scale 0       Crystal Clear, zero suspended solids    1 Very fine particles 1       Transparent - some suspended solids     2 to 9 Floc size increasing  2       Somewhat transparent solution   10 Very large flocs 3       Less cloudy, non-transparent solution 4       Very cloudy discernible solid/liquid interface 5       Opaque, no solid/liquid interface visible

 Unit Thickener Area Determination Modified Coe and Clevenger Method/ Oltman Technique

 Required Underflow Pulp Density: Compression Point: Slope (Settling Rate), R Feed Dilution, F Underflow Dilution, Du Liquid Relative Density, L Unit Thickener Area, A 60 w/w % solids 8.8 cm 9.0     min 49.4 m/d 4.1     (weight solution/weight solids) 0.67 (weight solution/weight solids) 1.00

Client: SilverCrest - Santa Elena       Date: 06/28/12

Test: ST-2      Project: 1103507

Sample: C17 slurry ( 20% solids with rake )     Page: 1 of 2

SETTLING TEST REPORT

Client: SilverCrest - Santa Elena       Date: 28-Jun-12

Test: ST-2      Project: 1103507

Sample: C17 slurry ( 20% solids with rake )     Page: 2 of 2

Interface Height, H versus Time, t

Hazen Research Inc. ● 3 of 4 

Client: Pulp density: 40% Date: 2-Nov-11

Sample id: ST-1 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-1

Test description: Sample slurried to 40% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-1 27 27 60 102 - 0.5 4.41 3.13 29.02 - - - 55 3-4

VF-1(Rep.) 27 27 30 53 - 0.3 4.05 2.85 29.63 - - - 30 4

Note:

Filter cloth blinded, after 5seconds no flow (no cake formed)

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 1.5 kg/m2/h 7.4 Filtrate Clarity Scale

Dry lbs/sq.ft./h= 2.7 kg/m2/h 13.1 0 Crystal Clear, zero suspended solids

2 Somewhat transparent solution

Filtrate Capacity Determination 3 Less cloudy, non-transparent solution

4 Very cloudy discernible solid/liquid interface

VF-1: Gal/sq.ft./h= 3.2 L/m2/h 130 5 Opaque, no solid/liquid interface visible

Gal/sq.ft./h= 3.4 L/m2/h 138 VF-1Rep

VACUUM FILTRATION TEST REPORT

  Hazen Research Inc. ● 4 of 4 

VACUUM FILTRATION TEST REPORT
SilverCrest - Santa Elena
Client: Pulp density: 50% Date: 3-Jul-12

APPENDIX 3

METALLURGICAL TEST RESULTS

Client: Pulp density: 60% Date: 3-Jul-12 SilverCrest - Santa Elena VACUUM FILTRATION TEST REPORT

Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-4

Test description: Sample slurried to 60% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-4 27 27 60 90 - 0.5 10.93 8.17 25.25 - - - 20 2-3

Filter cloth blinded, after 5 seconds no flow (no cake formed)

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 4.3 kg/m2/h 20.8 Filtrate Clarity Scale

0 Crystal Clear, zero suspended solids

Filtrate Capacity Determination 2 Somewhat transparent solution

3 Less cloudy, non-transparent solution

VF-1: Gal/sq.ft./h= 1.3 L/m2/h 51 4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

Client: Pulp density: 70% Date: 3-Jul-12 SilverCrest - Santa Elena VACUUM FILTRATION TEST REPORT

Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-5

Test description: Sample slurried to 60% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-5 27 27 60 90 - 1 15.02 12.78 14.92 1.336 1.211 9.36 25 2-3

Thin cake formed but fine solids into filtrate

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 6.7 kg/m2/h 32.6 Filtrate Clarity Scale

0 Crystal Clear, zero suspended solids

Filtrate Capacity Determination 2 Somewhat transparent solution

3 Less cloudy, non-transparent solution

VF-1: Gal/sq.ft./h= 1.6 L/m2/h 64 4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

Client: Pulp density: 50% Date: 3-Jul-12

Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech (PO-808HF) Test series: VF-6

Test description: Sample slurried to 50% solids and submitted to leaf filtration test using PO-808HF filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-6 27 27 60 90 - 0.3 3.32 2.315 30.27 - - - 45 5

Filter cloth blinded, after 5 seconds no flow (no cake formed)

Cake Capacity Determination

Filtrate Clarity Scale

VF-2: Dry lbs/sq.ft./h= 1.2 kg/m2/h 5.9 0 Crystal Clear, zero suspended solids

2 Somewhat transparent solution

Filtrate Capacity Determination 3 Less cloudy, non-transparent solution

4 Very cloudy discernible solid/liquid interface

VF-2: Gal/sq.ft./h= 2.8 L/m2/h 115 5 Opaque, no solid/liquid interface visible SilverCrest - Santa Elena VACUUM FILTRATION TEST REPORT

APPENDIX 4

MINERALOGICAL REPORT

Microscopical Investigation of Two Auriferous Composite Samples from the Cruz de Mayo Project, Sonora, Mexico (INSPECTORATE - METALLURGICAL DIVISION) November 2011 (11-11/128)

Summary

The following report presents the results of ore microscopic investigations carried out on two head composite samples from the Mexican Cruz de Mayo project. The investigations focussed on gold and silver mineralogy.

The first head composite consists of silica gangue that possesses a pronounced iron staining due to finely disseminated limonite. Silver carriers are argentite as well as pyrargyrite-proustite. In addition to these there is also native silver and/or electrum. Trace of gold occur in native

The second head composite is characterised by a distinct manganese mineralisation consisting essentially of pyrolusite. Specific silver minerals have not been observed. Silver is obviously contained in electrum which occurs together with native gold in association with limonite.

The small particle sizes of the noble metal phases imply rather resistant locking textures. Therefore a leaching process for the recovery of gold and silver is suggested. It is assumed that previously intense grinding of the siliceous materials will be required.

Resumen

El informe presenta los resultados de un estudio mineragrafico realizado en dos muestras colectivas de mineral de cabeza del proyecto mexicano de Cruz de Mayo. El estudio se centra en la m ineralogfa de oro y plata.

Una muestra esta formada poor ganga silicea que posee una marcada coloracion ferruginosa causada por limonita diseminada. Los portadores de plata son la argentita y la proustita- pirargirita. Ademas hay plata nativa y/o electrum. Trazas de oro se presentan en forma nativa.

La segunda muestra esta caracterizada por una pronunciada mineralizacion de manganeso que consiste principalmente de pirolusita. No se ha observado minerales especfficos de plata. Sin embargo hay ele ctrum que se presenta junto con oro nativo en asociacion con la limonita.

El pequeno tamano de grano de los minerales de oro y plata que implica texturas resistentes a la molienda , sugiere un p roceso de lixiviacion para la recuperacion de los metales nobles. Se supone que una previa molienda intensa de lass menas siliceas sera necesaria.

www.isogyre.com

Mannheim,

R. W. Lehne

November 1 £3, 2011

R. W. Lehne

Client: SilverCrest Mines Inc. Date: 19-Aug-11

Test: C4 Project: 1103507

Sample: Cyanidation Residue

Grind: 2kg grind for 28' @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.1 99.9

100 149 0.3 99.7

150 105 5.3 94.4

200 74 14.7 79.7

270 53 16.7 63.1

325 44 8.4 54.7

400 37 6.7 48.0

Undersize - 37 48.0 -

TOTAL: 100.0

80 % Passing Size (μm) = 75

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest Mines Inc. Date: 19-Aug-11

Test: C5 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: 72 hours bottle roll cyanide leach in 1.0 g/L NaCN at a target grind of 80% passing 75 microns

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,000 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 1.0g/L NaCN

Grind Size - P80: 73 μm - sampled at 2,4,7,24,30,48, 54 hours

Initial NaCN: 1.0 g/L - test ended after 72 hours

Target pH: 10.0 - filtered and displacement washed with hot cyanide solution

Test Duration: 72 hours followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 3.01 g/t 125.7 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.00 3.00 0.14 8.2 9.7 5,000 3,000

2 0.84 0.48 10.3 4,998 2,998 30 0.75 2.2 20.48 61.4

4 0.96 0.12 10.3 8.3 5,000 3,000 30 1.14 3.4 25.89 78.3

7 0.96 0.12 10.3 5,000 3,000 30 1.34 4.1 29.00 88.4

24 0.84 0.48 10.4 8.1 4,996 2,996 30 1.50 4.6 39.60 120.9

30 0.90 0.30 10.4 5,002 3,002 30 1.50 4.6 41.50 128.0

48 0.86 0.42 10.4 8.2 4,996 2,996 30 1.48 4.6 45.30 140.4

54 0.92 0.24 11.3 5,002 3,002 30 1.53 4.8 46.20 144.7

72 0.86 10.4 8.0 5,000 3,000 1.80 5.7 48.90 154.1

Total 5.16 0.14

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,000 0.17 0.35 48.7 97.4

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 37.4 24.4 0.24

4 57.2 31.1 0.30

7 67.7 35.1 0.36

24 76.3 48.1 0.60

30 77.2 50.9 0.75

48 76.8 55.8 0.96

54 80.2 57.5

72 94.2 61.3 1.29 0.07

Residue 5.8 38.7

Total 100.0 100.0

1.08

50

CYANIDATION TEST REPORT

Solution

Ag

Total Residue

Distribution Reagent Consumption

Client: SilverCrest Mines Inc. Date: 19-Aug-11

Test: C5 Project: 1103507

Sample: Cyanidation Residue

Grind: 2kg grind for 28' @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.0 100.0

100 149 0.1 99.8

150 105 4.0 95.9

200 74 15.2 80.6

270 53 17.6 63.0

325 44 8.7 54.3

400 37 6.4 47.9

Undersize - 37 47.9 -

TOTAL: 100.0

80 % Passing Size (μm) = 73

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest Mines Inc. Date: 19-Aug-11

Test: C6 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: 72 hours bottle roll cyanide leach in 1.5 g/L NaCN at a target grind of 80% passing 75 microns

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,000 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 74 μm - sampled at 2,4,7,24,30,48, 54 hours

Initial NaCN: 1.5 g/L - test ended after 72 hours

Target pH: 10.0 - filtered and displacement washed with hot cyanide solution

Test Duration: 72 hours followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 2.91 g/t 128.2 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 4.50 0.10 8.3 9.2 5,000 3,000

2 1.28 0.66 10.3 4,998 2,998 30 0.93 2.8 25.36 76.0

4 1.45 0.15 10.4 8.2 5,000 3,000 30 1.30 3.9 29.80 90.1

7 1.40 0.30 10.4 5,000 3,000 30 1.46 4.4 34.80 106.0

24 1.28 0.66 10.4 8.0 4,994 2,994 30 1.62 5.0 46.10 140.7

30 1.38 0.36 10.4 5,000 3,000 30 1.60 5.0 48.80 150.5

48 1.30 0.60 10.4 8.2 4,992 2,992 30 1.63 5.1 51.80 160.5

54 1.40 0.30 10.5 4,998 2,998 30 1.63 5.1 52.30 163.9

72 1.30 10.5 8.0 4,994 2,994 1.79 5.7 53.00 167.3

Total 7.53 0.10

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,000 0.08 0.15 44.6 89.2

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 47.9 29.6 0.33

4 67.5 35.1 0.41

7 76.5 41.3 0.56

24 85.3 54.8 0.89

30 85.3 58.6 1.07

48 87.4 62.6 1.37

54 88.4 63.9

72 97.4 65.2 1.82 0.05

Residue 2.6 34.8

Total 100.0 100.0

1.52

60

CYANIDATION TEST REPORT

Solution

Ag

Total Residue

Distribution Reagent Consumption

Client: SilverCrest Mines Inc. Date: 19-Aug-11

Test: C6 Project: 1103507

Sample: Cyanidation Residue

Grind: 2kg grind for 28' @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.0 100.0

100 149 0.2 99.7

150 105 4.6 95.1

200 74 14.8 80.2

270 53 17.2 63.1

325 44 8.5 54.5

400 37 6.6 47.9

Undersize - 37 47.9 -

TOTAL: 100.0

80 % Passing Size (μm) = 74

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 12-Oct-11

Test: as specified Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: to improve gold and silver recoveries with solution replacement after 24 hours at different cyanide strengths at a target P80 grind size of 100 microns

P80 Size, um % Solids NaCN, g/L Retention

Au (g/t) Ag (g/t) Au (g/t) Ag (g/t) Au (%) Ag (%) Au (g/t) Ag (g/t) NaCN Lime

C7

NIVEL 4 -

Santa Elena

project

99 40 1.0 72 hours 3.31 119.1 2.58 131.8 95.6 67.6 0.11 42.7 1.23 0.17

C8

NIVEL 4 -

Santa Elena

project

104 40

0.25kg/t

Pb(NO3)2

1.0 72 hours 3.31 119.1 2.63 126.9 95.8 70.3 0.11 37.7 1.31 0.16

C9

NIVEL 4 -

Santa Elena

project

109 40

0.1kg/t

H2O2

1.0 72 hours 3.31 119.1 2.80 133.7 95.5 67.9 0.13 42.9 1.31 0.15

C10

NIVEL 4 -

Santa Elena

project

105 40 1.5 72 hours 3.31 119.1 2.87 135.7 96.0 70.2 0.11 40.4 1.24 0.16

CYANIDATION TEST SUMMARY

Consumption (kg/t)

Test No Sample ID

Measured Head Calculated Head 72-h Extraction Test Conditions Residue Grade

0

10

20

30

40

50

60

70

80

90

100

110

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Gold Extraction Kinetics

C7 C8 C9 C10

0

10

20

30

40

50

60

70

80

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Silver Extraction Kinetics

C7 C8 C9 C10

Client: SilverCrest -Santa Elena project Date: 12-Oct-11

Test: C7 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: to improve Au and Ag recoveries in 1g/L NaCN with solution replacement after 24 hours at a target P80 grind of 100 microns

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,002 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 1.0g/L NaCN

Grind Size - P80: 99 μm - sampled at 2,4,7,24,26,28,31,48,54 hours

Initial NaCN: 1.0 g/L - replaced solution with fresh water after 24 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 2.58 g/t 131.8 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.00 3.00 0.34 8.2 10.1 5,002 3,000

2 0.84 0.48 10.5 4,998 2,996 30 0.67 2.0 20.02 60.0

4 1.00 10.5 5,004 3,002 30 1.02 3.1 24.69 74.7

7 0.94 0.18 10.5 7.7 5,000 2,998 30 1.15 3.5 28.62 87.2

24 0.82 10.4 7.9 4,996 2,994 30 1.18 3.6 35.80 109.4

Replaced solution with fresh water

24 1.00 3.00 10.3 5,004 3,002

26 1.00 10.6 8.0 5,004 3,002 30 0.39 4.8 12.52 147.0

28 1.00 10.6 5,000 2,998 30 0.39 4.8 14.07 152.0

31 1.00 10.6 5,000 2,998 30 0.39 4.8 15.23 155.9

48 0.82 0.54 10.5 7.9 4,988 2,986 30 0.43 4.9 19.34 168.4

54 0.94 0.18 10.6 4,998 2,996 30 0.42 4.9 20.09 171.4

72 0.82 10.7 8.2 5,002 3,000 0.42 4.9 22.16 178.3

Total 7.38 0.34

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,002 0.11 0.23 42.7 85.5

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 38.9 22.7 0.24

4 59.7 28.3 0.24

7 67.7 33.0 0.33

24 70.0 41.5 0.60

26 92.7 55.7 0.60

28 92.9 57.6

31 0.60

48 0.88

54 95.3 65.0

72 95.6 67.6 1.23 0.17

Residue 4.4 32.4

Total 100.0 100.0

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

30

0.96

Distribution Reagent Consumption

93.1

95.6

59.1

63.8

0.60

Client: SilverCrest -Santa Elena project Date: 7-Oct-11

Test: C7 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'10" @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.3 99.7

100 149 4.3 95.5

150 105 12.6 82.9

200 74 16.2 66.7

270 53 14.9 51.8

325 44 3.9 47.8

400 37 6.4 41.4

Undersize - 37 41.4 -

TOTAL: 100.0

80 % Passing Size (μm) = 99

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 12-Oct-11

Test: C8 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: similar to C7, but with 0.25kg/t Pb(NO3)2

TEST CONDITIONS TEST DESCRIPTION

Solids: 1,995 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - added 0.25kg/t Pb(NO3)2 at the beginning

Grind Size - P80: 99 μm - adjusted to and maintained at 1.0g/L NaCN

Pb(NO3)2: 0.25 kg/t - sampled at 2,4,7,24,26,28,31,48,54 hours

Initial NaCN: 1.0 g/L - replaced solution with fresh water after 24 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 2.63 g/t 126.9 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

Added 0.25kg/t Pb(NO3 ) 2

0 1.00 3.00 0.32 8.1 10.0 4,995 3,000

2 0.88 0.36 10.4 5,000 3,005 30 1.21 3.6 23.24 69.8

4 0.98 0.06 10.5 5,002 3,007 30 1.35 4.1 27.78 84.2

7 0.94 0.18 10.4 7.4 5,002 3,007 30 1.38 4.2 29.70 90.8

24 0.80 10.5 7.5 4,996 3,001 30 1.35 4.2 39.70 121.5

Replaced solution with fresh water

24 1.00 3.00 10.3 5,002 3,007

26 1.00 10.6 7.2 5,002 3,007 30 0.25 4.9 9.02 148.7

28 0.98 0.06 10.6 5,000 3,005 30 0.26 5.0 10.45 153.2

31 0.95 0.15 10.5 5,000 3,005 30 0.25 4.9 11.80 157.6

48 0.82 0.54 10.4 7.6 4,988 2,993 30 0.27 5.0 16.38 171.5

54 0.90 0.30 10.5 4,998 3,003 30 0.26 5.0 16.94 173.8

72 0.88 10.6 8.3 4,998 3,003 0.27 5.0 18.15 178.0

Total 7.65 0.32

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 1,995 0.11 0.22 37.7 75.3

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 69.4 27.6 0.18

4 78.2 33.3 0.21

7 80.7 35.9 0.30

24 79.6 48.0 0.60

26 93.9 58.7 0.60

28 94.7 60.5 0.63

31 0.70

48 0.98

54 95.1 68.6

72 95.8 70.3 1.31 0.16

Residue 4.2 29.7

Total 100.0 100.0

30

1.13

Distribution Reagent Consumption

94.2

95.5

62.2

67.7

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

Client: SilverCrest -Santa Elena project Date: 7-Oct-11

Test: C8 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'10" @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.2 99.8

100 149 4.3 95.4

150 105 15.1 80.3

200 74 16.5 63.9

270 53 15.0 48.9

325 44 4.3 44.6

400 37 6.7 37.9

Undersize - 37 37.9 -

TOTAL: 100.0

80 % Passing Size (μm) = 104

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 12-Oct-11

Test: C9 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: similar to C7, but with 0.1kg/t H2O2

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,005 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - added 0.1kg/t H2O2 at the beginning

Grind Size - P80: 99 μm - adjusted to and maintained at 1.0g/L NaCN

H2O2: 0.1 kg/t - sampled at 2,4,7,24,26,28,31,48,54 hours

Initial NaCN: 1.0 g/L - replaced solution with fresh water after 24 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 2.80 g/t 133.7 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

Added 0.1kg/t H 2O2

0 1.00 3.00 0.30 8.2 10.0 5,005 3,000

2 0.90 0.30 10.5 4,998 2,993 30 0.81 2.4 21.63 64.7

4 0.95 0.15 10.5 5,002 2,997 30 1.09 3.3 26.23 79.3

7 0.94 0.18 10.5 8.4 4,998 2,993 30 1.21 3.7 29.65 90.2

24 0.82 10.4 8.0 4,996 2,991 30 1.28 3.9 37.30 113.9

Replaced solution with fresh water

24 1.00 3.00 10.4 5,000 2,995

26 1.00 10.7 9.2 4,998 2,993 30 0.41 5.1 12.38 150.9

28 1.00 10.7 5,000 2,995 30 0.42 5.2 14.02 156.2

31 1.00 10.7 5,000 2,995 30 0.42 5.2 15.12 159.9

48 0.85 0.45 10.4 7.6 4,988 2,983 30 0.46 5.3 19.09 172.1

54 0.90 0.30 10.5 4,996 2,991 30 0.44 5.3 20.00 175.5

72 0.80 10.5 8.3 4,994 2,989 0.46 5.4 22.02 182.1

Total 7.38 0.30

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,005 0.13 0.25 42.9 86.0

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 43.2 24.1 0.15

4 58.6 29.6 0.23

7 65.5 33.6 0.32

24 69.9 42.5 0.59

26 91.7 56.3 0.59

28 92.5 58.3 0.59

31 0.59

48 0.82

54 94.2 65.5

72 95.5 67.9 1.26 0.15

Residue 4.5 32.1

Total 100.0 100.0

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

30

0.97

Distribution Reagent Consumption

92.7

95.0

59.7

64.2

Client: SilverCrest -Santa Elena project Date: 7-Oct-11

Test: C9 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'10" @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.4 99.6

100 149 4.4 95.3

150 105 17.0 78.2

200 74 16.4 61.8

270 53 14.4 47.5

325 44 5.8 41.7

400 37 4.7 37.0

Undersize - 37 37.0 -

TOTAL: 100.0

80 % Passing Size (μm) = 109

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 12-Oct-11

Test: C10 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: similar to C7, but in 1.5g/L NaCN solution

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,007 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 99 μm - sampled at 2,4,7,24,26,28,31,48,54 hours

Initial NaCN: 1.5 g/L - replaced solution with fresh water after 24 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 2.87 g/t 135.7 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 4.50 0.32 8.3 10.1 5,007 3,000

2 1.36 0.42 10.6 5,000 2,993 30 0.72 2.2 21.98 65.8

4 1.45 0.15 10.6 4,998 2,991 30 0.93 2.8 26.28 79.3

7 1.45 0.15 10.6 7.5 5,000 2,993 30 1.13 3.4 29.80 90.6

24 1.30 10.7 7.4 5,000 2,993 30 1.38 4.2 43.40 132.2

Replaced solution with fresh water 10.4

24 1.50 4.50 10.8 5,000 2,993

26 1.50 10.8 7.8 5,000 2,993 30 0.38 5.4 12.14 168.5

28 1.50 10.8 5,000 2,993 30 0.41 5.5 13.32 172.4

31 1.50 10.6 5,000 2,993 30 0.40 5.4 14.17 175.4

48 1.40 0.30 10.7 7.2 4,990 2,983 30 0.42 5.5 17.13 184.5

54 1.42 0.24 10.6 5,010 3,003 30 0.42 5.5 17.51 186.5

72 1.30 7.9 4,998 2,991 0.42 5.5 19.02 191.3

Total 10.26 0.32

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,007 0.12 0.23 40.4 81.1

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 37.4 24.1 0.21

4 48.6 29.1 0.29

7 59.5 33.3 0.36

24 73.1 48.5 0.66

26 92.8 61.9 0.67

28 94.6 63.3 0.67

31 0.67

48 0.82

54 95.8 68.5

72 96.0 70.2 1.24 0.16

Residue 4.0 29.8

Total 100.0 100.0

30

0.93

Distribution Reagent Consumption

94.3

95.5

64.4

67.7

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

Client: SilverCrest -Santa Elena project Date: 7-Oct-11

Test: C10 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'10" @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.2 99.8

100 149 3.9 95.9

150 105 16.0 79.9

200 74 15.6 64.2

270 53 13.7 50.5

325 44 3.5 47.1

400 37 6.0 41.1

Undersize - 37 41.1 -

TOTAL: 100.0

80 % Passing Size (μm) = 105

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 2-Nov-11

Test: C11 & C12 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective: To improve Au and Ag recoveries with solution replacement after 14 and 24 hours at a target P80 grind of 100 microns

P80 Size, um % Solids NaCN, g/L Retention Au (g/t) Ag (g/t) Au (g/t) Ag (g/t) Au (%) Ag (%) Au (g/t) Ag (g/t) NaCN Lime

C11

NIVEL 4 -

Santa Elena

project

97 40 1.5 72 hours 3.31 119.1 3.00 125.9 96.3 67.0 0.11 41.6 1.97 0.12

C12

NIVEL 4 -

Santa Elena

project

108 40 2.0 72 hours 3.31 119.1 2.85 127.6 96.5 69.2 0.10 39.3 2.68 0.16

CYANIDATION TEST SUMMARY

Consumption (kg/t)

Test No Sample ID

Measured Head Calculated Head 72-Test Conditions h Extraction Residue Grade

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Gold Extraction Kinetics

C11 C12

0

10

20

30

40

50

60

70

80

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Silver Extraction Kinetics

C11 C12

Client: SilverCrest -Santa Elena project Date: 2-Nov-11

Test: C11 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,002 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 97 μm - sampled at 2,4,7,24,26,28,31,48,54 hours

Initial NaCN: 1.5 g/L - replaced solution with fresh water after 14 & 24 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 3.00 g/t 125.9 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 4.50 0.24 8.2 10.0 5,002 3,000

2 1.32 0.54 10.5 4,998 2,996 30 0.79 2.4 22.09 66.2

4 1.50 10.5 5,000 2,998 30 1.17 3.5 28.33 85.6

7 1.38 0.36 10.5 8.6 4,994 2,992 30 1.40 4.2 31.60 96.1

10 1.50 10.5 5,004 3,002 30 1.46 4.5 32.30 99.4

14 1.40 10.4 4,996 2,994 30 1.57 4.8 36.30 111.2

Replaced solution with fresh water

14 1.50 4.50 10.0 5,000 2,998

16 1.36 0.42 10.4 5,000 2,998 30 0.14 5.3 6.48 130.6

18 1.40 0.30 10.5 8.5 5,034 3,032 30 0.17 5.4 7.93 135.4

24 1.40 10.6 5,018 3,016 30 0.22 5.5 11.67 146.8

Replaced solution with fresh water

24 1.50 4.50 10.4 5,000 2,998

26 1.46 0.12 10.7 4,998 2,996 30 0.03 5.6 2.47 154.2

28 1.48 0.06 10.7 8.0 5,000 2,998 30 0.03 5.6 2.92 155.8

48 1.36 0.42 10.6 8.5 4,990 2,988 30 0.07 5.7 5.43 163.6

54 1.24 0.78 10.6 4,998 2,996 30 0.08 5.8 5.68 164.6

72 1.38 10.6 8.3 5,004 3,002 0.08 5.8 7.02 168.8

Total 16.50 0.24

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,002 0.11 0.22 41.6 83.3

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 39.5 26.3 0.27

4 58.9 34.0 0.27

7 70.8 38.1 0.45

10 74.8 39.5 0.45

14 80.8 44.1 0.60

16 87.8 51.8 0.81

18 89.4 53.7

24 92.0 58.2

26 1.16

28 1.19

48 1.41

54 96.2 65.3

72 96.3 67.0 1.97 0.12

Residue 3.7 33.0

Total 100.0 100.0

61.8

64.9

0.94

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

To improve Au and Ag recoveries in 1.5g/L NaCN with solution replacement after 14 and 24 hours at a target P80 grind of 100

microns

30

1.79

Distribution Reagent Consumption

93.6

1.10

93.5 61.2

95.7

Client: SilverCrest -Santa Elena project Date: 2-Nov-11

Test: C11 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'10" @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.5 99.5

100 149 1.7 97.9

150 105 12.7 85.1

200 74 20.2 65.0

270 53 14.5 50.5

325 44 4.4 46.1

400 37 4.8 41.3

Undersize - 37 41.3 -

TOTAL: 100.0

80 % Passing Size (μm) = 97

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 2-Nov-11

Test: C12 Project: 1103507

Sample: NIVEL 4 - Santa Elena project

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,006 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 2.0g/L NaCN

Grind Size - P80: 97 μm - sampled at 2,4,7,24,26,28,31,48,54 hours

Initial NaCN: 2.0 g/L - replaced solution with fresh water after 14 & 24 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 2.85 g/t 127.6 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 2.00 6.00 0.32 8.5 10.1 5,006 3,000

2 1.78 0.66 10.6 5,002 2,996 30 0.86 2.6 24.73 74.1

4 2.00 10.6 4,998 2,992 30 1.19 3.6 30.10 90.8

7 1.80 0.60 10.7 7.5 4,998 2,992 30 1.37 4.2 35.10 106.7

10 2.00 10.6 7.5 4,990 2,984 30 1.43 4.4 37.70 115.2

14 1.88 10.6 7.4 5,000 2,994 30 1.51 4.7 42.10 128.7

Replaced solution with fresh water

14 2.00 6.00 10.2 5,000 2,994

16 1.80 0.60 10.6 7.8 4,998 2,992 30 0.25 5.4 9.00 155.7

18 1.90 0.30 10.7 5,036 3,030 30 0.24 5.4 9.61 158.1

24 1.90 10.6 5,020 3,014 30 0.25 5.4 11.43 163.8

Replaced solution with fresh water

24 2.00 6.00 10.6 5,000 2,994

26 1.80 0.60 10.8 4,998 2,992 30 0.02 5.5 1.72 168.9

28 1.96 0.12 10.9 5,000 2,994 30 0.03 5.5 1.96 169.9

48 1.78 0.66 10.7 7.2 4,988 2,982 30 0.03 5.5 3.41 174.6

54 1.80 0.60 10.7 4,998 2,992 30 0.03 5.5 3.62 175.3

72 1.80 10.6 7.9 5,008 3,002 0.02 5.5 4.18 177.1

Total 22.14 0.32

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,006 0.10 0.20 39.3 78.8

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 45.1 28.9 0.33

4 62.8 35.5 0.34

7 72.9 41.7 0.64

10 76.5 45.0 0.64

14 81.7 50.3 0.81

16 94.8 60.8 1.12

18 94.6 61.8 1.23

24 95.1 64.0 1.40

26 1.70

28 1.76

48 2.10

54 97.0 68.5

72 96.5 69.2 2.68 0.16

Residue 3.5 30.8

Total 100.0 100.0

30

2.39

Distribution Reagent Consumption

96.8

97.0

66.4

68.2

96.2 66.0

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

Client: SilverCrest -Santa Elena project Date: 2-Nov-11

Test: C12 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'10" @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.5 99.5

100 149 6.0 93.5

150 105 14.7 78.8

200 74 22.6 56.2

270 53 15.5 40.7

325 44 4.9 35.7

400 37 5.0 30.8

Undersize - 37 30.8 -

TOTAL: 100.0

80 % Passing Size (μm) = 108

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest Date: 30-Jan-12

Test: C11 & C12 Project: 1103507

Sample: as specified

Objective: To recover Au and Ag with solution replacement after 10 hours at a target P80 grind of 100 microns

P80 Size, um % Solids NaCN, g/L Retention Au (g/t) Ag (g/t) Au (g/t) Ag (g/t) Au (%) Ag (%) Au (g/t) Ag (g/t) NaCN Lime

C13 CM Comp. 101 40 3.0 72 hours 0.50 153.4 0.55 181.6 51.2 54.1 0.27 83.4 2.76 0.30

C14 Hi-Mn Comp. 103 40 3.0 72 hours 5.04 93.4 5.07 119.9 92.1 60.2 0.40 47.7 2.11 0.19

CYANIDATION TEST SUMMARY

Test No Sample ID

Test Conditions Measured Head Calculated Head 72-h Extraction Residue Grade Consumption (kg/t)

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Gold Extraction Kinetics

C13 C14

0

10

20

30

40

50

60

70

80

90

100

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Silver Extraction Kinetics

Client: SilverCrest -Santa Elena project Date: 24-Jan-12

Test: C13 Project: 1103507

Sample: CM Comp.

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 1,983 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 3.0g/L NaCN

Grind Size - P80: 101 μm - sampled at 2,4,7,10,12,14,34,40,58,66 hours

Initial NaCN: 3.0 g/L - replaced solution with fresh water after 10 hours

Target pH: 10.0 - test ended after 82 hours

Test Duration: 82 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 0.55 g/t 181.6 g/t

Measured Total: 0.50 g/t 153.4 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 3.00 9.00 0.60 7.9 10.0 4,960 2,977

2 2.80 0.60 10.4 4,952 2,969 30 0.14 0.4 38.60 114.6

4 2.90 0.30 10.6 4,952 2,969 30 0.15 0.4 43.30 129.7

7 2.86 0.42 10.6 10.0 4,950 2,967 30 0.15 0.5 47.10 142.2

10 2.86 10.5 4,954 2,971 30 0.16 0.5 52.10 158.7

Replaced solution with fresh water

10 3.00 9.00 10.7 5,000 3,017

12 3.16 11.0 5,000 3,017 30 0.02 0.5 7.86 182.4

14 3.10 10.9 5,002 3,019 30 0.03 0.6 8.56 184.8

17 3.05 10.8 9.9 5

34 2.80 0.60 10.9 9.6 4,998 3,015 30 0.03 0.6 11.41 193.6

40 2.85 0.45 10.9 4,998 3,015 30 0.02 0.6 10.99 192.7

58 2.65 1.05 11.0 9.8 4,996 3,013 30 0.03 0.6 11.47 194.5

66 2.70 0.90 11.0 5,092 3,109 30 0.03 0.6 10.37 192.5

72 2.68 10.9 9.6 5,096 3,113 0.02 0.6 11.02 194.9

Total 22.32 0.60

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 1,983 0.27 0.53 83.4 165.3

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 38.3 31.8 0.35

4 41.4 36.0 0.50

7 41.8 39.5 0.71

10 45.0 44.0 0.92

12 50.6 50.6 0.65

14 53.4 51.3 0.74

34 53.5 53.8

40 50.8 53.5

58 1.96

66 2.28

72 51.2 54.1 2.76 0.30

Residue 48.8 45.9

Total 100.0 100.0

53.7 54.0

53.4

1.20

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

To improve Au and Ag recoveries in 3.0g/L NaCN with solution replacement after 10 hours at a target P80 grind of 100 microns

25

Distribution Reagent Consumption

54.0

1.43

Client: SilverCrest -Santa Elena project Date: 24-Jan-12

Test: C13 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 36' @ 65% solids in stainless steel Mill #4

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.4 99.6

100 149 4.2 95.4

150 105 12.9 82.5

200 74 20.6 62.0

270 53 15.2 46.8

325 44 5.5 41.3

400 37 5.1 36.3

Undersize - 37 36.3 -

TOTAL: 100.0

80 % Passing Size (μm) = 101

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 24-Jan-12

Test: C14 Project: 1103507

Sample: Hi-Mn Comp.

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 1,978 g - repulped to 40% solids

Solution: 3,000 g - adjusted to and maintained pH 10.0

Solids: 40 % - adjusted to and maintained at 3.0g/L NaCN

Grind Size - P80: 103 μm - sampled at 2,4,7,10,12,14,34,40,58,66 hours

Initial NaCN: 3.0 g/L - replaced solution with fresh water after 10 hours

Target pH: 10.0 - test ended after 82 hours

Test Duration: 82 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 5.07 g/t 119.9 g/t

Measured Total: 5.04 g/t 93.4 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 3.00 9.00 0.38 8.1 10.0 4,952 2,974

2 2.82 0.54 10.8 4,952 2,974 30 1.99 5.9 35.00 104.1

4 3.00 10.9 4,954 2,976 30 2.22 6.7 37.00 111.1

7 2.90 0.30 10.9 9.4 4,950 2,972 30 2.38 7.2 38.90 117.8

10 2.90 10.9 4,954 2,976 30 2.59 7.9 39.20 120.0

Replaced solution with fresh water

10 3.00 9.00 10.7 5,000 3,022

12 3.20 11.1 4,996 3,018 30 0.40 9.1 6.28 138.9

14 3.16 11.1 5,004 3,026 30 0.41 9.2 6.30 139.2

17 3.12 11.0 9.4 5

34 2.88 0.36 11.0 9.2 4,996 3,018 30 0.42 9.2 7.02 141.6

40 2.85 0.45 11.1 5,018 3,040 30 0.43 9.3 6.85 141.4

58 2.68 0.96 11.1 9.5 4,996 3,018 30 0.44 9.3 7.20 142.5

66 2.74 0.78 11.1 5,130 3,152 30 0.39 9.2 6.57 141.7

72 2.72 11.1 9.6 5,134 3,156 0.40 9.2 6.86 142.9

Total 21.39 0.38

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 1,978 0.40 0.79 47.7 94.4

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 59.0 43.9 0.31

4 66.4 46.8 0.31

7 71.7 49.6 0.47

10 78.8 50.6 0.61

12 90.8 58.5 0.28

14 91.2 58.7 0.33

34 91.6 59.7

40 92.2 59.6

58 1.48

66 1.69

72 92.1 60.2 2.11 0.19

Residue 7.9 39.8

Total 100.0 100.0

CYANIDATION TEST REPORT

To improve Au and Ag recoveries in 3.0g/L NaCN with solution replacement after 10 hours at a target P80 grind of 100 microns

Solution

Ag

Total Residue

Distribution Reagent Consumption

25

0.77

0.96

92.5 60.1

91.6 59.7

Client: SilverCrest -Santa Elena project Date: 24-Jan-12

Test: C14 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21' @ 65% solids in stainless steel Mill #4

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.5 99.5

100 149 5.3 94.2

150 105 12.9 81.3

200 74 18.4 62.9

270 53 14.5 48.4

325 44 5.3 43.1

400 37 5.3 37.8

Undersize - 37 37.8 -

TOTAL: 100.0

80 % Passing Size (μm) = 103

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 9-Jul-12

Test: as specified Project: 1103507

Sample: Composite 4 Nivel

Objective: To recover Au and Ag with solution replacement after 10 hours at a target P80 grind of 100 microns

Solution

Replacement P80 Size, um % Solids NaCN, g/L Retention Au (g/t) Ag (g/t) Au (g/t) Ag (g/t) Au (%) Ag (%) Au (g/t) Ag (g/t) NaCN Lime

C15

Composite 4

Nivel

100% sol’n

replacement By

filtration

98 45 1.0 72 hours 1.46 38.5 1.39 39.4 97.1 43.7 0.04 22.2 2.02 0.24

C16

Composite 4

Nivel

100% sol’n

replacement By

filtration

98 45 1.5 72 hours 1.46 38.5 1.50 38.7 98.0 43.6 0.03 21.8 2.60 0.21

C17

Composite 4

Nivel

60% sol’n

replacement By

filtration

98 45 1.5 72 hours 1.46 38.5 1.35 38.9 97.8 43.5 0.03 22.0 2.09 0.21

CYANIDATION TEST SUMMARY

Consumption (kg/t)

Test No Sample ID

Test Conditions Measured Head Calculated Head 72-h Extraction Residue Grade

0

20

40

60

80

100

120

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Gold Extraction Kinetics

C15 C16 C17

0

5

10

15

20

25

30

35

40

45

50

0 10 20 30 40 50 60 70 80

Extractions (%)

Leach Time (hours)

Silver Extraction Kinetics

C15 C16 C17

Client: SilverCrest -Santa Elena project Date: 5-Jul-12

Test: C15 Project: 1103507

Sample: Composite 4 Nivel

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 1,980 g - repulped to 45% solids with site water

Solution: 2,444 g - adjusted to and maintained pH 10.0 with hydrated lime

Solids: 45 % - adjusted to and maintained at 1.0g/L NaCN

Grind Size - P80: 98 μm - sampled at 2,4,7,210,12,14,34,58 hours

Initial NaCN: 1.0 g/L - replaced solution with fresh water after 10 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 1.39 g/t 39.4 g/t

Measured Total: 1.46 g/t 38.5 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.00 2.44 0.44 7.8 10.0 4,424 2,444

2 0.88 0.29 10.1 4,448 2,468 30 0.73 1.8 8.60 21.2

4 0.92 0.20 10.1 7.8 4,442 2,462 30 0.87 2.2 9.40 23.4

7 0.95 0.12 10.1 4,444 2,464 30 0.94 2.4 10.10 25.4

10 0.82 10.1 4,440 2,460 30 0.97 2.5 10.90 27.7

Replaced solution with fresh site water

10 0.04 2.35 0.04 9.8 4,444 2,464

12 0.95 0.12 10.2 4,444 2,464 30 0.23 2.7 3.10 31.2

14 0.80 0.49 10.2 7.1 4,452 2,472 30 0.25 2.7 3.40 32.0

34 0.80 0.49 10.2 7.2 4,430 2,450 30 0.25 2.7 3.70 32.8

58 0.80 0.49 10.3 7.3 4,442 2,462 30 0.23 2.7 4.20 34.2

72 0.40 10.3 9.7 4,404 2,424 0.23 2.7 4.20 34.1

Total 6.99 0.48

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 1,980 0.04 0.08 22.2 43.9

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 65.3 27.2 0.14

4 78.4 30.0 0.24

7 85.7 32.6 0.30

10 89.3 35.4 0.52

12 96.4 39.9 0.53

14 98.5 41.0 0.77

34 98.6 42.0 1.03

58 1.27

72 97.1 43.7 2.02 0.24

Residue 2.9 56.3

Total 100.0 100.0

25

Distribution Reagent Consumption

97.2 43.8

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

Client: SilverCrest -Santa Elena project Date: 5-Jul-12

Test: C16 Project: 1103507

Sample: Composite 4 Nivel

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,000 g - repulped to 45% solids with site water

Solution: 2,444 g - adjusted to and maintained pH 10.0 with hydrated lime

Solids: 45 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 98 μm - sampled at 2,4,7,10,12,14,34,58 hours

Initial NaCN: 1.5 g/L - replaced solution with fresh water after 10 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 1.50 g/t 38.7 g/t

Measured Total: 1.46 g/t 38.5 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 3.67 0.42 7.9 10.0 4,444 2,444

2 1.32 0.44 10.1 4,442 2,442 30 0.75 1.8 9.30 22.7

4 1.40 0.24 10.2 8.0 4,442 2,442 30 0.92 2.3 10.10 24.9

7 1.42 0.20 10.3 4,442 2,442 30 1.02 2.5 11.00 27.4

10 1.30 10.2 4,444 2,444 30 1.08 2.7 11.50 29.0

Replaced solution with fresh site water

10 0.14 3.32 10.0 4,444 2,444

12 1.30 0.49 10.2 4,444 2,444 30 0.25 2.9 2.90 31.8

14 1.30 0.49 10.2 5.4 4,454 2,454 30 0.25 2.9 3.00 32.1

34 1.20 0.73 10.4 6.8 4,442 2,442 30 0.26 3.0 3.30 32.9

58 1.30 0.49 10.5 7.0 4,448 2,448 30 0.25 2.9 3.90 34.5

72 0.70 9.2 4,407 2,407 0.25 2.9 3.60 33.7

Total 10.07 0.42

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,000 0.03 0.06 21.8 43.6

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 61.0 29.4 0.22

4 75.5 32.3 0.34

7 84.6 35.5 0.44

10 90.5 37.5 0.68

12 97.3 41.1 0.76

14 97.6 41.5 1.00

34 98.6 42.5 1.37

58 1.61

72 98.0 43.6 2.60 0.21

Residue 2.0 56.4

Total 100.0 100.0

98.1 44.6

25

CYANIDATION TEST REPORT

To improve Au and Ag recoveries in 1.5g/L NaCN with solution replacement after 10 hours at a target P80 grind of 100 microns

Solution

Ag

Total Residue

Distribution Reagent Consumption

Client: SilverCrest -Santa Elena project Date: 5-Jul-12

Test: C15 Project: 1103507

Sample: Composite 4 Nivel

Objective:

TEST CONDITIONS TEST DESCRIPTION

Solids: 2,000 g - repulped to 45% solids with site water

Solution: 2,444 g - adjusted to and maintained pH 10.0 with hydrated lime

Solids: 45 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 98 μm - sampled at 2,4,7,10,12,14,34,58 hours

Initial NaCN: 1.5 g/L - replaced solution with fresh water after 10 hours

Target pH: 10.0 - test ended after 72 hours

Test Duration: 72 hours - filtered and displacement washed with hot cyanide solution

followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 1.35 g/t 38.9 g/t

Measured Total: 1.46 g/t 38.5 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 3.67 0.42 8.0 10.0 4,444 2,444

2 1.32 0.44 10.2 4,444 2,444 30 0.70 1.7 9.30 22.7

4 1.44 0.15 10.2 7.3 4,444 2,444 30 0.90 2.2 10.20 25.2

7 1.40 0.24 10.2 4,442 2,442 30 0.96 2.4 10.80 27.0

10 1.28 10.2 4,444 2,444 30 0.98 2.5 11.50 29.0

Replaced 60% solution with fresh site water and 40% solution with 10hours' PLS

10 0.52 2.40 10.1 4,444 2,444

12 1.30 0.49 10.3 4,444 2,444 30 0.47 2.6 5.60 31.1

14 1.30 0.49 10.3 7.1 4,444 2,444 30 0.46 2.6 5.80 31.8

34 1.20 0.73 10.3 7.2 4,428 2,428 30 0.47 2.7 6.30 33.0

58 1.30 0.49 10.6 7.3 4,430 2,430 30 0.47 2.7 6.80 34.5

72 0.74 10.6 8.6 4,412 2,412 0.46 2.6 6.50 33.8

Total 9.09 0.42

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

72 2,000 0.03 0.06 22.0 44.0

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 63.2 29.2 0.22

4 82.0 32.4 0.29

7 88.3 34.6 0.42

10 91.3 37.3 0.68

12 97.2 40.0 0.29

14 96.8 40.8 0.54

34 97.9 42.5 0.91

58 1.16

72 97.8 43.5 2.09 0.21

Residue 2.2 56.5

Total 100.0 100.0

98.5 44.3

25

CYANIDATION TEST REPORT

To improve Au and Ag recoveries in 1.5g/L NaCN with solution replacement after 10 hours at a target P80 grind of 100 microns

Solution

Ag

Total Residue

Distribution Reagent Consumption

Client: SilverCrest -Santa Elena project Date: 5-Jul-12

Test: C15-C17 Project: 1103507

Sample: Cyanidation Residue

Grind: 2.0kg grind for 21'30" with site water @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.7 99.3

100 149 3.6 95.7

150 105 12.9 82.8

200 74 14.1 68.7

270 53 11.4 57.3

325 44 5.4 51.9

400 37 4.1 47.8

Undersize - 37 47.8 -

TOTAL: 100.0

80 % Passing Size (μm) = 98

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 11-Sep-12

Test: C18 -C19 Project: 1103507

Sample: as specified

Objective: Tank leach tests to produce tailings for filtration

P80 Size, mm % Solids NaCN, g/L Retention Au (g/t) Ag (g/t) Au (g/t) Ag (g/t) Au (%) Ag (%) Au (g/t) Ag (g/t) NaCN Lime

C18

New

Composite 4

Nivel

97 48 1.5 20 hours 1.46 36.8 1.07 36.8 94.4 42.4 0.06 21.2 0.52 0.13

C19 Nivel 4 Comp. 111 48 1.5 8 hours 3.31 119.1 1.92 113.8 84.9 51.8 0.29 54.9 0.35 0.12

CYANIDATION TEST SUMMARY

Consumption (kg/t)

Test No Sample ID

Measured Head Calculated Head Test Conditions Extraction Residue Grade

0

20

40

60

80

100

0 4 8 12 16 20

Extractions, %

Leach time, hours

Gold Leach Kinetics

C18 C19

0

20

40

60

80

100

0 4 8 12 16 20

Extractions, %

Leach time, hours

Silver Leach kinetics

C18 C19

Client: SilverCrest -Santa Elena project Date: 11-Sep-12

Test: C18 Project: 1103507

Sample: New Composite 4 Nivel

Objective: Tank leach tests to produce tailings for filtration

TEST CONDITIONS TEST DESCRIPTION

Solids: 5,510 g - repulped to 48% solids with synthetic water

Solution: 5,969 g - adjusted to and maintained pH 10.2 with hydrated lime

Solids: 48 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 97 μm - sampled at 2 and 4 hours

Initial NaCN: 1.5 g/L - test ended after 20 hours

Target pH: 10.2 - filtered and displacement washed cut slurry with hot cyanide solution

Test Duration: 20 hours followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 1.07 g/t 36.8 g/t

Measured Total: 1.46 g/t 38.5 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 8.95 0.70 8.1 10.2

2 1.46 0.24 10.3 6.0 11,494 5,984 30 0.67 4.0 10.8 64.6

4 1.38 0.72 10.3 11,484 5,974 30 0.86 5.2 11.9 71.4

20 1.18 10.3 6.5 11,472 5,962 0.93 5.6 14.3 85.9

Total 9.91 0.70

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

20 5,510 0.06 0.33 21.2 116.8

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 67.7 31.9 0.04

4 87.1 35.2 0.17

20 94.4 42.4 0.52 0.13

Residue 5.6 57.6

Total 100.0 100.0

50

Distribution Reagent Consumption

CYANIDATION TEST REPORT

Total Residue

Solution

Ag

Client: SilverCrest -Santa Elena project Date: 11-Sep-12

Test: C18 Project: 1103507

Sample: Cyanidation Residue

Grind: 2kg grind for 21'30" with site water @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.5 99.5

100 149 1.7 97.9

150 105 12.7 85.1

200 74 20.2 65.0

270 53 14.5 50.5

325 44 4.4 46.1

400 37 4.8 41.3

Undersize - 37 41.3 -

TOTAL: 100.0

80 % Passing Size (μm) = 97

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 11-Sep-12

Test: C19 Project: 1103507

Sample: Nivel 4 Comp.

Objective: Tank leach tests to produce tailings for filtration

TEST CONDITIONS TEST DESCRIPTION

Solids: 20,000 g - repulped to 48% solids with synthetic water

Solution: 21,667 g - adjusted to and maintained pH 10.2 with hydrated lime

Solids: 48 % - adjusted to and maintained at 1.5g/L NaCN

Grind Size - P80: 111 μm - sampled at 2 and 4 hours

Initial NaCN: 1.5 g/L - test ended after 8 hours

Target pH: 10.2 - filtered and displacement washed cut slurry with hot cyanide solution

Test Duration: 8 hours followed by two hot water displacement washes

- solution and solids assayed for Au, Ag content

HEAD GRADE

Au Ag

Calculated Total: 1.92 g/t 113.8 g/t

Measured Total: 3.31 g/t 119.1 g/t

LEACH TEST DATA

Time NaCN Lime pH dO2 Slurry

Weight Vol. Assay Vol. Au

(hours) (g/L) (g) (g) before after (mg/L) (g) (mL) (mL) (mg/L) (mg) (mg/L) (mg)

0 1.50 32.50 2.30 8.5 10.4

2 1.32 3.90 10.6 9.6 41,667 21,667 30 1.23 26.7 36.0 780.0

4 1.40 2.17 10.6 41,667 21,667 30 1.46 31.7 45.6 989.1

8 1.46 10.7 41,667 21,667 1.50 32.6 54.3 1179.0

Total 38.57 2.30

SOLIDS

Time

Weight Au Ag

(hours) (g) (g/t) (mg) (g/t) (mg)

8 20,000 0.29 5.80 54.9 1098.0

CYANIDATION RESULTS

Time Reducing Power

Au Ag NaCN Ca(OH)2 0.1 N KMnO4/L

(hours) (%) (%) (kg/t) (kg/t) (mL)

2 69.4 34.3 0.19

4 82.5 43.4 0.30

8 84.9 51.8 0.35 0.12

Residue 15.1 48.2

Total 100.0 100.0

75

CYANIDATION TEST REPORT

Solution

Ag

Total Residue

Distribution Reagent Consumption

Client: SilverCrest -Santa Elena project Date: 11-Sep-12

Test: C19 Project: 1103507

Sample: Cyanidation Residue

Grind: 2kg grind for 21'30" with site water @ 65% solids in stainless steel Mill #3

Sieve Size Individual Cumulative

Tyler Mesh Micrometers % Retained % Passing

65 210 0.5 99.5

100 149 5.1 94.4

150 105 16.8 77.6

200 74 18.7 58.9

270 53 12.9 45.9

325 44 6.0 39.9

400 37 5.0 34.9

Undersize - 37 38.0 -

TOTAL: 103.1

80 % Passing Size (μm) = 111

SIZE ANALYSIS REPORT

0

20

40

60

80

100

0

10

20

30

40

50

60

70

80

90

100

- 37 37 44 53 74 105 149 210

Cumulative % Passing

Individual % Retained

Particle Size, μm

Client: SilverCrest -Santa Elena project Date: 7-Nov-11

Test: CD1 Project: 1103507

Sample: C11 final leach slurry

Objective: Scoping test to remove CN in C11 leach slurry using INCO cyanide destruction method

TEST CONDITIONS TEST DESCRIPTION

Solids: 554 g - sample placed in a 2L glass reactor with overhead agitation

Solution: 824 g - air sparged via 1 pump (~1L/min air)

Solids: 40 % - Na2S2O5 (MBS) added as 100g/L solution

Slurry weight: 1,378 g - maintained pH in range 8.5-9.0 during the test

Initial CN: 620.0 mg/L - maintained ORP above +70mV during the test

Initial ORP: 114.0 mV - monitored free cyanide concentration by titration with AgNO3

Initial pH: 10.2

CuSO4 addition: 0.7 g

Test Duration: 3.5 hours

LEACH TEST DATA

Time MBS Lime pH ORP Free CN

CN total (WAD) CN- SCN CNO

min g mV mg/L mg/L mg/L mg/L mg/L

0 10.2 114 690 620.00 600.00 5.62 11.44

15 9.0 196 239

30 0.10 8.5 211 127

45 0.30 8.5 212 85

60 0.46 8.6 112 42

90 0.38 8.5 290 3.22

120 0.36 8.7 265 1.79

150 0.30 8.5 332 1.39

180 0.22 8.5 325 1.09

210 0.25 8.5 362 0.37 0.35 1.05 381.08

Total 2.37

CYANIDE DESTRUCTION TEST REPORT

0.70

Final Solution Assay

g

1.00

0.60

6.10

Assay report

Client: SilverCrest -Santa Elena project Date: 7-Nov-11

Sample: as-specified Project: 1103507

Analytical

C11 72h Pls DC11 Final Pls Method

SO4 mg/L 16.9 6583.3 WetA

Ag mg/L 5.83 0.59 H2O-ICP

Al mg/L 0.8 <0.2 H2O-ICP

As mg/L <0.2 <0.2 H2O-ICP

Ba mg/L 0.12 <0.01 H2O-ICP

Bi mg/L <0.1 <0.1 H2O-ICP

Ca mg/L 2 459.4 H2O-ICP

Cd mg/L <0.01 <0.01 H2O-ICP

Co mg/L <0.01 <0.01 H2O-ICP

Cr mg/L <0.1 <0.1 H2O-ICP

Cu mg/L 1.46 0.88 H2O-ICP

Fe mg/L 8.43 0.17 H2O-ICP

Hg mg/L <0.05 <0.05 H2O-ICP

K mg/L 10 23 H2O-ICP

La mg/L 0.07 <0.05 H2O-ICP

Mg mg/L 0.6 8.7 H2O-ICP

Mn mg/L 2.03 <0.05 H2O-ICP

Mo mg/L 0.03 <0.02 H2O-ICP

Na mg/L 823.00 2308 H2O-ICP

Ni mg/L 0.12 0.25 H2O-ICP

P mg/L 0.20 0.8 H2O-ICP

Pb mg/L <0.05 0.25 H2O-ICP

Sb mg/L <0.1 <0.1 H2O-ICP

Sc mg/L <0.05 <0.05 H2O-ICP

Sr mg/L <0.01 1.85 H2O-ICP

Ti mg/L <0.1 <0.1 H2O-ICP

Tl mg/L <0.2 <0.2 H2O-ICP

V mg/L <0.01 0.07 H2O-ICP

W mg/L <0.1 0.3 H2O-ICP

Zn mg/L 2.54 0.45 H2O-ICP

Zr mg/L <0.01 <0.01 H2O-ICP

Client: Date: 2-Nov-11

Test: ST-1 Project: 1103507

Sample: C12 slurry ( 30% solids with rake ) Page: 1 of 2

Time Height Sludge Density Slurry pH: 10.5

(min.) (cm) (w/w % solids) Coagulant: n/a

0 34.9 29.0 Flocculant:

1 32.3 30.9

2 30.1 32.7 Dry Solids Density: 2.67 g/cm3

3 27.5 35.1 Liquid Density: 1.00 g/cm3

4 26.0 36.7 Weight of Dry Solids: 709.3 g

5 24.4 38.5

6 22.5 40.9 Initial Slurry Weight: 2456 g

8 20.9 43.2 Initial Slurry Volume: 2000 mL

10 19.3 45.8 Initial Slurry Height: 34.9 cm

15 17.3 49.4 Initial Weight Percent Solids: 28.9 w/w % solids

20 16.1 51.9 Initial Settling Rate: 1.6 m/h

40 13.7 57.7

60 13.0 59.7 Final Sediment Volume: 500 mL

120 11.4 64.7 Final Sediment Height: 8.8 cm

180 10.9 66.4

270 9.8 70.6 Supernatant Clarity: 3

420 9.2 73.1 Floc Size: 1

1440 8.8 74.8

Supernatant Clarity Scale Floc Size Scale

0 Crystal Clear, zero suspended solids 1 Very fine particles

1 Transparent - some suspended solids 2 to 9 Floc size increasing

2 Somewhat transparent solution 10 Very large flocs

3 Less cloudy, non-transparent solution

4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

Unit Thickener Area Determination

Modified Coe and Clevenger Method/ Oltman Technique

Required Underflow Pulp Density: 50 w/w % solids

Compression Point: 16.7 cm

17.5 min

Slope (Settling Rate), R: 15.0 m/d

Feed Dilution, F: 2.46 (weight solution/weight solids)

Underflow Dilution, Du: 1.00 (weight solution/weight solids)

Liquid Relative Density, L: 1.00

Unit Thickener Area, A: 0.10 m2/tpd solids A = (F - Du)

RL

SETTLING TEST REPORT

Magnafloc 368, 30g/t

SilverCrest - Santa Elena

Client: Date: 2-Nov-11

Test: ST-1 Project: 1103507

Sample: C12 slurry ( 30% solids with rake ) Page: 1 of 2

SETTLING TEST REPORT

SilverCrest - Santa Elena

Client: SilverCrest - Santa Elena Date: 2-Nov-11

Test: ST-1 Project: 1103507

Sample: C12 slurry ( 30% solids with rake ) Page: 2 of 2

SETTLING TEST REPORT

0

5

10

15

20

25

30

35

40

45

0 100 200 300 400 500 600 700 800 900 1000 1100 1200 1300 1400 1500

Interface Height (cm)

Time (minutes)

Client: Date: 06/28/12

Test: ST-2 Project: 1103507

Sample: C17 slurry ( 20% solids with rake ) Page: 1 of 2

Time Height Sludge Density Slurry pH: 10.5

(min.) (cm) (w/w % solids) Coagulant: n/a

0 39.7 19.8 Flocculant:

7 10.2 56.7

8 9.4 59.7 Dry Solids Density: 2.68 g/cm3

10 8.2 64.9 Liquid Density: 1.00 g/cm3

13 8.1 65.4 Weight of Dry Solids: 452.0 g

18 7.8 66.8

30 7.4 68.9 Initial Slurry Weight: 2264 g

60 6.9 71.6 Initial Slurry Volume: 2000 mL

129 6.5 74.0 Initial Slurry Height: 39.7 cm

312 6.2 75.9 Initial Weight Percent Solids: 20.0 w/w % solids

461 6.0 77.2 Initial Settling Rate: 2.5 m/h

1440 5.7 79.2

Final Sediment Volume: 290 mL

Final Sediment Height: 5.7 cm

Supernatant Clarity: 3.5

Floc Size: 1

Supernatant Clarity Scale Floc Size Scale

0 Crystal Clear, zero suspended solids 1 Very fine particles

1 Transparent - some suspended solids 2 to 9 Floc size increasing

2 Somewhat transparent solution 10 Very large flocs

3 Less cloudy, non-transparent solution

4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

Unit Thickener Area Determination

Modified Coe and Clevenger Method/ Oltman Technique

Required Underflow Pulp Density: 60 w/w % solids

Compression Point: 8.8 cm

9.0 min

Slope (Settling Rate), R: 49.4 m/d

Feed Dilution, F: 4.01 (weight solution/weight solids)

Underflow Dilution, Du: 0.67 (weight solution/weight solids)

Liquid Relative Density, L: 1.00

Unit Thickener Area, A: 0.07 m2/tpd solids A = (F - Du)

RL

SETTLING TEST REPORT

SETTLING TEST REPORT

Magnafloc 368, 40g/t

SilverCrest - Santa Elena

Client: SilverCrest - Santa Elena       Date: 06/28/12 Test: ST-2

Project: 1103507 Sample: C17 slurry ( 20% solids with rake )    Page: 1 of  Time

(min.)  Height

(cm)    Sludge Density

(w/w % solids)0 39.7    19.8 7  10.2    56.7 8  9.4     59.7 10 8.2     64.9

13      8.1     65.4

18      7.8     66.8

30      7.4     68.9

60      6.9     71.6

129     6.5     74.0

312     6.2     75.9

461     6.0     77.2

1440    5.7     79.2 Slurry pH: 10.5 Coagulant:      n/a

Flocculant: Magnafloc 368, 40g/t Dry Solids Density:     2.68 g/cm3 Liquid Density: 1.00 g/cm3 Weight of Dry Solids:   452.0 g Initial Slurry Weight:  2264 g Initial Slurry Volume:  2000 mL Initial Slurry Height:  39.7 cm Initial Weight Percent Solids:  20.0 w/w % solids Initial Settling Rate:  2.5 m/h Final Sediment Volume:  290 mL Final Sediment Height:  5.7 cm Supernatant Clarity:    3.5

Floc Size:      1 Supernatant Clarity Scale       Floc Size Scale 0       Crystal Clear, zero suspended solids    1 Very fine particles 1       Transparent - some suspended solids     2 to 9 Floc size increasing  2       Somewhat transparent solution   10 Very large flocs 3       Less cloudy, non-transparent solution 4       Very cloudy discernible solid/liquid interface 5       Opaque, no solid/liquid interface visible

 Unit Thickener Area Determination Modified Coe and Clevenger Method/ Oltman Technique

 Required Underflow Pulp Density: Compression Point: Slope (Settling Rate), R Feed Dilution, F Underflow Dilution, Du Liquid Relative Density, L Unit Thickener Area, A 60 w/w % solids 8.8 cm 9.0     min 49.4 m/d 4.1     (weight solution/weight solids) 0.67 (weight solution/weight solids) 1.00

Client: SilverCrest - Santa Elena       Date: 06/28/12

Test: ST-2      Project: 1103507

Sample: C17 slurry ( 20% solids with rake )     Page: 1 of 2

SETTLING TEST REPORT

Client: SilverCrest - Santa Elena       Date: 28-Jun-12

Test: ST-2      Project: 1103507

Sample: C17 slurry ( 20% solids with rake )     Page: 2 of 2

Interface Height, H versus Time, t

Client: Pulp density: 40% Date: 2-Nov-11

Sample id: ST-1 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-1

Test description: Sample slurried to 40% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-1 27 27 60 102 - 0.5 4.41 3.13 29.02 - - - 55 3-4

VF-1(Rep.) 27 27 30 53 - 0.3 4.05 2.85 29.63 - - - 30 4

Note:

Filter cloth blinded, after 5seconds no flow (no cake formed)

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 1.5 kg/m2/h 7.4 Filtrate Clarity Scale

Dry lbs/sq.ft./h= 2.7 kg/m2/h 13.1 0 Crystal Clear, zero suspended solids

2 Somewhat transparent solution

Filtrate Capacity Determination 3 Less cloudy, non-transparent solution

4 Very cloudy discernible solid/liquid interface

VF-1: Gal/sq.ft./h= 3.2 L/m2/h 130 5 Opaque, no solid/liquid interface visible

Gal/sq.ft./h= 3.4 L/m2/h 138 VF-1Rep

VACUUM FILTRATION TEST REPORT

Client: Pulp density: 40% Date: 2-Nov-11

Sample id: ST-1 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech (PO-808HF) Test series: VF-2

Test description: Sample slurried to 40% solids and submitted to leaf filtration test using PO-808HF filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-2 27 27 60 80 - 1.0 18.47 12.75 30.97 - - - 151 5

Note: VACUUM FILTRATION TEST REPORT

Cake formed but fine solids into filtrate

Cake Capacity Determination

Filtrate Clarity Scale

VF-2: Dry lbs/sq.ft./h= 7.1 kg/m2/h 34.8 0 Crystal Clear, zero suspended solids

2 Somewhat transparent solution

Filtrate Capacity Determination 3 Less cloudy, non-transparent solution

4 Very cloudy discernible solid/liquid interface

VF-2: Gal/sq.ft./h= 10.1 L/m2/h 413 5 Opaque, no solid/liquid interface visible SilverCrest - Santa Elena

VACUUM FILTRATION TEST REPORT
SilverCrest - Santa Elena
Client: Pulp density: 50% Date: 3-Jul-12
Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507
Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG
Filter cloth: Envirotech POPR-901F Test series: VF-3
Test description: Sample slurried to 50% solids and submitted to leaf filtration test using POPR 901F filter cloth.
Test Vacuum Filtration Time Cake Crust Filtrate
No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity
in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)
VF-3 27 27 60 90 - 0.3 3.17 2.42 23.82 - - - 30 2-3
Filter cloth blinded, after 5seconds no flow (no cake formed)
Cake Capacity Determination
VF-1: Dry lbs/sq.ft./h= 1.3 kg/m2/h 6.2 Filtrate Clarity Scale
0 Crystal Clear, zero suspended solids
Filtrate Capacity Determination 2 Somewhat transparent solution
3 Less cloudy, non-transparent solution
VF-1: Gal/sq.ft./h= 1.9 L/m2/h 76 4 Very cloudy discernible solid/liquid interface
5 Opaque, no solid/liquid interface visible

Client: Pulp density: 40% Date: 2-Nov-11

VACUUM FILTRATION TEST REPORTSilverCrest - Santa Elena

Client: Pulp density: 50% Date: 3-Jul-12

Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-3

Test description: Sample slurried to 50% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-3 27 27 60 90 - 0.3 3.17 2.42 23.82 - - - 30 2-3

Filter cloth blinded, after 5seconds no flow (no cake formed)

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 1.3 kg/m2/h 6.2 Filtrate Clarity Scale

0 Crystal Clear, zero suspended solids

Filtrate Capacity Determination 2 Somewhat transparent solution

3 Less cloudy, non-transparent solution

VF-1: Gal/sq.ft./h= 1.9 L/m2/h 76 4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

Client: Pulp density: 60% Date: 3-Jul-12 SilverCrest - Santa Elena VACUUM FILTRATION TEST REPORT

Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-4

Test description: Sample slurried to 60% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-4 27 27 60 90 - 0.5 10.93 8.17 25.25 - - - 20 2-3

Filter cloth blinded, after 5 seconds no flow (no cake formed)

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 4.3 kg/m2/h 20.8 Filtrate Clarity Scale

0 Crystal Clear, zero suspended solids

Filtrate Capacity Determination 2 Somewhat transparent solution

3 Less cloudy, non-transparent solution

VF-1: Gal/sq.ft./h= 1.3 L/m2/h 51 4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

Client: Pulp density: 70% Date: 3-Jul-12 SilverCrest - Santa Elena VACUUM FILTRATION TEST REPORT

Sample id: ST-2 slurry Pulp temperature: 20oC Project: 1103507

Filter area: 94.12cm2 (0.1013sq.ft.) Technician: JZ/BG

Filter cloth: Envirotech POPR-901F Test series: VF-5

Test description: Sample slurried to 60% solids and submitted to leaf filtration test using POPR 901F filter cloth.

Test Vacuum Filtration Time Cake Crust Filtrate

No. Form Dry Form Dry Cracks Thickness Wet Wt. Dry Wt. Moisture Wet Dry Moisture Volume Clarity

in Hg in Hg sec sec sec (mm) (g) (g) (%) (g) (g) (%) (mL)

VF-5 27 27 60 90 - 1 15.02 12.78 14.92 1.336 1.211 9.36 25 2-3

Thin cake formed but fine solids into filtrate

Cake Capacity Determination

VF-1: Dry lbs/sq.ft./h= 6.7 kg/m2/h 32.6 Filtrate Clarity Scale

0 Crystal Clear, zero suspended solids

Filtrate Capacity Determination 2 Somewhat transparent solution

3 Less cloudy, non-transparent solution

VF-1: Gal/sq.ft./h= 1.6 L/m2/h 64 4 Very cloudy discernible solid/liquid interface

5 Opaque, no solid/liquid interface visible

APPENDIX 4

MINERALOGICAL REPORT

APPENDIX 4

MINERALOGICAL REPORT

Microscopical Investigation of Two Auriferous Composite Samples from the Cruz de Mayo Project, Sonora, Mexico (INSPECTORATE - METALLURGICAL DIVISION) November 2011 (11-11/128)

Mineralogical Investigation of an Argentiferous
Leach Residue from the Santa Elena Mine, Sonora, Mexico

(INSPECTORATE - METALLURGICAL DIVISION)

November 2011 (11-11/130)

Summary

The following report presents the results of mineralogical investigations carried out on a silver-bearing leach residue from the Santa Elena Mine, Sonora, Mexico. The investigations comprised ore microscopy, X-ray diffractometry and microanalysis focussing on silver mineralogy and its potential deportment in a leaching process.

The leach residue consists mainly of silica and carbonate gangue (quartz and calcite). It carries rock fragments that show a distinct iron staining due to finely disseminated limonite. The general oxidation of the material is also reflected by the presence of manganese oxides. Sulphides do not exceed trace amounts. Specific silver minerals have not been detected.

Microanalysis proved a faint and erratic distribution of silver that could not be attributed to specific silver minerals or other silver carriers. It is assumed that the silver retained in the sample originates from reprecipitation of silver during the leaching process. Reprecipitation can be of the main reasons for poor silver recovery in leaching operations.

Resumen

El informe presenta los resultados de investigaciones mineralgicas realizadas en una muestra de residuo de lixiviacin argentfera de la mina de Santa Elena, Estado de Sonora, Mexico. Las investigaciones comprendieron mineragrafa, difractometra de rayos X y micronanlisis, enfocando la mineraloga y el comportamiento de plata.

El residuo est mayormente compuesto por cuarzo y calcita. Adems lleva fragmentos de roca que muestran una marcada coloracin ferruginosa causada por limonita diseminada. La oxidacin generalizada del material se manifiesta tambin en la presencia de xidos de manganeso. Los sulfuros no sobrepasan cantidades trazas. No se ha detectado minerales especficos de plata.

A travs de microanlisis se comprob una dbil y errtica distribucin de plata que no aparece con minerales especficos de plata u otros portadores de plata. Se supone que la plata retenida en la muestra resulta de una reprecipitacin de plata durante el proceso de lixiviacin. Reprecipitacin puede ser una de las causas principales para mermas de plata en operaciones de lixiviacin.

Mannheim,
November 22, 2011

R.W. Lehne

ORE MICROSCOPY (INSPECTORATE - METALLURGICAL DIVISION)
SAMPLE PROJECT
Ag [g/t]
Main Constituents ( > 5%) Accessories ( 5 - 1%) Traces ( < 1%)
Au [g/t]
C 10 1103507 (Santa Elena)
Leach Residue < 2 mm 1a
1d
1b
1c
3a
2a 2b
3b 3c
Apatite (apt)
Calcite
Scheelite (sch)
Silica Gangue (sil)
Fluorite (flu)
The sample is composed of predominant silica gangue and subordinate fluorite. Apatite
is a common accessory followed by calcite and minor scheelite. Traces are represented
by a series of minerals comprising sulphides, silicates and oxides. The presence of limonite
traces indicates a slight supergene alteration. Traces of graphite are restricted to shaly
rock particles that carry framboidal pyrite.
Just as in the previous sample rather coarse textures with relatively simple locking patterns
prevail in the ore. Scheelite is either unlocked (photo 6) or forms intergrowths with various
Continued on page 5.
x x x -- -- -- x -- --
20,71
4
Pyrite, Pyrrhotite
Rutile, Titanite
Hematite, Limonite
Ilmenite, Cassiterite (cas)
Zircon, Graphite
1,43
Ag Ag
250 microns 250 microns
MICROANALYSIS
The two element maps for silver show a discrete distribution of the metal. Silver has been detected in isolated spots. It cannot be attributed to specific
silver minerals or other silver carriers that could represent leaching-resistant minerals or remnants of an incomplete leaching process. In fact the faint and
erratic distribution of silver supports the idea of silver reprecipitation that occurred during the leaching process.

ORE MICROSCOPY (INSPECTORATE - METALLURGICAL DIVISION)
SAMPLE PROJECT
Ag [g/t]
Main Constituents ( > 5%) Accessories ( 5 - 1%) Traces ( < 1%)
Au [g/t]
C 10 1103507 (Santa Elena)
Leach Residue < 2 mm 1a
1d
1b
1c
3a
2a 2b
3b 3c
3
01-080-0743 (*) - Potassium Sodium Iron Aluminum Silicate Hydroxide - (K0.82Na0.1
01-089-4794 (A) - Fluorite - CaF2
01-086-1707 (*) - Calcium Aluminum Silicate - Ca(Al2Si2O8)
01-076-0823 (*) - Orthoclase - (K0.94Na0.06)(AlSi3O8)
00-041-0586 (*) - Ankerite - Ca(Fe+2,Mg)(CO3)2
00-005-0586 (*) - Calcite, syn - CaCO3
01-075-8322 (*) - Quartz - SiO2
File: D114301.raw - Laugungsrckstand
Sqrt (Counts)
0
1000
1e4
2e4
3e4
4e4
5e4
6e4
7e4
8e4
9e4
2-Theta - Scale
5 10 20 30 40 50 60 70
Estimated Phase Composition [wt.%]*)
Quartz 68,0
Calcite 25,0
Orthoclase 3,0
Fluorite 2,4
Ankerite 1,0
Anorthite < 1
Muscovite < 1
X-RAY DIFFRACTOMETRY
*) The phase composition represents a semiquantitative estimate based on data from
the ICDDs Powder Diffraction File and obtained by using a special software module
(Full Pattern Matching). While quartz and subordinate calcite a clearly the main
constituents, the given percentages for the listed minerals may vary considerably.

ORE MICROSCOPY (INSPECTORATE - METALLURGICAL DIVISION)
SAMPLE PROJECT
Ag [g/t]
Main Constituents ( > 5%) Accessories ( 5 - 1%) Traces ( < 1%)
Au [g/t]
C 10 1103507 (Santa Elena)
Leach Residue < 2 mm 1a
1d
1b
1c
3a
2a 2b
3b 3c
5
Ag Fe
MICROANALYIS
The third element map for silver has been obtained by employing two parallel spectrometers (elevated background
noise as compared to the previous maps). Nevertheless there are not more than three tiny spots where silver
has been clearly detected. The corresponding element maps for iron and silica show the distribution of the
two elements reflecting ferruginous rock fragments and quartz grains. In addition to the minerals already identified
by microscopy and XRD, microanalysis (EDS) detected dolomite, baryte, rutile, pyrophanite and zircon.

ORE MICROSCOPY (INSPECTORATE - METALLURGICAL DIVISION)
SAMPLE PROJECT
Ag [g/t]
Main Constituents ( > 5%) Accessories ( 5 - 1%) Traces ( < 1%)
Au [g/t]
C 10 1103507 (Santa Elena)
Leach Residue < 2 mm 1a
1d
1b
1c
3a
2a 2b
3b 3c
Apatite (apt)
Calcite
Scheelite (sch)
Silica Gangue (sil)
Fluorite (flu)
The sample is composed of predominant silica gangue and subordinate fluorite. Apatite
is a common accessory followed by calcite and minor scheelite. Traces are represented
by a series of minerals comprising sulphides, silicates and oxides. The presence of limonite
traces indicates a slight supergene alteration. Traces of graphite are restricted to shaly
rock particles that carry framboidal pyrite.
Just as in the previous sample rather coarse textures with relatively simple locking patterns
prevail in the ore. Scheelite is either unlocked (photo 6) or forms intergrowths with various
Continued on page 5.
x x x -- -- -- x -- --
20,71
4
Pyrite, Pyrrhotite
Rutile, Titanite
Hematite, Limonite
Ilmenite, Cassiterite (cas)
Zircon, Graphite
1,43
Ag Ag
250 microns 250 microns
MICROANALYSIS
The two element maps for silver show a discrete distribution of the metal. Silver has been detected in isolated spots. It cannot be attributed to specific
silver minerals or other silver carriers that could represent leaching-resistant minerals or remnants of an incomplete leaching process. In fact the faint and
erratic distribution of silver supports the idea of silver reprecipitation that occurred during the leaching process.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL 30, 2013 | RELEASED DATE: JULY 25, 2013

APPENDIX D

2012-2013 DRILL HOLE COLLAR LOCATIONS

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

APPENDIX D - 2012-2013 DRILL HOLE COLLAR LOCATIONS

	LOCATION_X	LOCATION_Y	LOCATION_Z	LENGTH	YEAR	DRILL TYPE
GT08-01	581296.36	3321255.30	782.00	300.54	2008	DD
GT08-02	581020.89	3321299.13	800.84	314.50	2008	DD
GT08-03	581579.70	3321391.77	802.02	301.90	2008	DD
GT08-04	581301.14	3321491.07	799.00	250.20	2008	DD
GT-12-09	581323.00	3321358.73	784.81	280.35	2012	DD
GT-12-10	581448.98	3321349.58	776.87	310.50	2012	DD
GT-12-11	581449.58	3321348.80	777.05	344.45	2012	DD
GT-12-12	581599.18	3321449.87	808.18	243.40	2012	DD
GT-12-13	581450.14	3321383.02	774.80	280.70	2012	DD
SE-06-01	580868.53	3321524.83	809.75	89.31	2006	DD
SE-06-02	580855.91	3321508.35	809.10	98.45	2006	DD
SE-06-03	580941.04	3321511.62	809.77	86.56	2006	DD
SE-06-04	580943.76	3321460.22	799.10	130.45	2006	DD
SE-06-05	581041.29	3321456.41	806.57	146.30	2006	DD
SE-06-06	581037.68	3321510.61	818.11	110.64	2006	DD
SE-06-07	580788.93	3321550.44	803.27	114.00	2006	DD
SE-06-09	581139.76	3321517.12	821.47	99.00	2006	DD
SE-06-10	581043.39	3321406.12	795.21	175.26	2006	DD
SE-06-11	581243.71	3321514.44	811.82	80.47	2006	DD
SE-06-12	581254.05	3321453.31	801.15	144.78	2006	DD
SE-06-13	581510.76	3321530.75	782.68	65.23	2006	DD
SE-06-14	581141.03	3321458.96	810.28	88.70	2006	DD
SE-06-15	581142.76	3321404.64	806.85	181.36	2006	DD
SE-06-16	580924.26	3321348.73	800.62	230.12	2006	DD
SE-06-17	581244.21	3321313.12	792.33	250.85	2006	DD
SE-06-18	581331.81	3321443.93	794.10	135.53	2006	DD
SE-06-19	580777.45	3321512.60	802.84	192.00	2006	DD
SE07-20	580966.47	3321092.87	777.17	491.00	2007	DD
SE07-21	581215.79	3321099.58	780.64	420.00	2007	DD
SE07-22	581204.11	3321452.92	800.11	139.00	2007	DD
SE07-23	581302.64	3321454.74	794.33	109.00	2007	DD
SE07-24	581374.71	3321451.92	786.92	100.00	2007	DD
SE07-25	581408.59	3321460.27	780.60	97.00	2007	DD
SE07-26	581463.17	3321437.00	776.40	83.55	2007	DD
SE07-27	580861.11	3321554.94	801.81	65.50	2007	DD
SE07-28	581202.27	3321498.26	810.00	78.10	2007	DD

1

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

LOCATION_X LOCATION_Y LOCATION_Z LENGTH YEAR DRILL TYPE

SE07-29	581291.70	3321533.95	804.85	106.00	2007	DD
SE07-30	581346.13	3321516.57	797.00	100.00	2007	DD
SE07-31	581395.77	3321529.00	792.00	100.00	2007	DD
SE07-32	581445.16	3321519.72	782.00	100.00	2007	DD
SE07-33	581411.22	3321390.20	775.00	185.00	2007	DD
SE07-34	581411.25	3321388.11	774.53	220.50	2007	DD
SE07-35	581462.96	3321437.02	776.42	125.00	2007	DD
SE07-36	581123.10	3321444.74	805.26	115.60	2007	DD
SE07-36B	581123.10	3321444.74	805.26	134.00	2007	DD
SE07-37	581192.10	3321391.99	798.32	151.00	2007	DD
SE07-38	581039.35	3321508.23	814.89	115.40	2007	DD
SE07-39	581254.79	3321406.42	791.14	152.00	2007	DD
SE07-40	581303.13	3321408.17	790.07	180.80	2007	DD
SE07-41	581343.75	3321408.38	787.06	170.00	2007	DD
SE07-42	580943.79	3321517.16	806.49	60.00	2007	DD
SE07-42B	580943.79	3321517.16	806.00	90.00	2007	DD
SE07-43	580903.04	3321517.11	806.69	48.38	2007	DD
SE07-43B	580903.04	3321517.11	806.69	56.00	2007	DD
SE07-44	581083.78	3321608.60	865.65	58.00	2007	DD
SE07-45	581083.86	3321610.87	865.69	50.00	2007	DD
SE07-46	581085.21	3321609.77	865.65	50.00	2007	DD
SE07-47	581084.38	3321610.51	865.66	50.70	2007	DD
SE07-48	581134.87	3321611.11	859.09	50.00	2007	DD
SE07-49	581136.92	3321611.02	859.05	52.00	2007	DD
SE07-50	581136.21	3321611.63	859.10	50.00	2007	DD
SE07-51	581136.26	3321610.26	859.09	50.00	2007	DD
SE07-52	580913.91	3321629.70	814.18	58.00	2007	DD
SE07-53	580990.29	3321544.51	817.87	58.50	2007	DD
SE07-54	580990.98	3321544.35	817.86	31.70	2007	DD
SE07-55	580990.44	3321545.02	817.92	26.80	2007	DD
SE07-56	580991.08	3321544.98	817.89	75.10	2007	DD
SE07-57	581188.12	3321558.21	827.86	31.80	2007	DD
SE07-58	581189.19	3321558.42	827.85	52.50	2007	DD
SE07-59	581188.83	3321559.76	827.88	50.00	2007	DD
SE07-60	581187.95	3321559.23	827.89	46.50	2007	DD
SE07-61	581189.33	3321559.40	827.86	61.00	2007	DD
SE07-62	581523.02	3321506.38	778.40	116.40	2007	DD
SE07-63	581655.55	3321580.27	782.90	197.70	2007	DD

2

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

LOCATION_X LOCATION_Y LOCATION_Z LENGTH YEAR DRILL TYPE

SE07-64	581655.79	3321579.18	782.77	122.00	2007	DD
SE-08-68	581453.42	3321266.65	777.11	352.30	2008	DD
SE-08-69	581420.80	3321105.11	771.04	520.70	2008	DD
SE-08-70	581453.24	3321269.19	777.04	285.00	2008	DD
SE-08-71	581037.72	3321648.82	842.55	219.90	2008	DD
SE-08-72	581600.63	3321388.66	807.54	292.20	2008	DD
SE-08-73	581356.49	3321312.44	783.01	263.00	2008	DD
SE-08-75	581483.52	3321384.19	779.90	239.00	2008	DD
SE-08-79	581608.04	3321467.37	809.78	252.10	2008	DD
SE-08-80	581607.06	3321300.87	802.94	363.40	2008	DD
SE-08-81	581711.42	3321449.82	815.75	300.60	2008	DD
SE-08-82	581791.94	3321477.93	824.93	268.50	2008	DD
SE-08-83	581783.88	3321344.38	872.43	462.50	2008	DD
SE-08-84	581304.27	3321216.70	781.13	339.30	2008	DD
SE-08-85	581186.70	3321210.10	786.70	337.20	2008	DD
SE-08-86	581105.14	3321219.57	789.11	395.30	2008	DD
SE-08-87	580977.73	3321243.95	787.27	287.40	2008	DD
SE-08-88	580885.80	3321214.10	786.06	364.40	2008	DD
SE-08-91	581593.93	3321210.17	818.53	498.00	2008	DD
SE-08-93	581301.57	3321313.27	784.65	190.95	2008	DD
SE-08-95	581640.60	3321200.59	828.35	452.50	2008	DD
SE-08-96	580719.83	3321593.42	793.03	110.00	2008	DD
SE-12-39	581450.64	3321386.59	774.91	191.00	2012	DD
SE-12-40	581451.64	3321386.53	774.96	203.00	2012	DD
SE-12-41	581452.16	3321386.65	774.96	234.00	2012	DD
SE-12-42	581515.32	3321508.28	777.42	118.25	2012	DD
SE-12-43	581541.94	3321520.13	778.29	111.95	2012	DD
SE-12-44	581541.85	3321519.04	778.31	130.85	2012	DD
SE-12-45	581524.67	3321403.89	789.04	250.45	2012	DD
SE-12-46	581524.65	3321404.32	789.04	201.70	2012	DD
SE-12-47	581525.48	3321403.16	789.35	246.80	2012	DD
SE-12-49	581607.86	3321466.93	808.51	212.00	2012	DD
SE-12-50	581607.93	3321466.44	808.57	278.00	2012	DD
SE-12-51	581608.06	3321467.02	808.61	275.00	2012	DD
SE-12-52	581608.28	3321467.51	808.64	212.00	2012	DD
SE-12-53	581512.52	3321507.69	777.40	126.85	2012	DD
SE-12-54	581516.82	3321507.34	777.44	155.05	2012	DD
SE-12-55	581525.90	3321398.44	788.87	224.90	2012	DD

3

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

LOCATION_X LOCATION_Y LOCATION_Z LENGTH YEAR DRILL TYPE

SE-12-56	581525.30	3321398.78	788.85	239.95	2012	DD
SE-12-57	581525.87	3321402.93	789.04	274.10	2012	DD
SE-12-59	581735.74	3321183.20	853.19	424.90	2012	DD
SE-12-62	581640.60	3321200.59	828.00	443.00	2012	DD
SE-12-63	581649.79	3321451.13	811.82	299.00	2012	DD
SE-12-64	581580.13	3321393.08	800.85	401.00	2012	DD
SE-12-65	581579.66	3321393.52	800.94	275.00	2012	DD
SE-12-66	581579.64	3321392.92	800.80	359.00	2012	DD
SE-12-67	581579.78	3321393.18	800.87	255.00	2012	DD
SE-12-70	581579.13	3321393.18	800.78	302.00	2012	DD
SE-12-71	581578.29	3321392.72	800.72	332.00	2012	DD
SE-12-72	581736.14	3321182.84	853.28	600.00	2012	DD
SE-12-73	581735.77	3321183.05	853.18	540.00	2012	DD
SE-12-74	581735.95	3321183.01	853.25	615.00	2012	DD
SE-12-75	581735.95	3321183.24	853.35	500.60	2012	DD
SE-12-76	581606.63	3321301.43	802.08	358.15	2012	DD
SE-12-77	581605.14	3321300.85	802.17	392.70	2012	DD
SE-12-78	581605.19	3321301.47	802.21	356.00	2012	DD
SE-12-79	581604.85	3321301.86	802.25	351.70	2012	DD
SE-12-80	581648.74	3321451.10	811.72	290.00	2012	DD
SE-12-81	581604.21	3321300.23	802.13	417.30	2012	DD
SE-12-82	581640.60	3321200.59	828.00	473.00	2012	DD
SE-12-83	581586.51	3321217.78	817.52	426.15	2012	DD
SE-12-84	581640.60	3321200.59	828.00	455.00	2012	DD
SE-12-85	581586.31	3321216.40	817.46	441.60	2012	DD
SE-12-86	581640.60	3321200.59	828.00	425.00	2012	DD
SE-12-87	581493.81	3321324.89	786.15	402.05	2012	DD
SE-12-89	581640.60	3321200.59	828.00	479.00	2012	DD
SE-12-90	581323.00	3321358.73	784.00	218.78	2012	DD
SE-12-91	581640.60	3321200.59	828.00	455.00	2012	DD
SE-12-92	581526.29	3321403.19	789.23	305.00	2012	DD
SE-12-93	581525.84	3321403.94	789.13	311.05	2012	DD
SE-12-94	581448.34	3321301.38	775.97	337.40	2012	DD
SE-12-95	581447.47	3321301.13	775.90	316.60	2012	DD
SE-12-96	581449.20	3321302.62	776.11	317.00	2012	DD
SE-12-102	581449.43	3321302.96	776.06	401.90	2012	DD
SE-12-103	581323.00	3321358.73	784.00	270.90	2012	DD
SE-12-104	581323.00	3321358.73	784.00	277.00	2012	DD

4

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

LOCATION_X LOCATION_Y LOCATION_Z LENGTH YEAR DRILL TYPE

SE-12-105	581380.18	3321362.61	782.21	193.95	2012	DD
SE-12-106	581379.96	3321361.75	782.21	240.00	2012	DD
SE-12-107	581649.25	3321451.36	811.73	239.00	2012	DD
SE-12-108	581649.41	3321450.83	811.74	254.00	2012	DD
SE-12-109	581649.30	3321451.13	811.86	287.00	2012	DD
SE-12-110	581648.94	3321451.38	811.74	263.00	2012	DD
SE-12-111	581736.16	3321183.09	853.33	527.60	2012	DD
SE-12-112	581640.60	3321200.59	828.00	450.00	2012	DD
SE-12-113	581379.88	3321361.42	782.34	267.85	2012	DD
SE-12-114	581379.16	3321359.94	782.27	289.20	2012	DD
SE-13-60	581640.60	3321200.59	828.00	474.00	2013	DD
SE-13-61	581640.60	3321200.59	828.00	494.00	2013	DD
SE-13-115-2	581301.57	3321313.30	784.50	384.00	2013	DD
SE-13-116	581734.30	3321182.34	853.00	491.60	2013	DD
SE-13-117	581301.57	3321313.30	784.50	291.90	2013	DD
SE-13-118	581301.57	3321313.30	784.50	345.10	2013	DD
SE-13-119	581301.57	3321313.30	784.50	383.95	2013	DD
SE-13-120	581640.60	3321200.59	828.00	460.00	2013	DD
SE-13-121	581734.30	3321182.34	853.00	495.00	2013	DD
SE-13-122	581301.60	3321313.30	784.50	301.55	2013	DD
SE-13-123	581640.60	3321200.59	828.00	470.00	2013	DD
SE-13-124	581301.60	3321313.30	784.50	432.75	2013	DD
SE-13-125	581734.30	3321182.34	853.00	522.00	2013	DD
SE-13-126	581640.60	3321200.59	828.00	492.40	2013	DD
SE-13-127	581225.93	3321364.26	791.07	350.30	2013	DD
SE-13-128	581734.30	3321182.34	853.00	551.80	2013	DD
SE-13-129	581225.93	3321364.26	791.07	396.30	2013	DD
SE-13-130	581640.60	3321200.59	828.00	452.00	2013	DD
SE-13-131	581226.97	3321365.13	795.05	301.60	2013	DD
SE-13-132	581640.60	3321200.59	828.00	310.50	2013	DD
SE-13-133	581734.30	3321182.34	853.00	500.80	2013	DD
SE-13-134	581640.60	3321200.59	828.00	502.00	2013	DD
SE-13-135	580964.57	3321310.84	795.05	332.05	2013	DD
SE-13-136	581734.30	3321182.34	853.00	501.00	2013	DD
SE-13-137	580964.57	3321310.80	795.00	337.85	2013	DD
SE-13-138	581735.77	3321183.00	853.18	497.00	2013	DD
SE-13-139	580964.57	3321310.80	795.05	331.90	2013	DD
SE-13-140	580964.57	3321310.80	795.05	316.30	2013	DD

5

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

LOCATION_X LOCATION_Y LOCATION_Z LENGTH YEAR DRILL TYPE

SE-13-141	581735.77	3321183.00	853.18	501.00	2013	DD
SE-13-142	580964.57	3321310.80	795.05	347.00	2013	DD
SE-13-144	581735.77	3321183.00	853.18	519.00	2013	DD
SE-13-145	580964.57	3321310.80	795.05	350.25	2013	DD
SE-13-152	581735.77	3321183.00	853.18	531.00	2013	DD
SERC09-22	581150.28	3321556.71	823.24	76.00	2009	RC
SERC09-23	581111.56	3321538.91	818.96	53.00	2009	RC
SERC09-24	581111.46	3321540.00	819.04	58.00	2009	RC
SERC09-25	581069.88	3321530.86	816.01	62.00	2009	RC
SERC09-26	581069.92	3321532.26	816.15	58.00	2009	RC
SERC09-27	581036.90	3321554.54	813.94	40.00	2009	RC
SERC09-28	581038.63	3321554.02	813.90	42.67	2009	RC
SERC09-29	580948.33	3321550.47	811.66	40.00	2009	RC
SERC09-30	580948.38	3321548.64	811.77	52.00	2009	RC
SERC09-31	580900.73	3321542.37	807.07	45.72	2009	RC
SERC09-32	580900.66	3321540.63	806.96	52.00	2009	RC
SERC09-33	581124.39	3321679.65	850.46	150.00	2009	RC
SERC09-36	580965.12	3321595.37	835.92	100.00	2009	RC
SERC09-37	580964.10	3321597.21	835.88	76.00	2009	RC
SERC09-38	581017.75	3321601.58	852.45	74.00	2009	RC
SERC09-39	581016.82	3321604.00	852.51	100.00	2009	RC
SERC09-40	580743.15	3321634.44	806.02	34.00	2009	RC
SERC09-41	580741.76	3321633.05	806.17	40.00	2009	RC
SERC-12-22	581734.13	3321182.89	853.34	602.00	2012	RC/DD
SERC-12-23	581734.30	3321182.45	853.50	596.00	2012	RC/DD
SERC-12-24	581735.02	3321184.49	853.41	602.00	2012	RC/DD
SERC-12-25	581740.24	3321126.90	858.40	532.00	2012	RC/DD
SERC-12-26	581739.77	3321123.90	859.27	542.00	2012	RC/DD
SERC-12-27	581449.94	3321300.20	776.36	486.50	2012	RC/DD
SERC-12-28	581448.74	3321300.04	775.95	465.40	2012	RC/DD
SERC-12-29	581451.52	3321299.83	776.28	385.00	2012	RC/DD
SERC-12-30	581450.06	3321452.45	780.37	120.00	2012	RC
SERC-12-31	581550.17	3321350.84	790.45	326.00	2012	RC/DD
SERC-12-32	581551.57	3321351.80	790.34	428.00	2012	RC/DD
SERC-12-33	581548.55	3321351.81	790.33	350.00	2012	RC/DD
SERC-12-34	581650.09	3321453.93	811.81	344.70	2012	RC/DD
SERC-12-35	581599.72	3321454.78	808.25	209.15	2012	RC/DD
SERC-12-36	581225.97	3321364.24	791.13	180.00	2012	RC

6

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EBA File: V15101045 | July 2013

LOCATION_X LOCATION_Y LOCATION_Z LENGTH YEAR DRILL TYPE

SERC-12-37	581068.99	3321366.63	807.52	204.00	2012	RC
SERC-12-38	581169.02	3321362.98	803.66	186.00	2012	RC
SERC-12-58	581734.47	3321181.04	853.51	517.35	2012	RC/DD
SERC-12-88	581640.60	3321200.59	828.00	500.00	2012	DD
SERCH08-01	581304.27	3321216.70	781.13	204.00	2008	RC
SERCH08-02	581186.70	3321210.10	786.70	234.00	2008	RC
SERCH08-03	581105.14	3321219.57	789.11	237.00	2008	RC
SERCH08-04	580977.73	3321243.95	787.27	225.00	2008	RC
SERCH08-05	580885.80	3321214.10	786.06	237.00	2008	RC
SERCH08-15	581593.93	3321210.17	818.53	254.00	2008	RC
SERCH08-18	581301.57	3321313.27	784.65	163.00	2008	RC
SERCH08-19	580838.04	3321586.49	800.83	175.00	2008	RC
SERCH08-20	580909.42	3321564.37	816.79	10.00	2008	RC
SERCH08-21	581408.64	3321193.75	773.22	151.00	2008	RC
SERCP-11A	581354.15	3321483.88	769.96	50.00	2012	RC
SERCP-11A-TWIN	581354.55	3321483.58	765.09	56.90	2012	RC
SERCP-11V	581354.47	3321481.64	770.00	85.00	2012	RC
SERCP12-A	581100.00	3321494.66	770.00	66.50	2012	RC
SERCP-1A	580896.15	3321516.18	764.53	45.00	2012	RC
SERCP-2A	580990.37	3321492.57	769.00	70.00	2012	RC
SERCP-2V	580990.65	3321491.37	769.99	96.00	2012	RC
SERCP-4A	581100.02	3321494.66	769.85	62.00	2012	RC
SERCP-4A-TWIN	581100.99	3321501.97	765.29	66.50	2012	RC
SERCP-4V	581099.97	3321492.21	769.98	84.00	2012	RC
SERCP-5A	581134.56	3321492.38	769.98	62.00	2012	RC
SERCP-5V	581133.92	3321489.91	770.08	90.00	2012	RC
SERCP-7A	581226.72	3321486.91	770.24	102.00	2012	RC
SERCP-7V	581226.74	3321480.57	770.25	102.00	2012	RC
SERCP-9A	581283.64	3321483.25	770.07	55.00	2012	RC
SERCP-9V	581283.74	3321481.34	770.10	108.00	2012	RC

7

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL 30, 2013 | RELEASED DATE: JULY 25, 2013

APPENDIX E

GEOTECHNICAL DATA

July 11, 2013	ISSUED FOR USE
	EBA FILE: V15101045-011
SILVERCREST MINES INC.	Via Email: eric@silvercrestmines.com
570 Granville Street, Suite 501
Vancouver, BC  V6C 3P1

Attention:	Eric Fier, Brent McFarlane
Subject:	Santa Elena Expansion Plan,  Underground Pre-Feasibility Level Geotechnical Rock Mechanics Evaluation

1.0   INTRODUCTION

1.1   General

EBA Engineering Consultants Ltd. operating as EBA, A Tetra Tech Company (EBA) was requested by SilverCrest Minerals Inc. (SilverCrest) to assess the geotechnical conditions at the Santa Elena mine as part of a Prefeasibility Study. The following report summarizes the geotechnical data obtained and design parameters implemented and focusses on both the footwall development and production stopes of the Santa Elena Mine. Geotechnical development design and production design are summarized individually in this report.

The Santa Elena Gold and Silver Mine property is approximately 150 km northeast of the state capital city of Hermosillo, Sonora, Mexico. Currently, ore is extracted from an open pit mine, and gold and silver are recovered by heap leaching. Before SilverCrest became involved with the property, ore was extracted underground, and there are several open stopes, a mine shaft, and an open mine portal.

Geotechnical data collection consisted of a background review of information from Vector Engineering (Vector 2009), a geotechnical drilling program undertaken in 2011 where six geotechnical boreholes were drilled to assess conditions for permanent ramp and raise development, and a geotechnical drilling program undertaken in 2012 to assess the orebody, footwall and hangingwall conditions for production design. The 2012 program also consisted of UCS, Brazilian and Point Load Testing of rock samples at the Sonora University.

1.2   Report Organization

This report is organized into six parts:

1.   Geotechnical assessment of permanent ramps and raises (Sections 2 and 3);

2.   Stability assessment and support for permanent ramps and raises (Sections 4 and 5);  

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE  

3.   Geotechnical analysis of production stope design (Section 6);

4.   Geotechnical analysis of pillar design (Section 7);

5.   Stope design summary (Section 8); and

6.   Recommendations for further geotechnical work at feasibility level (Section 9).

1.3   Investigation and Assessment Objectives and Methodology

The following are the objectives of the pre-feasibility geotechnical evaluation for the Santa Elena Expansion Plan:

§   Collect geotechnical information pertaining to the in situ materials;

§   Geotechnically characterize the in situ materials;

§   Geotechnically characterize the rock mass;

§   Divide the mine into zones, if deemed necessary, to serve as the basis of geotechnical analysis; and

§   Assess the stability of the underground openings (ramps, raises, and stopes) to make recommendations pertaining to permanent ramp and raise support and stope and pillar dimensions.

The principal stages of the pre-feasibility geotechnical evaluation work program were split into two sections and comprised the following:

Section 1 – Permanent ramp access and raise assessment utilizing data from the 2011 drilling program:

§   Conduct QA/QC of geotechnical core logging and discontinuity orientation of core recovered from the boreholes;

§   Interpret and analyze the collected geotechnical data to divide the ramp and raise development into geotechnical domains; and

§   Compile a pre-feasibility geotechnical evaluation incorporating recommendations for ramp and raise support design.

Section 2 – Stope and pillar design utilizing data from the 2012 drilling program:

§   Conduct QA/QC of geotechnical core logging and discontinuity orientation of core recovered from the boreholes;

§   Select representative drill core samples from the different lithological units encountered in the boreholes to undertake field and laboratory testing of geotechnical properties of the in situ materials;

§   Submit representative samples to the Sonora University rock mechanics laboratories, Mexico, for geotechnical testing;

§   Interpret and analyze the collected geotechnical data and test results to divide the rockmass into geotechnical domains; and

2

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§   Compile a pre-feasibility geotechnical evaluation incorporating recommendations for stope and pillar design.

2.0   GEOTECHNICAL ASSESSMENT OF FOOTWALL ACCESS RAMPS AND RAISES

Initial ramp development and ventilation raise location was considered for the eastern end of the open pit only, however, this has subsequently been revised. The main access ramp which is currently under development is located at the eastern end of the open pit. This eastern ramp is currently at the 655 Level and will continue down to the base of the orebody at the 328 Level. Two shaft raises will be used at this eastern end of the orebody for ventilation purposes. A future western decline is proposed at the base of open pit together with a separate ventilation shaft. This western shaft and ramp will extend down to the base of the current orebody in this area of the orebody at the 471 Level. Figure 1 shows an overview of the ultimate pit, underground workings and borehole locations drilled in 2011, and the perceived location of the Tinaja Fault Zone (located based on topography). The proposed underground workings are within the footwall of the deposit where the predominant rock types are Andesite and Rhyolite. The workings can be separated into two distinct zones, as described below.

Zone 1 – Eastern Ramp

The Eastern ramp spirals down from surface to the base of the current orebody at 328 Level. Figure 2 shows the eastern ramp and the relevant boreholes that have been drilled specifically to characterize the rock. The gradient of the ramp is 12%. Boreholes GT11-08, GT11-05, GT-06 and GT-07 intersect the ramp.

The Tinaja Fault is anticipated to intersect the Eastern Ramp. Borehole GT11-06 intersects the Tinaja Fault. Boreholes GT11-06a and GT11-06b intersect the Tinaja Fault at a higher elevation. These holes were planned to be drilled to completely intersect the Tinaja Fault, but due to surface conditions were moved by SilverCrest. The perceived fault is shown in Figure 1.

Adjacent to the eastern ramp is an as-built fresh air raise that connects to each of the different levels to provide ventilation during development of the ramp access. This raise is shown on Figure 2. The proposed main ventilation shaft at the eastern end of the orebody is also shown on Figure 2 and will extend down to the base of the proposed workings.

Zone 2 – Western Ramp and Ventilation Raise

The western ramp is proposed at the base of the open pit together with a separate ventilation shaft as shown in Figure 1. This development is a recent addition and was proposed after initial geotechnical planning and investigation which focussed on the eastern side of the open pit. Therefore, this ramp and raise are not considered in detail within this report. Whilst it is anticipated that geotechnical conditions are similar throughout the footwall it is recommended to undertake geotechnical drilling at this proposed location in order to ascertain ground support recommendations for this access and raise development.

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Note on Dip and Dip Direction

Within this report a structural orientation has been designated a dip and dip direction. For example 40°/ 090° indicates a structure inclined at 40 degrees from the horizontal trending towards the east.

2.1   Background Review

As part of the review of background information, the report prepared by Vector Engineering (Vector 2009) was reviewed. The pertinent information from this report is summarized below.

The primary rock types observed on the property are Tertiary Andesite and Rhyolite flows.

Near quartz veining argillic and silicification is present, whereas in the Andesite beds chloritic alteration increases away from the mineralized zones.

§   The mineralized zone is approximately 1.2 km long, with a width from 1 to 35 m, averaging 15 m.

§   The trend is between 40° and 60° towards the south.

§   Within the footwall Andesite, the primary joint set is oriented at 28/085.

§   The average geomechanical and geotechnical properties of the Andesite as reported by Vector are:

-   Strength 80 to 160 MPa;

-   Average RQD = 65%;

-   Average joint spacing = 0.25 m;

-   Average RMR joint condition = 10 (out of a possible 30); and

-   Average RMR is 51 (fair rock).

§   A stereonet of the discontinuities measured underground indicates two major joint sets one at 65/170, the second at 74/258. These measurements are consistent with surface recordings and measurements undertaken from oriented core.

2.2   Geotechnical Investigation

SilverCrest geologists undertook diamond drilling and geotechnical borehole logging during September and October 2011 to gain geotechnical data for the ramp. EBA was on site before the start of the drilling program to train SilverCrest geologists on evaluating geotechnical parameters. Field data collected by SilverCrest geologists was reviewed on site by the senior mine geologist as the drilling progressed.

The drilling program consisted of drilling, logging and sampling six boreholes. EBA in conjunction with SilverCrest chose the location, length, azimuth, and inclination of the boreholes. EBA recommended that all six boreholes have oriented core to allow kinematic analysis, and the Reflex® ACT II orientation system was used to orient the drill core.

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Drill core from the geotechnical holes was oriented using the Reflex ACT II RD core orientation system. With the ACT II tool, the core is marked at this lowermost point from which a bottom reference line can be drawn.

The core is carefully pieced back together, and the reference line is transferred along the length of the core run. Alpha and beta angles are measured and recorded for the dominant and sub-dominant joint sets for each core run. The alpha angle is the acute angle between the core axis and the plane of the joint surface, while the beta angle is the circumferential angle between the Reflex ACT II RD reference line and the downhole apex of the ellipse formed by the joint cutting through the core, measured in a “looking downhole” clockwise direction. Alpha angles are measured using a clear plastic protractor held parallel to the axis of the core, and beta angles are measured using a wraparound protractor scaled for the core diameter.

Oriented core data was processed using Rocscience DIPS™ software. Each alpha and beta angle pair, in conjunction with downhole survey data was used to determine a true dip and dip direction for dominant and subdominant joint surfaces that were observed in core logging.

Geotechnical parameters for rock mass classification were collected using two widely used empirical systems: the NGI Q-system after Barton et al. (1974) and the Rock Mass Rating (RMR) system after Bieniawski (1973). Sections 2.5.1 and 2.5.2 present an overview of the Q and RMR systems respectively.

Table 1 presents the location of the six boreholes for footwall development.

Table 1 - Location of 2011 Geotechnical Boreholes
Hole ID	Easting (m)	Northing (m)	Elevation at ground surface (m)	Length (m)	Azimuth (bearing degrees)	Plunge (degrees)
GT11-05	581469.76	3321557.72	785.03	100.0	022	-70
GT11-06	581575.03	3321577.82	778.78	40.6	315	-70
GT11-06-A	581575.47	3321577.04	778.67	114.7	315	-70
GT11-06-B	581577.93	3321572.17	778.67	299.5	315	-60
GT11-07	581650.00	3321630.50	782.17	449.5	180	-60
GT11-08	581491.93	3321653.34	800.05	301.3	165	-60

2.3   Quality Assurance of Geotechnical Logs and Resolution of Ambiguities

A quality assurance (QA) review of the geotechnical core logs was undertaken to identify any potential issues or concerns with the dataset. The review of the logs included a spot check of the core photos to cross-check data. A few minor inconsistencies and issues were identified, but as a whole the QA review indicates that the dataset is in good order.

Some data was missing from the logs, including joint parameters as well as oriented core orientation angles. Q and RMR values were only calculated in runs for which the necessary data was available and complete.

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Two issues were identified with respect to the recorded RQD values. First, in selected runs, the RQD exceeded 75%, yet the number of natural fractures and Jn value recorded were high, indicating the opposite may be true. Second, and more commonly, in some runs, only one or two natural fractures were recorded, but the RQD length was lower than is potentially feasible. These runs typically listed 100% recovery over a 3 m run length, but with only one natural fracture, and yet had an RQD length of less than 2.90 m. In these cases, the correct procedures for measuring RQD may not have been followed. The RQD was still within the correct range of 80-100%, but the RQD was perhaps slightly lower than it should be based on fracture data. Potential issues in the RQD length are highlighted in 51 out of the 490 runs (10.4%). Due to the limited number of highlighted runs and since the RQD was still within the correct range, this issue did not warrant re-logging of the core, and no changes were made to this part of the data.

As a whole the Ja values were noted as being slightly lower than expected given the lithologies and alteration styles encountered. In some runs, the comments section notes the presence of chlorite and other non-softening mineral coatings, but the Ja values of 1 or 2 do not reflect this. A spot check of the photos suggests that the mineral coatings may be minor. The runs in which this concern is noted are irregular and interspersed and not deemed significant in quantity.

In the lower six runs from 28.75 to 39.05 m of borehole GT-11-06, the values of Ja for the representative and worst case joints are reversed. This is simply noted as a typo and the Ja values have been interchanged for the purpose of EBA’s analyses.

Some discrepancies exist between the values recorded and the associated comments, particularly as it relates to the assignment of a weathering value. This was noted in some runs; however, as the weathering value does not influence the Q or RMR calculation, no changes were made to this part of the dataset.

A nominal RQD of 10% was assigned to all runs for which the listed RQD was <10%. This change is consistent with the guidelines presented by Barton et al. (1974) in assessing the NGI Q-value.

2.4   Assessment of Potential Structural Domains

The main ramp and fresh air raise developments are within the footwall of the deposit which is formed from both Rhyolite and Andesite rock types. Within these rock types, differing structural domains could exist. A structural domain is classified as a zone or area of rock, which has similar structural characteristics and geomechanical properties, and consequently, the rock behaviour is also similar in terms of excavation (potential failure mechanisms and deformation).

The Tinaja Fault Zone is thought to be a 100 m wide strip of rock striking northeast to southwest. The location of the TFZ is based on surface topography where a shallow valley feature has formed. A review of the borehole logs was unable to definitively locate the Tinaja Fault Zone; therefore, this feature may not be as significant as previously suggested from surface mapping.

Based on the foregoing, there are potentially three structural domains on site:

1.           Eastern ramp in Rhyolite;

2.           Eastern ramp in Andesite; and

3.           The Tinaja Fault Zone.

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To reconcile whether these three areas form structural domains, the structural discontinuity orientations, as well as the average geotechnical parameters, were compared within each of the zones.

The alpha and beta angles recorded on the geotechnical logs were converted to true dip and dip directions using the program DIPS (by Rocscience 2011) and the orientation of the borehole at the collar.

Deviation by downhole survey of the borehole was not undertaken, therefore, downhole orientation was assumed to be a straight path from the hole collar. In total, EBA recorded 114 discontinuity measurements. This is an unusually low success rate for the orientation tool. By comparing all the data, four general structural joint sets were identified, as shown on Figure 3.

The average geotechnical parameters were compared and showed that overall geotechnical characteristics are very similar throughout the footwall of the mine. Therefore, at this stage, we do not believe there is sufficient data, or the basis, to discretize the rock mass into structural domains.

2.5   Summary of Rock Mass Quality and Geotechnical Conditions

2.5.1   Note on the NGI Q-System

The Q-system uses six parameters to determine a numerical value (Q) which represents the quality of a given rock mass. The parameters assign numerical values to various features of the rock mass, including RQD, number of joint sets, joint alteration, joint surface roughness, stress conditions, and water inflow intensity. These are combined in a simple formula to derive a “Q-value”, which indicates the overall quality of the rock mass for underground excavation and support design. This formula can be thought of as the function of three geomechanical rock properties (1) block size given by RQD/Jn; (2) inter-block shear strength given by Jr/Ja; and (3A)A active stress given by Jw/SRF.  The formula for determining Q is as follows:

                                                                       (1)

Where:                   RQD = rock quality designation

Jn = joint set number

Jr= joint roughness number

Ja = joint alteration number

Jw= joint water reduction number

SRF = stress reduction factor

The description of the six parameters can be found in the report by Barton et al. (1974). The numerical value of Q ranges from 0.001 (for exceptionally poor quality, squeezing ground) to 1000 (for exceptionally good quality, massive rock).

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The Q-system presents a relationship between the rock mass quality and the amount and type of permanent excavation support. This relationship is summarized on the NGI Q-system excavation or tunnel support chart. An updated version of the tunnel support chart is shown on Figure 4. The tunnel support chart takes into account not only the rock mass quality, but also the size of the excavation in terms of span or wall height and the purpose of the excavation. The size and purpose of the excavation are incorporated through a parameter known as the equivalent dimension (De). The equivalent dimension is determined by dividing the span (when analyzing for roof support) or height (when analyzing for wall support) of the excavation by the excavation support ratio (ESR). The ESR is a value, determined by the function and purpose of the excavation, and it is akin to a factor of safety. The ESR is higher for some excavations (e.g., ESR = 1.6 for permanent mine openings) and lower for more critical structures (e.g., ESR = 0.8 for underground nuclear power stations).

For the underground ramps at Santa Elena, the excavation dimensions are a span of 4.5 m wide by a vertical height of 4 m. In this case, the span is the controlling factor. A span of 4.5 m divided by an ESR of 1.6 is equal to an equivalent dimension of 2.8. This equivalent dimension value is shown as a green line on Figure 4.

2.5.2   Note on the Rock Mass Rating (RMR) System

The Rock Mass Rating (RMR) system is a geomechanical classification system for rocks, developed by Bieniawski (1973, with major updates in 1976 and 1989). The following six parameters are used to classify a rock mass using the RMR system:

1.   Uniaxial compressive strength of rock material;

2.   Rock Quality Designation (RQD);

3.   Spacing of discontinuities;

4.   Condition of discontinuities;

5.   Groundwater conditions;

6.   Orientation of discontinuities.

Each of the six parameters is assigned a value corresponding to the characteristics of the rock. The sum of the six parameters is the “RMR value”.

2.6   Assumptions for Evaluating Rock Mass Quality

Most of the parameters for evaluating the RMR and NGI Q-Values were collected on site. However, some assumptions were made during the evaluation specifically related to groundwater and stress conditions.

2.6.1   Groundwater

The groundwater table at Santa Elena is reported to be 60 m below ground surface. In the underground excavations, water inflow will be controlled by hydrogeological conduits such as fault zones, open joints, and fractured rock. Hydro-geological packer testing to determine the hydraulic conductivity was undertaken as part of the geotechnical investigation. The average of 12 tests successfully undertaken gave 1.3 x 10-7 m/s. The value of 1.3 x 10-7 m/s can be described as “low discharge or poor drainage” based on a qualitative comparison presented by Hoek and Bray (1977).

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Although open to conjecture, in finer grained igneous rocks, the somewhat brittle nature of the rock tends to produce a lower persistence of joints, but with a higher fracturation than other coarser igneous rocks. For a rock mass to be permeable, both conditions (i.e., fractured rock and highly persistent joint sets) are potentially required to increase the inter-connectivity. Other aspects that control permeability are joint spacing and the aperture of joints. The average joint spacing is moderate, and hardly any joints were noted as being open on the logs.

At this stage, the groundwater is assumed not to be significant during excavation and the water inflows that do occur should dissipate with time. For the RMR calculation, a groundwater condition of 10 was assumed (equivalent to < 10 litres of water inflow for a 10 m length of tunnel). For the NGI Q-system, a JW value of 1.0 was assumed (equivalent to a dry excavation or minor inflow). These inflow estimates have been verified by the Global Resource Engineering (November 2011) report which estimates inflow for the ramp as between 17.4 and 59.4 litres per minute of water inflow.

We understand that for the full feasibility study a groundwater model may be developed to more accurately predict groundwater conditions. The assessment of assumed groundwater parameters used in the RMR and Q-value analysis should be updated following this study.

2.6.2   Stress Reduction Factor

The stress reduction factor (SRF) categorizes the stress conditions into four categories: a) weakness zones, b) competent rock stress problems, c) squeezing rock, and, d) swelling rock. At this stage only competent rock stress problems are envisaged. One method of estimating SRF is by σc / σ1 (the uniaxial compressive strength (UCS) of the rock divided by the principal stress).

Based on the average description of strong rock (indicating a UCS of ~75 MPa) a medium stress (under SRF category b) is envisaged in the eastern ramp and eastern raise bore. A medium stress condition is based on the σc / σ1 ratio ranging between 10 and 200. A medium stress condition equates to a SRF value of 1.

We have also assumed an SRF of 1 for the lower eastern ramp. Although in theory the in-situ stress and average rock strength gives a σc / σ1 ratio lower than 10 in some instances, given only one borehole for the lower eastern ramp and the number of unknown or assumed parameters, this SRF is believed to be reasonable at this time. If stress problems occur in the lower levels after the eastern ramp is excavated, the SRF should be re-evaluated in conjunction with additional geotechnical boreholes drilled into the eastern ramp and geomechanical testing (including rock strength testing).

3.0   SUMMARY OF GEOTECHNICAL PARAMETERS

The geotechnical parameters throughout the mine were assessed based on the borehole logs. To achieve this aim, the exploration declines were split into four zones. Boreholes relevant to each zone were then grouped, and a weighted average was used to determine the average parameter value.

3.1   Eastern Ramp and Fresh Air Raise

EBA visited the site in 2012 and undertook geotechnical mapping of the first 400 m of ramp development. To assess the weighted averages for the Eastern ramp and fresh air raise and to compare underground geotechnical mapping from the as-built decline, the following boreholes were used: a) all of GT11-05 below mineralization, b) GT11-08 from 0-170m, c) GT11-06B (below mineralization to 170m), and, d) GT11-06A (below mineralization to depth). Table 2 summarizes the data. The values in brackets were obtained from the underground mapping exercise.

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Table 2: Summary of Geotechnical Parameters in the Eastern Ramp and Fresh Air Raise
Parameter	Minimum Value	Maximum Value	Average Value
Recovery (%)	60	100	98.54
RQD (%)	10 (10)	100 (90)	69.35 (63)
Hardness	R2 (weak rock 5 to 25 MPa)	R6 (Extr. strong rock >250 MPa)	R4 (Strong rock 50 to 100 MPa)
Joint Number (Jn)	1 (9)	15 (20)	6.65 (9.8)
Joint Condition (Jc)	10 (10)	25 (25)	19.48 (21.5)
Joint Roughness (Jr)	1.5 (1)	3 (1.5)	1.92 (1.48)
Joint Alteration (Ja)	1 (1)	3 (4)	1.45 (1.24)
JRC	1 (3)	11 (13)	5.85 (8)
Representative Q	0.5 (0.1)	142.5 (20)	27.49 (9.6)
Representative RMR	49 (35)	87 (80)	69.01 (69)

The average geotechnical parameters from core logging in the Eastern ramp and fresh air raise show that the rock is good quality based on both the Q and RMR values. The jointing and rock strength will control excavation stability. The fair RQD means that locally the rock is expected to be fractured, but generally the rock mass contains two joint sets (with one random set), slightly altered, rough planar joints.

In July 2012 the first 420 m of the ramp was geotechnically mapped by EBA staff. The ramp was found to be generally dry with minor seepage in the last 70 m of ramp mapped. The figures in brackets in Table 2 are the parameters obtained from the mapping exercise as a comparison to the parameters obtained from core logging. It can be seen that the representative average Q value is lower than that obtained from core logging. This is mainly due to a far lower maximum Q value obtained from the mapping exercise and the fact that generally three joint sets are apparent in the actual section of ramp mapped.

It is imperative that regular geotechnical mapping of the ramp access is carried out to confirm these initial Q values and to correlate Q values obtained from core logging and mapping. Support recommendations remain the same for both the values obtained – see Section 4.4.

A useful tool that is currently used for QA/QC correlation of mapping and borehole log data is the Televiewer. It is a valuable tool in obtaining 360o digital images of the borehole walls and therefore in-situ fractures and other defects along the borehole. As the Televiewer logs the hole in-situ it provides a far more accurate representation of actual rockmass conditions than traditional core evaluation which has been disturbed from its in-situ state. The Televiewer can accurately provide structural orientation data for the borehole and can provide a means of QC/QA for traditional orientated boreholes. It can also provide valuable structural data in jointed zones where core has returned broken or lost.

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The Televiewer can be used in the orientated exploration boreholes for accurate QA/QC correlation and subsequently be used in non-orientated boreholes where gaps in orientation data have been highlighted without the additional expense and time for further orientated drilling.

3.2   Eastern Ramp Between 597 Level and 471 Level, and Planned Main Air Raise

To assess the weighted averages for this section of the ramp and the main air raise, the following boreholes were used: a) all of GT11-06A below mineralization, and b) GT11-06B (below mineralization to 190 m). Table 3 summarizes the data.

Table 3: Summary of Geotechnical Parameters in Eastern Ramp Between 597 Level and 471 Level, and Planned Main Air Raise
Parameter	Minimum Value	Maximum Value	Average Value
Recovery (%)	88	100	99.65
RQD (%)	10	100	71.9
Hardness	R3 (Medium strong rock 25 to 50 MPa)	R6 (Extr. strong rock >250 MPa)	R4 (Strong rock 50 to 100 MPa)
Joint Number (Jn)	1	15	8.4
Joint Condition (Jc)	25	25	20
Joint Roughness (Jr)	0.5	3	2.01
Joint Alteration (Ja)	1	2	1.61
JRC	3	9	6.61
Representative Q	0.5	276.2	28.73
Representative RMR	50	87	72.59

The rock in this section of ramp development and main air raise is of good quality in terms of the Q and RMR rock mass classification systems. The rock is well jointed with almost three joint sets on average being present. The joints are generally unaltered and rough planar.

3.3   Eastern Ramp – Lower

To assess the weighted averages for the lower eastern ramp, borehole GT11-07 was used between 275.5 to 449.5 m. Table 4 summarizes the data.

Table 4: Summary of Geotechnical Parameters in the Lower Eastern Ramp
Parameter	Minimum Value	Maximum Value	Average Value
Recovery (%)	94	100	99.51
RQD (%)	22	98	76.22
Hardness	R2 (weak rock 5 to 25 MPa)	R5 (Very strong rock 100 to 250 MPa)	R3.5 (Medium strong to strong rock 25 to 100 MPa)
Joint Number (Jn)	1	15	6.3
Joint Condition (Jc)	20	25	17.49
Joint Roughness (Jr)	1	1.5	1.43
Joint Alteration (Ja)	1	2	1.13
JRC	3	13	6.57
Representative Q	1	146	40.52
Representative RMR	45	87	69.36

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Only one borehole was used to estimate the geotechnical characteristics within the lower eastern ramp.

On average the rock mass quality is very good based on the Q-system and good based on the RMR system. The joints are unaltered, rough planar with two joint sets plus a random being present at any one place. The rock strength on average is 3.5 indicating medium strong to strong rock. This average is slightly lower than elsewhere.

4.0   UNDERGROUND STABILITY ASSESSMENT AND SUPPORT OF PERMANENT ACCESS

4.1   Kinematic Stability

This section of the report assesses the effect of joint orientation on the ramp and raise stability. Joints or other discontinuities can interact to form trapezoid or pyramidal shaped blocks of rock that may become unstable when the ramp or raise is excavated. A stereonet analysis was carried out on the measured joint sets (having adjusted the alpha and beta angles to give true dip and dip direction orientations) to assess if the joint orientations could potentially form blocks of unstable rock. Given that the ramps spiral down the sidewalls have all possible orientations. The vertical raises also have all possible orientations for vertical kinematic stability.

The total number of successful orientation measurements was relatively low at 114 recorded orientations, and all structural data was combined onto a single stereonet. Figure 3 presents a stereonet of the major joint sets from all data; these joint sets are summarized below.

1 = 73 / 101.

2 = 89 / 017.

3 = 68 / 338.

4 = 62 / 034.

4.2   Kinematic Stability of Underground Ramps

The underground stability of the ramps was analyzed using the Rocscience program Unwedge

(Rocscience 2011). Within this program, the dimensions and shape of the excavation were entered as well as the assumed in situ stress field at a nominal depth of 100 m. The shear strength properties of the discontinuities were assigned a friction of 35° with no cohesion. The Unwedge program assigns an orientation of the excavation to assess if kinematically feasible wedges are formed. Given that the ramps vary in orientation as they spiral down, the azimuth of the alignment of the ramps was varied in 15 degree increments from 000 degrees (whole circle bearing) through to 180 degrees (180 to 360 was not assessed as it is a mirror image of the first assessment). The results of this analysis are shown in Table 5. Wedges less than 0.1 m3 in size have been ignored, as have wedges formed in the floor (sill/invert) of the excavation. Typical output plots from the Unwedge program are shown in Figures 5 and 6.

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Table 5: Results of Underground Kinematic Stability Analysis
Ramp Orientation (whole circle bearing)	Number of wedges reported	No. of wedges where factor of safety is less than 1.0	Mechanism of failure and volume (m3)
000 & 180	4	1	Falling wedge; volume = 0.77 m3
015 & 195	4	1	Sliding on joint 2; volume = 0.66 m3
030 & 210	4	1	Falling wedge; volume = 0.66 m3
045 & 225	4	1	Falling wedge; volume= 1.125 m3
060 & 240	4	1	Falling wedge; volume = 2.3 m3
075 & 255	4	0	(One falling wedge identified where factor of safety was 1.18, but wedge volume was 5.29 m3).
090 & 270	4	1	Sliding on joint 4; volume = 0.5 m3
105 & 285	4	0	NA
120 & 300	4	1	Sliding on joint 3; volume = 0.278 m3
135 & 315	4	1	Falling wedge volume = 0.435 m3
150 & 330	4	1	Falling wedge volume = 0.38 m3
165 & 345	4	1	Falling wedge volume = 1.2 m3
180 & 000	4	1	Falling wedge volume = 0.77 m3

Generally, the wedge size is relatively small, although wedges falling from the crown or back of the tunnel are a potential mode of failure. Only one sizable wedge was found at an orientation of 075 and 255 degrees, where a low factor of safety for a wedge formed between joint sets 2, 3 and 4 may give rise to a wedge 5.29 m3 in size.

The analysis summarized in Table 5 is sensitive to the horizontal stress component. As horizontal stress increases, it tends to ‘lock in’ potential kinematically feasible blocks of rock. Therefore, wedge failure is considered to be significantly less for the mid and lower levels within the eastern ramp than for the near surface in the eastern ramp.  In theory, locating the excavations within the upper eastern ramp on top of one another, channels any horizontal stress present into the rock between the excavations, potentially assisting to ‘lock in’ kinematic wedges.

The sizes of the wedges formed within the ramp were used to check the sufficiency of the standard rock bolt length to support the rock (detailed in Section 4.6).

4.3   Kinematic Stability Underground Raises

The program Unwedge was also used to assess the kinematic stability of the raises. In this case, the raises were modelled as semi-vertical, circular excavations with a 3.5 m diameter. The results of this analysis showed that a combination of joint sets 1, 2, and 3 form potential wedges with a factor of safety lower than 1.0. However, the volume of the wedges formed from joint sets 1, 2, and 3 was relatively small (0.37 m3). Figure 7 shows the results of this analysis. Joint set combination 2, 3, and 4 could also form potential wedge failures, but the factor of safety of this combination is above 1.4, and the wedge size is again relatively small and less than 0.75 m3. Figure 8 shows the results of this second analysis.

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Based on these initial results, it does not appear that sizeable wedges will form within the raises.  Critically, in the case of the raise bore, the wedge size is small enough to fall down the raise without causing a potential blockage. The size of the wedges has in part been used to assess the supporting requirements within the raises.

4.4   Support Design

Based on the equivalent dimension of 2.8 calculated in Section 2.5.1, the following support is envisaged for differing ranges of Q-value. Figure 4 presents the Q-system stability graph.

Q-Value> 1.75 = Classes 1 & 2: No support, spot bolting only.

Q-Value range 0.6 to 1.75 = Class 3: Systematic bolting at a spacing of 1.7 m and 50 mm steel fibre reinforced shotcrete.

Q-Value range 0.15 to 0.6 = Class 4: Systematic bolting at a spacing of 1.4 m and 60 to 90 mm steel fibre reinforced shotcrete.

Q-Value range 0.04 to 0.15 = Class 5: Systematic bolting at a spacing of 1.3 m and 90 to 120 mm steel fibre reinforced shotcrete.

Q-Value range 0.01 to 0.04 = Class 6: Systematic bolting at a spacing of 1.1 m and 120 to 150 mm steel fibre reinforced shotcrete.

Q-Value range < 0.01 = Class 7: Systematic bolting at a spacing of 1.0 m, installation of reinforced ribs of shotcrete (or steel sets with invert closure if preferred) and more than 150 mm of steel fibre reinforced shotcrete.

4.5   Scaling and Ground Support for Personnel Safety

Although the NGI Q-system indicates that for Q-value greater than 1.75 (i.e., Classes 1 and 2) either no support or only spot bolting is required, such a recommendation may not conform to local practice. At some mine sites where a scheme of thorough scaling and routine detailed inspection occurs, having the back or crown of the excavation left unsupported is feasible in the context of an overall rigorous quality control and assurance ground support program. It is also appreciated that some mine sites prohibit workers from entering unsupported ground, and the back or crown of the excavation is always supported.

At Santa Elena, it may be preferable to support the back with wire mesh fixed in place with split sets. If this is the case, the welded wire mesh should have an aperture of 100 mm by 100mm and a section gauge of 5 mm. Panels are generally sized by imperial units (8 by 4 feet) and should have at least six split sets per panel (including overlap of 300 mm/ 1 foot, such that the adjacent panel has a minimum of three split sets installed). The split sets should be a minimum of 1.2 m long.

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The choice of a rigorous scaling program or the use of mesh should be left to the owner’s discretion. Based on the Q-values, we do not foresee the need to spray shotcrete down the entire ramp.

4.6   Rockbolt Details and Design

The length of rock bolts can be estimated from the excavation width and the excavation support ratio. Based on this assessment and the potential size of wedge failure (detailed in Section 4.2 of this report), two alternative types of rock bolt are envisaged at this feasibility stage.

§   Alternative A = 2 m long super-swellex bolts installed and then inflated as per Atlas Copco manufacturers recommendations.

§   Alternative B = 2 m long, 25 mm diameter fully threaded bar (steel grade 517 / 690 MPa), fully resin grouted in place.

A stock of 3 m long rock bolts should also be on site and available for either rock bolting or fore-poling difficult and poor ground. At intersections, using longer rock bolts up to 3 m should also be considered, due to the increased span of the excavation

4.7   Shotcrete Design

The shotcrete design will be developed by the contractor undertaking the mining work to suit the equipment and logistics of the site. Although specifications for the shotcrete will be developed during detailed design, the following preliminary specification has been prepared for this report.

§   Minimum 28 day strength from sprayed panels on site = 35 MPa

§   Minimum 28 day flexural strength from sprayed panels on site = 5 MPa

§   Minimum toughness indices index I20 from sprayed panels on site = 12.0

§   Minimum steel fibre reinforcement content = 50 kg / m3

4.8   Poor Ground and Fault Zone Support Design

The contractor should allow for poor and faulted ground and should have suitable equipment and materials on site to maintain the stability and integrity of the excavation if such ground conditions are found.

Support within these fault zones will be critical as potentially weak and fractured ground could affect the stability of the excavation and the safety of mine workers. The ability to support and reinforce difficult ground should include:

§   Ability to spray shotcrete quickly and efficiently

§   Suitably designed steel sets, on hand at site

§   3 to 4 m long forepoles

§   Ability to grout up to 20 m ahead of the excavation (with down the hole hydraulic packers as well as shorter packers).

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4.9   Summary of Support Requirements for Ramps

To estimate the support requirements within the excavation, the average NGI Q-value (from nearby and relevant boreholes) was assessed. Table 6 summarizes the average Q-values in different parts of the proposed underground workings.

Table 6: Summary of Anticipated Q-Values for Excavations
Location	Elevation		Rock Type	Average Q - value	Rock Support Class	Comments
	From	To
Eastern ramp	785	750	Rhyolite	7.55	Class 1 or 2 (unsupported/spot bolting)
Eastern ramp	750	700	Rhyolite/Andesite	16.72	Class 1 or 2 (unsupported/spot bolting)
Eastern ramp	700	650	Rhyolite/Andesite	18.85	Class 1 or 2 (unsupported/spot bolting)
Eastern ramp	650	625	Rhyolite	20.65	Class 1 or 2 (unsupported/spot bolting)
Eastern ramp	625	550	Rhyolite/Andesite	19.74	Class 1 or 2 (unsupported/spot bolting)	Based on GT-11-08 average Q is 44.3, whereas, GT-11-06B average Q is 10.3.
Eastern ramp	550	500	Andesite/Rhyolite	10.71	Class 1 or 2 (unsupported/spot bolting)
Eastern ramp	500	450	Andesite	16.78	Class 1 or 2 (unsupported/spot bolting)
Eastern ramp	450	328	Andesite	41.74	Class 1 or 2 (unsupported/spot bolting)
Main Eastern Air Raise	750	650	Andesite/Rhyolite	11.17	Class 1 or 2 (unsupported/spot bolting)	Lower raise section has an anticipated higher Q value similar to the bottom section of the ramp of 41.74
Eastern Fresh Air Raise as-built	800	625	Andesite	17.47	Class 1 or 2 (unsupported/spot bolting)

Based on these average values, in general, no support or only spot bolting support will be required throughout the mine. However, the distribution of all relevant Q-values calculated shows that poorer rock than indicated by the average can be expected. Figure 9 shows the distribution of all Q-values in both the Andesite and Rhyolite within the footwall of the deposit. Based on Figure 9 and the distribution of Q-values as shown above, Table 7 presents the lengths of excavation that should be assumed to occur for each support class. This table should be compared to the actual support installed to the 625 Level as a comparison.

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Table 7: Expected Lengths of Support Classes Within the Ramp to 625 Level
Q-Value Range	Support Class	Percentage %	Length of Excavation1 (m)	Recommended Support
> 1.75	1 & 2	87.2	1177	No support, spot bolting only2
0.6 to 1.75	3	9.4	127	Systematic bolting at a spacing of 1.7 m and 50 mm steel fibre reinforced shotcrete.
0.15 to 0.6	4	3.4	46	Systematic bolting at a spacing of 1.4 m and 60 to 90 mm steel fibre reinforced shotcrete.
0.04 to 0.15	5	0	0	Systematic bolting at a spacing of 1.3 m and 90 to 120 mm steel fibre reinforced shotcrete.
0.01 to 0.04	6	0	0	Systematic bolting at a spacing of 1.1 m and 120 to 150 mm steel fibre reinforced shotcrete.
< 0.01	7	0	0	Systematic bolting at a spacing of 1.0 m, installation of reinforced ribs of shotcrete (or steel sets with invert closure if preferred) and more than 150 mm of steel fibre reinforced shotcrete.
Notes: 1      Length of excavation is within standard ramp having a width of 4.5 m and height of 4 m. Based on a 12% gradient the length of the Eastern ramp is 1350 m. This has been used as the baseline for the excavation length. This does not include the air raises or the proposed development at the base of the open pit. At intersections additional support may be required both in terms of shotcrete and longer rock bolts. 2      As discussed in Section 4.9, although no support might be recommended by support classes 1 and 2 of the NGI Q-system, some optional support for worker safety may be desirable such as mesh fixed with split sets.

4.10   Fresh Air Raise Excavation and Support

This air raise has been built to 625 Level and the following support recommendations should be compared to actual as-built design. The fresh air raise is expected to be excavated using a short 10 to 15 m long horizontal access excavated from each level of the ramp. The 2.1 x 3 m diameter raise will essentially connect these access points together by excavating upwards in 20 m vertical, sequential stages as the ramp is progressed. Within the raise, initial support will consist of spot rock bolting only, until the 20 m section has holed through with the access above. Good quality scaling is critical to the safety of this part of the works following each blast. Support for the temporary platform constructed at the start of each shift, as the raise progresses, will also be needed using resin-grouted rebar shaped to suit the heavy timber used for the platform (this platform is dismantled prior to blasting).

Once a section of the raise has holed through, then the sidewalls of the raise should be supported with welded wire mesh panels and split sets. This will help ensure that rock does not fall down the entire raise if instability occurs and creates a rock fall. Having supported a particular section of the raise, a concrete bulkhead wall should be constructed within the access way leading off the ramp. The purpose of this concrete bulkhead wall is to maintain the correct air flow throughout the ramp excavations (with a door for emergency access). The intention is to mount one or two ventilation fans into the concrete bulkheads.

The access points to the fresh air raise from each level of the ramp will make ideal and safety bays during excavation.

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4.11   Proposed Main Eastern Air Raise Excavation and Support

It is assumed that the 3.4 m diameter eastern fresh air raise will be excavated using raise boring techniques. A shaft chamber will need to be excavated at the proposed shaft base off the Eastern ramp. The shaft chamber should at least be 6 m wide and 6 m long to accommodate the raising cutter head and provide an ample target for the probe hole and accompanying drill string.

The stability of the raise bore walls will likely be controlled by joint sets forming wedge shaped instabilities. At this stage, the rock appears to be of sufficient quality that mass or large scale instability from the walls of the shaft will not occur. However, this should be verified by drilling a vertical geotechnical hole offset (by 5 to 15 m) from the location of the raise bore.

During raise boring, the type and consistency of the cuttings should be carefully recorded in case poor ground is struck. Support of the raise bore can only occur once the cutter head has reached surface and is removed. Support will be installed from the top down within the shaft, following thorough scaling of a particular section. An allowance should be made to have the entire shaft supported with mesh and split sets so as to prevent rock fall. Only in the event of massive, unfractured rock can the shaft be left unsupported. Spot bolts and potentially pattern bolts (in poor ground) may also be required. An additional support assessment should be undertaken once the vertical investigation hole has been drilled.

5.0   SUMMARY AND CONCLUSIONS OF GEOTECHNICAL AND STABILITY CONDITIONS FOR RAMPS AND RAISES

5.1   Eastern Ramp As-Built and Fresh Air Raise

The following conclusions in this section should be compared to the as-built ramp and fresh air raise:

§   Both Rhyolite and Andesite rock types can be expected.

§   The average geotechnical parameters in the eastern ramp and fresh air raise show that the rock is good quality based on both the Q and RMR values. The jointing and rock strength will control excavation stability. The fair RQD means that locally the rock is expected to be fractured, but generally the rock mass contains two joint sets (with one random set), slightly altered, rough planar joints.

§   The groundwater conditions are considered to be “damp” in that less than 10 litres of inflow per 10 m of excavation can be expected.

§   The stress conditions within the eastern ramp are considered to be favorable for excavation, with stress related to overburden being the principal component. Stress is considered to be less than 5 MPa.

§   There is the potential for kinematic wedge instability, although in general the size of the wedges formed is relatively small (< 1.2 m3). Wedge-related instability is likely to be most prolific within the western ramp, where low horizontal stress does not ‘lock in’ potential kinematic wedges.

§   Based on the average overall Q-value, the support class will be either 1 or 2, indicating no support or only localized spot bolts. However, poorer rock is expected to occur, and shotcrete will be required in just over 10% of the excavation.

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§   Additional support may be required for worker safety, at the discretion of the owner in coordination with the mining contractor. Such support may consist of installing mesh fixed with split sets on the back (crown) of the excavation throughout the mine.

5.2   Eastern Ramp Between 597 Level and 471 Level, and Planned Main Air Raise

The following summary conclusions can be made about the geotechnical conditions within this ramp section and planned main air raise:

§   Both Rhyolite and Andesite rock types can be expected within the raise bore. Rhyolite can be expected within the ramp.

§   The rock in the ramp and air raise is considered to be good quality in terms of the Q and RMR rock mass classification systems. The rock is well jointed with almost three joint sets on average being present. The joints are generally unaltered and rough planar.

§   The groundwater conditions are considered to be “damp” in that less than 10 litres of inflow per 10 m of excavation can be expected.

§   There is the potential for kinematic wedge instability, although in general the size of the wedges formed is relatively small (< 1.2 m3). Wedge instability may also occur within the raise bore, although the size of these wedges is considered to be less than 0.75 m3.

§   Based on the average overall Q-value, the support class will be either 1 or 2, indicating no support or only localized spot bolts. However, poorer rock is expected to occur, and shotcrete will be required in just over 10% of the excavation.

§   Additional support may be required for worker safety, at the discretion of the owner in coordination with the mining contractor. Such support may consist of installing mesh fixed with split sets on the back (crown) of the excavation throughout the mine.

5.3   Lower Eastern Ramp

The following summary conclusions can be made about the geotechnical conditions within the eastern ramp:

§   Both Rhyolite and Andesite rock types can be expected.

§   Based on only one borehole, within the lower eastern ramp the rock mass quality is very good based on the Q-system and good based on the RMR system. The joints are unaltered, rough planar with two joint sets plus a random being present at any one place. The rock strength on average is 3.5 indicating medium strong to strong rock. This average is slightly lower than elsewhere within the mine.

§   The groundwater conditions at this stage are considered to be “damp” in that less than 10 litres of inflow per 10 m of excavation can be expected. However, since the eastern ramp will develop the highest hydraulic gradient within the mine, this assessment will need to be updated based on additional hydro-geological modelling and actual mine pumping data. Section 9.3 of this report gives specific recommendations to assess groundwater.

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§   The stress conditions within the eastern ramp are considered to be favorable for excavation. Some stress related problems may occur where the rock is locally weaker (UCS less than 50 MPa).

§   There is the potential for kinematic wedge instability, although in general the size of the wedges formed is relatively small. Wedge-related instability is likely to be least prolific within the lower eastern ramp, due to the horizontal stress component being relatively high and tending to ‘lock in’ potential kinematic wedges.

§   Based on the average overall Q-value, the support class will be either 1 or 2, indicating no support or only localized spot bolts. However, poorer rock is expected to occur, and shotcrete will be required in just over 10% of the excavation.

§   Additional support may be required for worker safety, at the discretion of the owner in coordination with the mining contractor. Such support may consist of installing mesh fixed with split sets on the back (crown) of the excavation throughout the mine.

6.0   PRODUCTION STOPE DESIGN

6.1   General

The geotechnical analysis of production stoping at Santa Elena is based on geotechnical data collected from exploration drill holes advanced as part of the 2012 subsurface exploration program and geotechnical test work conducted on core samples at Sonora University.

This section provides the following information:

§   Summary of the range of variance of the rock mass conditions (by means of RMR76 and NGI-Q classification systems) at the following five zones as represented in Figure 10:

1.   At the ore zone (referred herein as Zone D1).

2.   Country rock above the hanging wall, from 5-6 m to 50-60 m above the upper ore zone/country rock interface (referred here in as Zone D2).

3.   Country rock above the hanging wall, between the upper ore zone/country rock interface and about 5-6 m above it (referred here in as Zone D3).

4.   Country rock below the foot wall, from 5-6 m to 50-60 m below the lower ore zone/country rock interface (referred here in as Zone D4).

5.   Country rock below the foot wall, between the lower ore zone/country rock interface and about 5 6 m below it (referred here in as Zone D5).

§   The results of the assessment for three underground methods:

-   Non-man entry longitudinal sublevel open stoping method with pillars (no fill) for ore body dipping greater than 45-50°. This approach includes:

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-   Point line anchor hanging wall cable support installed in fan rings drilled into the hanging wall of a stope from sublevel drill drifts.

-   External (unplanned) dilution estimations.

-   Stope pillars - vertical (rib) and horizontal (sill) pillars associated with the sublevel stoping mining method.

§   Mechanized cut and fill mining method (man entry) for ore body dipping less than 45-50° or at steeper angles when the hanging wall of the stope is weak.

§   Transverse stoping for wide areas of the orebody.

§   EBA QA/QC all the core logs provided by Silver Crest. Three of the core logs were already retrieved on site on September 5th and 6th, 2012. It was found that the RQD values were lower than the actual values. This can be attributed to the fact that the core boxes were only 60 cm in length. The driller’s off-sider had to break the core in order to enable the core to fit in the boxes. EBA corrected the RQD values before processing the data for the purpose of the pre-feasibilty report.

§   Crown Pillars.

6.2   Geotechnical Data Collection

6.2.1   Drill Program

The geotechnical investigation consisted of:

§   Geotechnical data collection on 16 exploration drill holes advanced as part of the 2012 subsurface exploration program. The coordinate and elevation of the 16 drill holes are presented in the Table 8.

Geological and geotechnical logging of the drill holes listed in the Table 8.  This work was conducted by Silver Crest’s mine geologist prior to splitting the core. The geotechnical logging includes collection of geotechnical parameters to estimate the Rock Mass Rating (RMR) system (Bieniawski 1976) and the Norwegian Geotechnical Institute’s NGI Q-System (after Barton et al. 1974). EBA undertook QA/QC of the rock core during site visits at the Santa Elena project. During this process, EBA reviewed (QA/QC) the geotechnical logs against the core photos and made corrections where appropriate.

§   Six (6) of the 16 drill holes with core orientation (refer to Table 8).

§   Collection of rock samples for geotechnical laboratory purposes.

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Table 8: 2012 Borehole Summary
Hole ID	Northing (m)	Easting (m)	Surface Elevation (m)	Azimuth (°)	Plunge (°)	Total Length (m)	Oriented Core Data
GT-12-09	3321358.7	81322.9	784.8	360	60	280.35	Y
GT-12-10	3321349.6	581448.9	776.8	360	70	310.50	Y
GT-12-11	3321348.7	581449.5	777.0	315	70	344.45	Y
GT-12-12	3321449.8	581599.1	808.1	315	45	240.35	Y
GT-12-13	3321383.0	581450.1	784	360	70	280.70	Y
SERC-12-32	3321351.7	581551.5	790.3	045	70	380.00	Y
SE-12-41	3321385	581450	784	360	45	234	N
SE-12-47	3321404	581525	793	015	60	228.45	N
SE-12-48	3321403.4	581525.8	788.9	010	50	215.15	N
SE-12-50	3321466.4	581607.9	808.6	345	78	278	N
SE-12-51	3321467.0	581608.0	808.7	015	70	275	N
SE-12-66	3321391.7	581579.7	802.0	358	86	359	N
SE-12-71	3321391.7	581579.7	802.0	320	82	332	N
SE-12-72	3321182.3	581734.3	853	030	65	600	N
SE-12-90	3321358.7	581323.0	784	020	55	218.75	N
SE-12-91	3321200.5	581640.6	853	013	59	455	N

6.2.2   Laboratory Geotechnical Testing

During the core logging in 2012, samples of the core were collected to provide specimens for geotechnical laboratory testing for Uniaxial Compressive Strength, Direct Shear Testing, Brazilian tensile strength testing and point load testing (PLT). The results of these tests (except for direct shear tests) have been made available to EBA. The UCS tests did not include strain measurement data.

6.3   Rock Mass Characterization

EBA evaluated the rock mass quality using the NGI-Q and RMR systems for the 16 drill holes listed in Table 8. For this purpose, EBA processed the geotechnical logs in raw spread sheets.

The rock mass was subdivided into five geotechnically distinct zones as described in Section 6.1. Summaries of the rock mass properties are presented in the back of this memo in the following tables:

§   Table 9 – Summary of the average RMR76, NGI-Q and RQD values for each of the 16 drill holes listed in Section 6.2.1.

§   Table 10 – Summary of the average geotechnical rock mass properties for each of the five subject zones.

§   Table 11 – Summary of the NGI-Q distributions for each of the five subject zones.

§   Table 12 – Summary of the RMR76 distributions for each of the five subject zones.

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In addition to the tables listed above, the following figures serve to summarize the rock mass properties at the proposed underground location:

§   Figure 11 – Summary of the average NGI-Q values for each of the five zones listed in Section 6.1.

§   Figures 12 through 16 – Plots showing the cumulative percentage of the RMR76 and NGI-Q for each of the five zones described in Section 6.1.

In summary based on the tables and figures above, the following is interpreted for each of the five zones:

§   At the ore zone (referred herein as Zone D1) – The rock quality of this zone is mainly rated as (i) fair to good rock based on the RMR rating classification system and (ii) fair to extremely good rock above El. 580 m and poor to very good rock below El. 580 according to the Q rating classification system.

§   Country rock above the hanging wall, from 5-6 m to 50-60 m above the upper ore zone/country rock interface (referred here in as Zone D2) – The rock quality of this zone is mainly rated as (i) fair to good rock based on the RMR rating classification system and (ii) poor to good rock according to the Q rating classification system.

§   Country rock above the hanging wall, between the upper ore zone/country rock interface and about 5 6 m above it (referred here in as Zone D3) – The rock quality of this zone is rated as (i) poor to good rock above El. 580 and fair to good rock below El. 580 m according to the RMR rating classification system and (ii) poor to very good rock according to the Q rating classification system.

§   Country rock below the foot wall, from 5-6 m to 50-60 m below the lower ore zone/country rock interface (referred here in as Zone D4) – The rock quality of this zone is rated as (i) fair to good rock according to the RMR rating classification system and (ii) poor to very good rock according to the Q rating classification system.

§   Country rock below the foot wall, between the lower ore zone/country rock interface and about 5-6 m below it (referred here in as Zone D5) – The rock quality of this zone is rated as (i) fair to good rock according to the RMR rating classification system and (ii) poor to very good rock according to the Q rating classification system.

6.4   Design Parameters

6.4.1   Rock Mass Parameters

Figures 12 through 16 present plots of the cumulative percentage distributions of the NGI-Q and RMR values. The cumulative percentage distribution plots provide the means of progressively estimating the likelihood that an RMR or NGI-Q value will be equaled or exceeded. For this preliminary prefeasibility design the NGI-Q and RMR values at 50% were chosen. It should be understood that the design values at 50% could be over conservative for some areas of the mine where the actual NGI-Q and RMR values are considerably higher, and therefore would not likely be the most economical option, or may not work

(i.e. unstable) for areas of the mine where the NGI-Q and RMR values are considerably lower. The purpose of adopting the 50% design values for prefeasibility design is to provide stope dimension and reinforcement that are not too conservative or too optimistic, and therefore serve for economic estimation purposes.

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Table 13 summarizes the NGI-Q and RMR values used for the preliminary design.

Table 13: RMR’ and NGI-Q’ Values for Design – at 50%
Zone	Elevation	RMR’76	NGI-Q’
D1	Above El. 580 m	75	29
	Below El. 580 m	76	11
D2	From surface to bottom of planned workings	52	4
D3	Above El. 580 m	67 (711)	12 (211)
	Below El. 580 m	52	3
D4	Above El. 580 m	70	13
	Below El. 580 m	67	13
D5	Above El. 580 m	67	7
	Below El. 580 m	72	14
1      Two values are presented; (i) the first value considers the data from all the drill holes, and (ii) the second (in brackets) does not consider three of the drill holes where potential faults were encountered.

In addition to this assessment set out in Table 13, NGI-Q values were assessed specifically for the 655 Level. Core runs were isolated over a 45m run length at this level to give an individual rock mass characterization for stoping at this level. NGI-Q values of 15 and 6 were calculated for zones D1 and D3 for stopes at this level.

From core logging during the 2012 drilling program it would appear that the rock mass quality is degrading with depth. However, from footwall cores summarized in Table 6 (2011 drill holes), the rock mass quality is good and relatively consistent with depth. This degradation in rock mass quality needs investigating further as it impacts on the recommended stope dimensions below the 580 level. There is concern that due to limited drilling, rock mass quality and hence stope lengths are being underestimated here.

It is recommended that further drilling be conducted underground during development of the current 655 Level in order to build on drilling data and resolve this issue in a timely manner. Drilling could be conducted in remuck bays driven into the hangingwall side of the stope access to avoid extra exploration development and drilled to depth below the 580 Level to obtain NGI-Q values for the hangingwall and orebody. This program could be combined with exploration drilling to further define the orebody at depth. Geotechnical mapping of the ramps and current orebody and hangingwall exposure at the 655 Level is also recommended in order to correlate mapped NGI-Q values to core logging data.

6.4.2   Rock Strength

The uniaxial laboratory compressive results submitted an average value of 71 MPa which appears to be low for the intact material of the site – refer to Appendix A.

The average orebody strength from UCS testing at Sonora University was 72MPa. Point load testing from Sonora indicated an average equivalent UCS of 136MPa. Average Brazilian tests at Sonora indicated a tensile strength of 10 MPa in the orebody. Typical tensile strengths in rock are approximately 10% of the UCS value. An initial estimate of orebody UCS strength was taken as 100MPa.

From core sent to EBA from Santa Elena four samples were deemed suitable for UCS testing and three samples for Brazilian testing. These were sent to UBC in Vancouver for appropriate testing. See Appendix A for UCS and Brazilian test results from Sonora University, Mexico and UBC, Vancouver.

Two UCS samples from the orebody were tested at UBC, Vancouver with an average strength result of 87MPa. These samples are from the same drill hole SE-13-145 and tie in closely with the average UCS result of 81MPa from drill hole SE-13-145 conducted at Sonora University. It was decided to weight the orebody UCS in favour of Sonora UCS values. The UCS orebody strength was subsequently revised to a value of 87MPa.

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The strength of the hangingwall for the prefeasibility analysis is taken as 92MPa. This is based on the average point load testing data from Sonora University. Three UCS tests were carried out in hangingwall material at Sonora giving an average of 95MPa which ties in closely with the PLT results. UCS testing of Andesite at UBC indicated a strength of 199 MPa however, it should be noted that this is only based on one result.

The strength results obtained so far for the orebody appear to be low for a quartzite rock type. This could be due to local rock conditions or the presence of micro fracturing in the samples taken due to extraction from the in-situ stress regime.

It is strongly recommended that tri-axial testing be carried out on hangingwall, footwall and ore samples together with further UCS testing with strain measurements in order to determine site specific strength and deformation characteristics of the rock types at Santa Elena for inclusion into future geotechnical modelling of the mine for stope and pillar design refinement. See Section 9.0-Recommendations.

As UCS results from Sonora University tie in closely with the limited number of test carried out at UBC it is recommended that 80% of future testing be carried out at Sonora University, Mexico and 20% at UBC, Vancouver as a reference check.

6.5   Design Approach

6.5.1   General

The design approach presented in this report consisted of empirical and analytical methods using the rock mass quality approach. In this respect, it should be noted that as with all the empirical methods, assumptions should be validated once mining starts by comparing actual with anticipated results.  Below is presented the empirical method used for the assessment.

6.5.2   Methodology of Stability Graph for Stope Design

The stability graph method (developed by Mathews et al. (1981) and modified by Potvin (1988) and Nickson (1992) is widely used by designers to qualitatively assess the stability of open stopes. The stability number N is calculated from selected geotechnical factors as described below to give an associated hydraulic radius from the stability graph for two cases of stable with support and stable without support utilizing the following formulae.

Hydraulic Radius =	Cross Section Area	(5)
	Perimeter

The stability number N is defined as:

N = Q' x A x B x C (6)

Where A is the stress factor (ratio of intact rock strength to induced stress), B is rock defect orientation factor, and C is the gravity factor related to orientation of design surface.

The underground design at this stage of the analysis is deemed to be governed by the stability of the hanging wall and stope back surfaces.  The following steps were taken to calculate N values for each level:

§   Q’: From an initial assessment of Q’-Values, two lithology dependent domains “Waste Rock” of the hangingwall immediately above the ore zone (D1) and “Mineralized Zone” (D1) were initially identified for stope design considerations. It was also found from the geotechnical assessment of the borehole data that these domains can be sub-divided into two portions of “above 580 m elevation”, and “below 580 m elevation”. Table 14 shows the Q’ associated with these four domains at 50% level of cumulative frequency. In addition to this NGI-Q values were assessed for the 655 Level. Core runs were isolated over a 45m run length at this level to give an individual rock mass characterization for stoping at this level. NGI-Q values of 15 and 6 were calculated for zones D1 and D3 for stopes at this level.

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Table 14: NGI-Q’ Values for Stope Design
Zone	Elevation	NGI-Q’ (50% level of cumulative frequency)
Mineralized Vein (D1)	Above El. 580 m	29
	Below El. 580 m	11
Waste Rock (D3)	Above El. 580 m	12
	Below El. 580 m	3

§   Factor A: Induced stresses were estimated at the back and hanging-wall of the stope using boundary element computer program Examine 2D.

-   Uniaxial compressive strength of the ore and waste material was considered as 87MPa and 92 MPa, respectively.

-   Four cases 7 m, 12 m, 17 m, and 22 m for the span width were analyzed in examine 2D, considering the variation of ore width at each level (with four classes of ore thickness: 0 m to 10 m, 10 m to 15 m, 15 m to 20 m, and greater than 20 m). The average back and hangingwall span stress at each level was taken for calculation of the A factor.

§   Factor B: From review of the GEMS model, the mineralized vein was subdivided into three sub-domains “West”, “Centre” and “East” with dip and strike of 55°/90°, 62°/72°, and 47°/90°, respectively.

In the case of stope length design, the joint set J5 with the orientation of 79°/178° was found to have the greatest impact in the western and central stopes while the joint set J4 with the orientation of 63°/168° is the most influencing discontinuity set when the hanging-wall is entirely located in the waste material. However, when the ore material contacts the hanging-wall (i.e. when the low grade ore is the barrier between the mined-out span and the waste material), the discontinuity set J1 with the orientation of 43°/222° is the most influencing factor on instability of the eastern and western stopes. The joint set J4 with the orientation of 81°/133° is the dominating factor in central stopes within mineralized vein material.

The associated B factors derived from the available empirical chart for the two scenarios of hanging-walls are presented in Table 15.

Table 15: B Factor for Stope Length Design
Hanging Wall Material		Waste Rock	Mineralized Vein
Stope Location	West	0.21	0.45
	Centre	0.24	0.29
	East	0.22	0.43

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§   Factor C: Factor C is calculated using the empirical relationship 8 – 6 × Cos (dip of the wall) at each location.

Table 16 presents this factor for different domains / locations.

Table 16: C Factor for Stope Length Design
Stope Location	Stope Length Design	Stope Back Design
West	4.6	2.0
Centre	5.2
East	3.9

6.6   Span Length Design

The following key mining parameters / assumptions were used as the basis of the initial stope design:

§   Non-man entry to production level.

§   With an initial assessment of the effect of both hanging wall and the stope back it was found that the hanging wall effect will govern the stope size. Therefore, the hanging wall effect was considered throughout the design process.

§   Height of stopes = 45 m; width of stope dependent on ore body; length of stope dependent on acceptable sloughage limits of 15%. The stope length was refined when the sloughage level exceeded 15% to fall within acceptable sloughage limits.

§   Seven levels were considered for stress analysis and underground stope design (including stope length, back length, sill pillar thickness, and rib pillar width as well as crown pillar thickness). These levels were named as the elevation of the bottom line of the stope considering the recommended sill thickness and the stope height of 45 m.

§   Cable bolted stopes and stable stope spans were assessed, where deemed necessary.

§   Two scenarios were considered for stope length design: When the hanging wall is excavated in waste rock, and when mineralized vein is left in place as part of the hanging wall.

§   Stope widths of 7 m, 12 m, 17 m and 22 m were considered.

Sloughage was considered using the ELOS graph developed by Clark and Pakalnis (1997). This seeks to quantify the degree of dilution anticipated for a given span length and corresponding shape factor.

It must be noted that this graph was developed for weak rock masses. There is no measure of time dependent behaviour but nevertheless this approach provides an indication as to predicted dilution at this stage and should be used in conjunction with observational data during mining. Acceptable sloughage levels of 15% or less were also calculated for stope design utilizing the ELOS criteria for the stability graph after Clark and Pakalnis (1997).

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The predicted hydraulic radius with acceptable sloughage can then be utilized to back-calculate the maximum span for the surface being analyzed with a given design stope height.

Tables 17 to 23 show the stope configuration summaries for each level. They indicate a stable span without support, a stable span with support and a measure of anticipated dilution with greater span lengths. The recommended cablebolt pattern is 5 cables per ring every 2.2 m in stope sublevels. Dilution levels for span lengths greater than those recommended in the stability graph method are indicative only and it is strongly recommended that a site based dilution data base be established to monitor sloughage based on stope spans and rockmass quality.

6.7   Stope Back Width Design

The stability graph method was utilized for designing the “stable” back length similar to hangingwall stope length design. It can be seen in Tables 17 - 23 that the hangingwall stability is the overall controlling factor for stope stability.

At this prefeasibility stage it may be advisable to consider supporting stope backs with cablebolts at the 536 Level and below due to stress effects. However, the magnitude and orientation of the in-situ stress regime is recommended beyond prefeasibility level to verify stress effects on the mine openings.

6.8   Cut and Fill

The dimensions for the cut and fill mining method were estimated using the span design curve (Lang, 1994).

A maximum span of 15 m is recommended in current cut and fill mining levels. The spans refer to spans which may not require intensive ground support such as timber sets, cable bolts or post pillars, but still requires local support within the immediate back to confine potential blocks (loose) which may open due to nearby blasting and/or redistribution caused by subsequent mining activity. 2 m x 2 m pattern bolting is recommended in slashed stopes with spans less than these stable back span figures. Due to the nature of the orebody, spans will vary across levels and also in individual stopes. An assessment of the need of support required for individual stopes needs to be undertaken at a stage beyond the scope of this prefeasibility study.

With a 5 m cut and a maximum stope length of 100 m dilution from blast damage only along the hangingwall should be expected.

A value of Jw=1 was assigned in calculating Q as was an SRF of 1 as described in Section 2.6.2.

6.9   Transverse Stoping

With reference to Table 24 assuming a 20 m stope width and a maximum length of 30 m the stope back should remain stable. Loose from the stope back may be created during the last 10 m of stope production at 471 and 536 Level. Timely mucking and backfill placement is recommended. Stope sidewalls should remain stable. Expect a maximum dilution of 7% from hangingwall sloughage at the 471 and 536 Levels. However, as the hangingwall forms the stope end, dilution will be dependent upon the speed at which the stope is blasted and mucked. If low grade mineralized vein forms the hangingwall then expect blast damage only.

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Table 24: Transverse Stoping

7.0   PILLAR DESIGN

7.1   Sill Pillar Design

Horizontal (sill) pillars were assessed by means of Rocscience Examine 2D software program and by an empirical failure criterion (Hoek and Brown, 2002).

The strength of the sill rock mass depends on the compressive strength of the rock mass material as well as the shape factor of the sill, i.e. ore width and sill thickness, while the sill thickness is being designed. Sill stress depends on the depth and the sill thickness and the width of the opening. For pre-feasibilty purposes, sill stress was estimated in Examine 2D in four cases of stope width 7 m, 12 m, 17 m, and 22 m.

As maximum sill pillar stress is acting in the horizontal direction the ‘width’ of the pillar is its thickness and its ‘height’ corresponds to the orebody width. A constant sill pillar thickness was taken across the entire mining level for practical mining purposes. This constant thickness corresponds to the maximum ore width analyzed. Additionally blast damage of 1 m was also assumed on each side of the pillar.

For two dimensional analysis brittle Hoek Brown parameters were utilized for sill pillar design as the GSI values for the ore vary between 65 to 74 based on converted modified Q’ values for the ore body, Maybee (2000).

Figure 17 shows the pillar stability chart used for the analysis. Average pillar stress was obtained using the RocScience Examine 2D software package with sill pillar width reduced by 2 m at each level due to blast damage consideration.

Table 25 shows the corresponding sill thickness stability above each mine level with the greatest ore span considered having the most effect on stability due to the ratio of width to height.

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Table 25: Sill Pillar Stability
Level	Sill thickness (m)	W/H ratio		Av. Pillar stress		Pillar stress/UCS *		Pillar stability graph
		7m ore width	22m ore width	7m ore width	22m ore width	7m ore width	22m ore width	7m ore width	22m ore width
655	10	1.1	0.4	8.5	10.5	0.10	0.12	Stable	Stable
597	13	1.6	0.5	12	13.5	0.14	0.16	Stable	Stable
536	16	2.0	0.6	21.3	15.8	0.24	0.18	Stable	Stable F.o.S 1.4
471	20	2.6	0.8	17	20.4	0.20	0.23	Stable	Stable F.o.S 1.4
401	25	3.3	1.0	20.6	20	0.24	0.23	Stable	Stable F.o.S 1.4
328	28	3.7	1.2	26	22	0.30	0.25	Stable	Stable F.o.S 1.4
* Ore UCS = 87 MPa

This analysis assumes a constant ore width along ore strike and that the rockmass is homogenous and isotropic. In analyzing pillar design a K stress ratio of one was taken at prefeasibility level. The actual in-situ stress ratio, its inclination and its variation with depth needs to be determined beyond prefeasibility level.

As the orebody varies significantly in thickness, strike and dip, both along ore strike and down dip, a three dimensional analysis of stope and pillar design is recommended as are stress measurements at Santa Elena. See Section 9 - Recommendations.

7.2   Crown Pillar Design

Underground mining will result in a temporary surface crown pillar below the open pit, as well as permanent crown pillars in some areas. The minimum thickness for crown pillar design was assessed as 25 m using the empirical crown pillar stability graph (Carter, 1990) taking a typical Q value of 11.6 for the ore at near surface stress conditions. The maximum span of the orebody was taken as 25 m giving a thickness / span ratio of 1.

7.3   Rib Pillar Design

Rib pillar widths have been assessed based on two dimensional analysis utilizing brittle Hoek Brown parameters and the pillar stability chart as used for sill pillar analysis.

For two dimensional modelling a horizontal cross section was taken across the pillar midpoint in order to incorporate the pillar shape factor in the stress analysis. A constant in-situ stress field dependent on depth was imposed onto the cross section.

For efficiency purposes at the stage of prefeasibility level the rib pillar widths were modelled at the mid-point of the mine at 536 level. EBA modeled five “fixed” stope span length classes of 25, 50, 75, 100 and 125 m. An average pillar height of 56m was used during the analysis based on average dip of the orebody across the West, Central and East domains. Pillar results are shown in Table 26. Stress analysis was conducted on a pillar width less 2 m of final pillar width to take into consideration pillar blast damage.

Table 26: Rib Pillar Stability
Level	Span length	Pillar height	Pillar width	Width/Height	Average stress	Pillar stress/UCS*	Pillar stability graph
536	25	56	15	0.27	13.8	0.16	F.o.S 1.6
	50	56	18	0.32	17	0.20	F.o.S 1.4
	75	56	20	0.36	18.5	0.21	F.o.S 1.4
	100	56	25	0.45	20	0.23	F.o.S 1.4
	125	56	30	0.54	20	0.23	F.o.S 1.4
* Ore UCS = 87 MPa

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These numbers are considered as an initial estimation of rib width at PFS level at mine mid-point. Pillar widths are likely to be reduced at shallower levels and increased at lower levels. However, in order to define pillar widths on a mine wide scale, as the orebody varies significantly in thickness, strike and dip, both along strike and down dip, it is advised that the in-situ stress ratio, its inclination and its variation with depth be determined beyond prefeasibility level along with site specific strength, deformation and Hoek-Brown characteristics of the rock mass. Tri-axial and UCS testing with strain are recommended along with shear strength testing in order to determine site specific strength and deformation characteristics.

A three dimensional analysis of stope and pillar design is recommended once site specific in-situ stress and rock mass deformation and Hoek-Brown characteristics have been determined. This will enable rib and sill widths to be refined at each level and stope boundary beyond pre-feasibilty level in order to maximize ore extraction. Extraction of rib pillars for longhole, cut and fill and transverse mining methods have not been considered at this level and require further investigation for mining sequencing and backfill operations. See Section 9.0 - Recommendations.

8.0   STOPE DESIGN SUMMARY TABLES

Tables 17 to 23 present show the stope configuration summaries for each level for longitudinal and cut and fill mining. They indicate a stable span without support, a stable span with support and a measure of anticipated dilution with greater span lengths on stope surfaces in hangingwall “waste” and stope surfaces where the hangingwall is in low grade “mineralized vein”. Effective cablebolt support has been taken into consideration when designing maximum span lengths for stable spans with support. Stope backs are considered in ore. Maximum critical back spans are listed for cut and fill mining. The stope design summary tables 17 to 23 present the results of the analysis with respect to the rectangular stope design. They are split into West, Central and East locations at each level dependent on average strike and dip of the orebody.

These charts are meant for general design purposes only at prefeasibility stage as the stope spans assume continuous average width, strike and dip profiles. As such a maximum span of 120 m has been applied to the tables. As actual stopes will vary in width, strike and dip along their span further geotechnical study of individual stopes is advised beyond prefeasibility level.

Dilution levels for span lengths greater than those recommended from Potvin’s method are indicative only and it is strongly recommended that a site based dilution data base be established to monitor sloughage based on stope spans and rockmass quality. Further rockmass classification through mapping and logging can assist in defining stope dimensions to a higher level of detail.

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Actual stope and pillar performance should be consistently monitored during mining. This will allow for any modifications of initial design to be made based on actual mining of the orebody.

9.0   RECOMMENDATIONS

Recommendations are split into three sections for the pre-feasibilty report:

§   Key recommendations to improve stope size and refine pillar dimensions;

§   Geotechnical considerations in overall mine design; and

§   General recommendations.

9.1   Stope Size and Pillar Dimensioning

The key factors governing stope size at Santa Elena are:

§   Hangingwall and orebody NGI-Q values;

§   Joint set orientation;

§   In-situ stress regime; and

§   Hangingwall and orebody strength and deformation characteristics.

9.1.1   Hangingwall and orebody NGI-Q Values

§   Due to the apparent degradation of rock mass quality with depth in the 2012 drilling program it is recommended that further drilling be conducted underground. This degradation could be due to limited availability of geotechnical data at present as footwall rock quality appears fairly consistent at depth. This apparent degradation could create overly conservative stope design parameters and it is recommended that rockmass quality at depth requires further clarification through a drilling program.

§   This drilling program could be undertaken during development of the current 655 Level in order to build on drilling data and resolve this issue in a timely manner. Drilling could be conducted in remuck bays driven into the hangingwall side of the stope access to avoid extra exploration development and drilled to depth below the 580 Level to obtain NGI-Q values for the hangingwall and orebody. This program could be combined with exploration drilling to further define the orebody at depth. The use of a Televiewer and sonic probing down hole could provide a tool for accurate geotechnical data collection of the in-situ undisturbed rock mass.

§   Geotechnical mapping of the ramps and current orebody and hangingwall exposure at the 655 Level is also recommended in order to correlate mapped NGI-Q values to core logging data.

§   Silver Crest’s mine geologist recorded NGI-Q values during geological logging of exploratory boreholes in 2012. It is recommended to construct a NGI-Q value data base and a 3D-NGI-Q value model using data from all the 2011 and 2012 exploration holes and underground mapping data. This could allow for refinement of stope design by creating NGI-Q zones at a local level through the mine. Incorporation of observational data during actual mining will be highly advantageous in refining initial design constraints.

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§   Initial reports indicate that running water up to 450 gpm has been intersected within the orebody at the 625 Level. During ramp and ore development reports indicate that any water makes intersected were pockets of water and had drained away naturally. The water make within the orebody needs to be monitored in order to establish if a similar situation occurs here. If this water make is persistent this could have a negative impact on rock mass quality and NGI-Q values and impact on stope dimensions and sloughage levels. The source of water needs to be established whether its drainage from the water table or if it is artisanal which may be the case as the water temperature is at 30o Centigrade. Further exploratory work within the orebody would need to be conducted in order to determine the continuation of the water make along strike and down dip and a de-watering program may be recommended in order to improve on stability conditions within the orebody.

9.1.2   Joint Set Orientation

Joint set orientation can play a key role in stope stability. The orientation data at Santa Elena is limited with six orientated holes gleaning a limited number of measurements to gain joint set data for stope layout.

§   Further joint set measurements can be taken through consistent geotechnical mapping of the orezone and hangingwall during mining activities.

§   Further orientated boreholes can be undertaken during subsurface and surface exploration. During the 2012 surface drilling program a limited number of joint orientation data was able to be retrieved. It is strongly recommended that the use of a Televiewer be incorporated into further data collection. This allows for in-situ joint orientation data to be collected that will aid in stope definition.

§   The Televiewer is able to acquire 360o digital images of the borehole walls and therefore in-situ fractures and other defects along the borehole. As the Televiewer logs the hole in-situ it provides a far more accurate representation of actual conditions than traditional core evaluation which has been disturbed from its in-situ state. The Televiewer can accurately provide structural orientation data for the borehole and can provide a means of QC/QA for traditional orientated boreholes. It can also provide valuable structural data in jointed zones where core has returned broken or lost.

§   The Televiewer can be used in the orientated exploration boreholes for accurate QA/QC correlation and subsequently be used in non-orientated boreholes where gaps in orientation data have been highlighted without the additional expense and time for further orientated drilling.

9.1.3   In-situ Stress Regime

The in-situ stress regime at Santa Elena needs clarification as the stress regime has been assumed at pre-feasibility level.

§   It is strongly advised that in-situ stress analysis be conducted at Santa Elena in order to determine principal stress magnitude and trajectory and horizontal to vertical stress ratio. This should be done in order to maximize ore extraction, minimize pillar dimensions, assist in pillar recovery design and minimize support cost to local rather than regional zones based on the in-situ stress regime. EBA can assist in developing a cost effective stress measurement program in conjunction with Silver Crest for this purpose.

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9.1.4   Hangingwall and Orebody Strength and Deformation Characteristics

Clarification of hangingwall and orebody strength characteristics is recommended beyond pre-feasibility level as there is concern that the UCS strength results that have been returned so far are relatively low for the general rock types encountered.

§   In addition to simple UCS strengths for the footwall, hangingwall and orebody the Hoek-Brown parameters and deformation characteristics need to be determined for the rock conditions found at Santa Elena in order to refine stope and pillar design to a higher degree of accuracy beyond pre-feasibilty level.

§   It is strongly recommended that tri-axial testing be carried out on hangingwall, footwall and ore samples together with further UCS testing with strain measurements in order to determine site specific strength and deformation characteristics of the rock types at Santa Elena for inclusion into future geotechnical modelling of the mine and for stope and pillar refinement. It is recommended that 80% of future testing be carried out at Sonora University, Mexico and 20% at UBC, Vancouver as a reference check.

9.2   Geotechnical Considerations in Overall Mine Design

§   As the orebody varies greatly in dip and dip direction and thickness, both along strike and down dip, it is strongly advised that three dimensional geotechnical analysis be conducted on the orebody once the stress regime and rock deformation characteristics have been determined. This can be highly advantageous to mine sequencing, backfill addition and pillar recovery in order to further maximise ore extraction.

§   It is recommended to space rib pillars across levels so as to avoid stacking of pillars. This will minimize the creation and effect of tall slender pillars on the overall mine wide excavation.

§   Since the ground conditions are not consistently uniform throughout the ore body, it is recommended that in further stages (beyond prefeasibility level) the site be subdivided into detailed geotechnical domains, as rock mass characteristics are further defined through drilling, mapping and actual stope dilution data. This will enable in-depth geotechnical analysis to be conducted to maximize ore recovery from stope and pillar dimensioning on a stope by stope basis, where each domain is represented by a Q’ value. Stability within a domain could be selected to ensure that approximately 80 % (or other percentage) of the stopes are stable, and the remaining 20% or so can then be dealt with individually using specific ground support or different extraction strategies.

§   This prefeasibility study does not consider the interaction between the pit and the proposed underground stopes. This should be reviewed once stoping outlines have been defined for feasibility or design stages.

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9.3   General Recommendations

§   Both the Contractor undertaking the work and a representative from SilverCrest should carry out joint geological and geotechnical mapping of the face and sidewalls of each round of excavation. The representatives, in coordination with the miners, should also agree on the type of support class, and the NGI Q-value. The record could also be used to assess whether geological overbreak has occurred. The actual quantity of overbreak should be based on survey profile records.

§   A tape extensometer should be routinely used by the mine geotechnical department to measure deformation near the face, at intersections and especially within poor ground conditions. These measurements can be used to back calculate the stress conditions and will prove very useful for detailed mine design.

§   The ability to spray shotcrete quickly and efficiently from the top of a muck pile (at short notice) is considered a necessity. If poor ground is struck, stabilizing the face and sidewalls is vital to control the ground and maintain the safety and integrity of the works.

§   Careful records should be made of groundwater inflow being pumped out of the ramp during excavation. Accurate flow meters should be installed at the ramp portal to measure both water inflow and outflow; the difference shows the quantity of groundwater inflow. Also, during daily geological and geotechnical mapping, records should be kept of any localized water makes and the volume of ingress. The records should also indicate if the inflow is structurally controlled and the structure orientation and dip direction and whether the water flow declines over time. These records will be invaluable when assessing the macro hydro-geological conditions within the mine.

§   The performance of the support recommended in this report should be evaluated. The evaluation should be based on instrumentation as well as observations made within the ramp. Refinement of the recommended support should occur based on these evaluations and on site assessments.

10.0   LIMITATIONS OF REPORT

This report and its contents are intended for the sole use of Silver Crest and their agents.  EBA Engineering Consultants Ltd. does not accept any responsibility for the accuracy of any of the data, the analysis, or the recommendations contained or referenced in the report when the report is used or relied upon by any Party other than Silver Crest, or for any Project other than the proposed development at the subject site.  Any such unauthorized use of this report is at the sole risk of the user.  Use of this report is subject to the terms and conditions stated in EBA’s Services Agreement.  EBA’s General Conditions are attached to this memo.

11.0   CLOSURE

We trust this report meets your present requirements. Should you have any questions or comments, please contact the undersigned at your convenience.

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Yours Sincerely,

EBA Engineering Consultants Ltd.

Prepared by:

Ben Howden, M.Sc., M.C.S.M.

Senior Geotechnical Engineer

Direct Line: 604.685.0017 x325

bhowden@eba.ca

Reviewed by:

                                                                                                         

Carlos Chaparro, P.Eng. (BC)	James Barr, P.Geo. (BC)
Senior Geotechnical Engineer	Senior Mining Geologist
Direct Line: 604.685.0017 x242	Direct Line: 604.685.0017 x360
cchaparro@eba.ca	jbarr@eba.ca

36

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

REFERENCES

Barton, N., Lien, R., and Lunde, J., (1974). Engineering Classification of Rock Masses for the Design of Tunnel Support. Published in the Journal of Rock Mechanics.

Bieniawski, Z.T., (1973). Engineering Classification of Jointed Rock Masses. Published in Transaction of the South African Institution of Civil Engineers.

Carter T. G., (1989). Design lessons from evaluation of old crown pillar failures, Procd. Surface Crown Pillar Evaluation of Active and Abandoned Metal Mines, Timmins, pp 177 - 187

Clark, L.M. and Pakalnis, R.C. (1997) An Empirical Design Approach for Estimating Unplanned Dilution from Open Stope Hangingwalls and Footwalls. Proceedings of the 99th AGM - CIM. Vancouver, pp. 25.

Global Resource Engineering Ltd., (November 2011). Santa Elena Underground Mine Dewatering Estimates.

Hoek-Brown Failure Criterion (2002).

Hoek, E., and Bray J. (1977). Rock Slope Engineering. Revised second edition, published by the Institute of Mining and Metallurgy, London.

Lang B., (1994) Span Design for Entry Type Excavations. M.A.Sc Thesis, University of British Columbia, Vancouver, B.C.

Mathews, K., E Hoek, DC Wyllie, DC Stewart, SBV (1980) Prediction of stable excavation spans for mining at depths below 1000 metres in hard rock - Report to Canada Centre for Mining and Energy Technology (CANMET), Department of Energy and Resources; DSS File No 17SQ23440-0-90210 Ottawa, 1980

Maybee W.G., (2000) Pillar Design in Hard Brittle Rocks, M.A.Sc Thesis, Laurentian University

Potvin, Y., (1988) Empirical Stope Design in Canada Ph.D. Thesis Department of Mining and Minerals Processing, University of British Columbia

Rocscience (2011). Dips. Version 5.108.

Rocscience (2011). Unwedge. Version 3.020.

Vector Engineering Inc. (2009). Open Pit Geotechnical Study for the Proposed Santa Elena Mine.

37

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

TABLES

Table 9 Average Geotechnical Properties per Boreholes

Table 10 Zones Average Characteristic properties

Table 11 Zones Representative Q’ Statistics

Table 12 Zones Representative RMR’ Statistics

Table 17 710 West/Central and East Stope Configuration Summary

Table 18 655 West/Central and East Stope Configuration Summary

Table 19 597 West/Central and East Stope Configuration Summary

Table 20 536 West/Central and East Stope Configuration Summary

Table 21 471 West/Central and East Stope Configuration Summary

Table 22 401 West/Central and East Stope Configuration Summary

Table 23 328 West/Central and East Stope Configuration Summary

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

Table 9: Domains Average Geotechnical Properties per Boreholes
Borehole ID	Borehole Surface Elevation (m)	Borehole Inclination (°)	Deposit Ore Interval (m)		Deposit Ore Centre Point Elevation (m)	Geotechnical Properties Per Domains
						Ore			HW' 6-60 m			HW 0-6 m			FW' 6-60 m			FW 0-6 m
						RMR76	Q'	RQD	RMR76	Q'	RQD	RMR76	Q'	RQD	RMR76	Q'	RQD	RMR76	Q'	RQD
GT-12-12	808	45	160	174	690	50	12	47	48	2	45	42	2	19	62	8	59	57	2	57
SE-12-41	784	45	151	163	673	73	69	72	58	22	52	53	33	48	69	18	68	70	12	50
SE-12-48	789	50	175	190	649	87	98	98	43	2	42	77	36	87	86	47	95	---	---	93
SE-12-90	784	55	164	180	643	74	18	62	57	9	55	69	32	85	60	10	67	45	4	27
GT-12-09	785	60	158	180	638	57	7	69	54	8	60	68	10	86	74	89	84	71	7	88
SE-12-47	793	60	191	210	619	88	121	89	57	15	56	67	32	84	89	67	91	79	64	92
SE-12-51	809	70	203	209	615	55	9	36	46	7	48	32	1	25	67	29	62	67	12	48
GT-12-13	784	70	197	218	589	75	50	81	42	3	35	60	26	43	48	7	38	70	34	69
SE-12-50	809	78	221	233	587	75	90	85	50	15	46	54	39	68	77	32	81	72	46	79
GT-12-10	777	70	222	243	558	75	9	79	58	6	57	53	8	72	62	8	54	48	5	50
GT-12-11	777	70	237	255	546	79	12	69	62	3	54	55	3	38	62	4	56	57	3	44
SE-12-71	802	82	287	299	512	64	19	70	38	5	44	50	4	47	70	21	73	74	12	85
SERC-12-32	790	70	305	317	498	55	17	60	48	3	33	49	1.0	21	50	2	30	69	4	56
SE-12-91	853	59	407	428	495	76	48	97	61	16	76	49	23	63	81	29	97	83	44	100
SE-12-66	802	86	317	332	478	81	29	78	42	5	40	50	3	31	43	7	50	74	25	76
SE-12-72	853	65	468	480	423	83	20	70	42	4	37	43	17	53	80	25	81	79	29	86

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

Table 10: Domains Average Characteristic Properties
Characteristic (Average of All Intervals)	D1: Ore		D2: HW' 6-60 m	D3: HW 0-6 m		D4: FW' 6-60 m		D5: FW 0-6 m
	Above 580 m Elevation	Below 580 m Elevation		Above 580 m Elevation	Below 580 m Elevation	Above 580 m Elevation	Below 580 m Elevation	Above 580 m Elevation	Below 580 m Elevation
RQD (%)	76	77	49	62	46	71	59	62	72
Hardness	4	4	3	3	3	4	3	4	4
Jn	5	9	11	8	11	7	10	8	8
JRC	11	12	11	11	12	10	11	12	11
Joint Condition	20	21	18	19	19	20	20	20	20
Jr	3	3	3	3	3	3	3	3	3
Ja	2	2	3	3	3	3	3	3	2
Representative RMR	72	74	54	62	53	69	64	64	70
St. Dev.	13	11	13	15	12	12	15	13	11
Worst RMR	76	76	50	59	48	69	67	63	75
St. Dev.	10	11	16	18	17	13	17	16	5
Representative Q'	49.5	20.1	7.4	23.3	8.3	34.4	14.6	12.3	22.4
St. Dev.	56.4	17.6	8.7	23.0	12.3	50.0	16.6	17.9	19.1
Worst Q'	52.2	32.0	8.0	22.4	8.9	21.4	21.5	13.6	31.8
St. Dev.	47.3	15.9	9.3	23.8	14.8	19.3	18.2	13.9	16.6

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

Table 11: Domains Representative Q’ Statistics
Domain	Elevation	Percentage of Q' Value Range								Design Q' (Rep. >60%)	Design Q' (Rep. >50%)	Reference Figure
		Exceptionally Poor	Extremely Poor	Very Poor	Poor	Fair	Good	Very Good	Extremely Good
		0.001 to 0.01	0.01 to 0.1	0.1 to 1	1 to 4	4 to 10	10 to 40	40 to 100	> 100
D1: Ore	Above 580 m	0.0	0.0	4.7	2.3	16.3	30.2	32.6	14.0	22	29	Figure 2
	Below 580 m	0.0	0.0	0.0	18.9	24.3	35.1	21.6	0.0	9	11
D2: HW' 6-60 m		2.2	0.0	9.6	35.4	30.9	20.8	1.1	0.0	3	4	Figure 3
D3: HW 0-6 m	Above 580 m	0.0	0.0	5.6	22.2	11.1	33.3	27.8	0.0	9 (13)	12 (21*)	Figure 4
	Below 580 m	0.0	0.0	14.3	42.9	21.4	14.3	7.1	0.0	2	3
D4: FW' 6-60 m	Above 580 m	2.6	0.0	2.6	10.4	24.7	36.4	11.7	11.7	9	13	Figure 5
	Below 580 m	0.8	0.0	9.8	21.8	22.6	36.8	8.3	0.0	11	13
D5: FW 0-6 m	Above 580 m	0.0	0.0	9.1	18.2	45.5	18.2	9.1	0.0	5	7	Figure 6
	Below 580 m	0.0	0.0	0.0	22.2	16.7	44.4	16.7	0.0	10	14
* Data from the boreholes GT-12-12, SE-12-51 and SERC-12-32 were excluded from the analysis.

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

Table 12: Domains Representative RMR’ Statistics
Domain	Elevation	Percentage of RMR’ Value Range								Design RMR’ (Rep. > 60%)	Design RMR’ (Rep. > 50%)	Reference Figure
		Very Poor	Poor		Fair		Good		Very Good
		< 20	21 to 30	31 to 40	41 to 50	51 to 60	60 to 70	70 to 80	81 to 100
D1: Ore	Above 580 m	0.0	0.0	0.0	9.3	9.3	16.3	39.5	25.6	72	75	Figure 2
	Below 580 m	0.0	0.0	0.0	8.1	5.4	5.4	54.1	27.0	75	76
D2: HW' 6-60 m		0.0	3.4	6.2	38.8	14.0	25.3	12.4	0.0	47	52	Figure 3
D3: HW 0-6 m	Above 580 m	0.0	0.0	16.7	11.1	11.1	16.7	38.9	5.6	61 (69*)	67 (71*)	Figure 4
	Below 580 m	0.0	7.1	0.0	35.7	35.7	14.3	7.1	0.0	50	52
D4: FW' 6-60 m	Above 580 m	0.0	0.0	0.0	10.3	14.1	30.8	37.2	7.7	68	70	Figure 5
	Below 580 m	0.0	2.3	1.5	23.3	13.5	21.1	27.8	10.5	60	67
D5: FW 0-6 m	Above 580 m	0.0	0.0	0.0	27.3	0.0	54.5	9.1	9.1	63	67	Figure 6
	Below 580 m	0.0	0.0	0.0	5.6	16.7	16.7	55.6	5.6	70	72
* Data from the boreholes GT-12-12, SE-12-51 and SERC-12-32 were excluded from the analysis.

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

FIGURES

Figure 1 Overview of mine and 2011 borehole locations

Figure 2 Western and Eastern ramp and relevant boreholes

Figure 3 Joint sets for footwall development

Figure 4 Tunnel support chart

Figure 5 UnWedge output plots

Figure 6 UnWedge output plots

Figure 7 Kinematic stability analysis results

Figure 8 Kinematic stability analysis results

Figure 9 Q value distribution for ramps

Figure 10 Schematic of domains for stope design

Figure 11 Domains Average Q’ Per Boreholes

Figure 12 Geotechnical Properties Distribution at Domain D1 (Ore)

Figure 13 Geotechnical Properties Distribution at Domain D2 (HW’ 6-60 m)

Figure 14 Geotechnical Properties Distribution at Domain D3 (HW 0-6 m)

Figure 15 Geotechnical Properties Distribution at Domain D4 (FW’ 6-60 m)

Figure 16 Geotechnical Properties Distribution at Domain D5 (FW 0-6 m)

Figure 17 Pillar Stability Chart

 

SANTA ELENA MINE UNDERGROUND PRE-FEASIBILITY LEVEL GEOTECHNICAL ROCK MECHANICS EVALUATION PRELIMINARY ANALYSIS EBA FILE: V15101045-011 | JULY 11, 2013 | ISSUED FOR USE

APPENDIX A

UNIAXIAL COMPRESSIVE STRENGTH AND BRAZILIAN TESTING CARRIED OUT AT SONORA UNIVERSITY, MEXICO AND UBC, VANCOUVER

EBA: St. Elena Project

UBC Laboratory Testing Report

Report No: 13_001

June 14, 2013

LABORATORY TESTING

June 3, 2013

Testing date:

May, 2013

Tested by:

Sheila Ballantyne

Tested at:

Norman B Keevil Institute of Mining Engineering, University of British Columbia

Testing Program:

•   Four (4) Ultimate Compressive Strength (UCS) Tests

•   Three (3) Brazilian Tensile Tests

Equipment:

MTS Electro-Hydraulic Testing Machine

Procedures:

Samples were wet cut, polished, and tested per the following standards:

•   UCS: ASTM D2938

•   Brazilian Tensile : ASTM D3967

Summary

Table One: Uniaxial Compressive Strength Test Results

				Height	Diameter	Ratio	UCS
UBC ID	Borehole	Depth	Description	H	d	H/d
		(m)			(mm)		(MPa)	(psi)
U1	SE-12-64	320.08 - 320.28	Andesite	157.51	60.62	2.60	199.2	28877
U2	SE-12-107	238.73 - 239.06	Andesite	154.85	60.48	2.56	131.9	19126
U3	SE-13-145	275.81 - 276.00	Quartzite	149.07	62.91	2.37	82.2	11919
U4	SE-13-145	263.03 - 263.25	Quartzite	150.78	63.00	2.39	91.2	13218

Table Two: Brazilian Tensile Test Results

		Diameter	Thickness	Ratio	Peak Load		Tensile Strength
Sample #		(d)	(t)
		(mm)	(mm)	t/d	(kN)	(lbs)	(MPa)	(psi)
	B1A	63.00	20.34	0.32	8.75	1269	4.35	406.7
B1	B1B	63.01	18.61	0.30	11.95	1733	6.49	606.8
	B1C	62.84	19.10	0.30	16.49	2391	8.75	818.1
	B2A	62.99	16.78	0.27	16.08	2331	9.68	905.8
B2	B2B	63.04	16.48	0.26	8.85	1283	5.42	507.1
	B2C	63.11	17.35	0.27	11.76	1705	6.84	639.5
	B3A	62.83	16.27	0.26	9.54	1383	5.94	555.8
B3	B3B	62.75	19.58	0.31	9.32	1351	4.83	451.7
	B3C	62.85	17.69	0.28	8.96	1299	5.13	479.9

APPENDIX I: UNIAXIAL COMPRESSIVE STRENGTH TEST RESULTS

APPENDIX II: BRAZILIAN TENSILE TEST RESULTS

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL 30, 2013 | RELEASED DATE: JULY 25, 2013

APPENDIX F

ECONOMIC ANALYSIS

SilverCrest - Santa Elena Expansion
Financial Inputs
Item	Value	Notes
Gold price used - USD	$1,450.00
Gold price for Sandstorm Deliveries	$350.00
Silver price used - USD	$28.00
Ag:Au ratio	51.79

Costs
Item	Cost	Units
Decline development	$2,300.00	$/m	SilverCrest, Quotes with $500 for equipping
Raise development by raise bore	$2,104.00	$/m	Costmine
Conventional raise development (for fresh air raises)	$1,300.00	$/m	SilverCrest Financials
Drift development by contractor (initial phase)	$2,300.00	$/m	SilverCrest, Quotes with $500 for equipping
Drift development by Santa Elena (year 5 onwards)	$1,950.00	$/m	EBA/Costmine
Open pit Ore	$2.90	$/tonne	SilverCrest Financials
Open pit Waste	$2.15	$/tonne	SilverCrest Financials
Heap leach rehandle cost	$0.62	$/tonne	EBA/SilverCrest
Tailings handling cost	$1.00	$/tonne	EBA
Processing costs ore	$21.50	$/tonne	John Fox
Processing costs heap leach reprocess	$19.50	$/tonne	John Fox
Refining and smelting costs AU	$3.40	$/Oz. Au	EBA/SilverCrest
Refining and smelting costs AG	$0.29	$/Oz. Ag	SilverCrest Financials
G and A costs per tonne	$3.75	$/tonne	Average over LOM
Monthly G and A	$304,229.89	$/month	SilverCrest Financials 2013
Annual G and A	$3,650,758.69	$/year	SilverCrest Financials 2013
Metal in dore	95%	%	EBA
Process recoveries
Underground ore AU	92%		John Fox
Underground ore Ag	68%		John Fox
Heap leach ore AU	92%		SilverCrest
Heap leach ore AG	68%		SilverCrest
Conversion factors
Grams per troy ounce	31.1	g/Oz

Assumptions

3 year trailing average gold price and silver price as of 30 April 2013 has been used as reference and deducted to $1450

Capital costs relating to the expansion plan are included as of Apil 30th 2013, and production is considered to be all production relating to the processing plant from commissioning in January 2014

No inflation of costs or revenues has been used in the model

Revenues are only calculated on gold equivalent grades

Contractor development and haulage of ore has been considered

Contractor development will continue to year 2 during which a jumbo will be purchased to do ongoing development

Ore haulage by contractor has been included for life of mine

Contractor development rates are estimated at $3,000/m based on Santa Elena cost information for contractor

In house development is estimated at $1,950/m based on Costmine and cost models - EBA expects that SE will purchase a jumbo in year 3 and after that the development cost will be an average of $3000 and $900/m

Conventional Raise development costs are estimated at $1300/m, based on costmine, Santa Elena numbers which show a range from $417 to $1700 for raise development

Raise boring is estimated at $2,104 /m

Mining costs are based on cost modelling

G and A costs are based on Santa Elena financial reports

Contingencies on capital
0%	Sunk costs and current commitments
5%	On quotes requiring no additional engineering, current Santa Elena budgets
10%	Budgetary estimates requiring no additional engineering
15%	Material estimates with conservative historic cost correlation such as ramp and drift development costs
20%	Material estimates using cost modelling to determine unit costs
25%	Material estimates using cost references based on limited engineering work
30%	Estimates requiring based on cost data requiring further engineering
35%	Estimates requiring requiring further engineering based on experience only

Santa Elena Project
Estimate of Income Tax					>> Begin							End <<
Description	Units	Total	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
		Average	-2	-1	1	2	3	4	5	6	7	8	9
Production Summary (Production from UG, Heap Leach Pad and Open Pit)
RoM Ore to Process	kt	7,309,720		0	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	253,720
Grade:
Gold	g/t	1.2			1.2	1.3	1.0	0.9	0.9	1.8	1.3	1.5
Silver	g/t	76			61	73	64	80	71	89	83	128
Contained Metal
Gold	oz	285,586		0	39,319	43,050	31,915	29,954	30,352	57,996	40,799	12,202
Silver	oz	17,953,964		0	1,992,961	2,362,434	2,088,354	2,581,838	2,306,469	2,890,783	2,690,545	1,040,581
Recovered Metal
Gold	oz	262,739		0	36,173	39,606	29,361	27,557	27,924	53,357	37,535	11,226
Silver	oz	12,118,926		0	1,345,248	1,594,643	1,409,639	1,742,741	1,556,867	1,951,279	1,816,118	702,392
Metal Deductions
Gold	0.002	(525)	0	0	(72)	(79)	(59)	(55)	(56)	(107)	(75)	(22)
Silver	0.0015	(18,178)	0	0	(2,018)	(2,392)	(2,114)	(2,614)	(2,335)	(2,927)	(2,724)	(1,054)
Payable Metal
Gold Spot	Oz	231,214	0	0	28,881	31,622	23,442	17,488	27,868	53,250	37,460	11,204	0
Gold Sandstorm	Oz	31,000		0	7,220	7,905	5,861	10,014	0	0	0	0
Silver	Oz	12,100,747	0	0	1,343,231	1,592,251	1,407,525	1,740,126	1,554,531	1,948,352	1,813,393	701,339	0
Estimate of Cash Flow
Market Prices			0
Gold Spot	1.0	$1,450		$1,450	$1,450	$1,450	$1,450	$1,450	$1,450	$1,450	$1,450	$1,450	$1,450
Gold Sandstorm		$350		$350	$350	$350	$350	$350	$350	$350	$350	$350	$350
Silver	1.0	$28.00		$28	$28	$28	$28	$28	$28	$28	$28	$28	$28
Net Revenue
Gold Sales	$000s	335,260	0	0	41,877	45,851	33,991	25,358	40,409	77,212	54,316	16,245	0
Gold sales Sandstorm	$000s	10,850		0	2,527	2,767	2,051	3,505	0	0	0	0
Silver Sales	$000s	338,821	0	0	37,610	44,583	39,411	48,724	43,527	54,554	50,775	19,637	0
NSR	$000s	684,931	0	0	82,014	93,201	75,453	77,586	83,936	131,766	105,091	35,883	0
Freight & Refining	$000s	(5,579)		0	(626)	(739)	(650)	(797)	(714)	(893)	(837)	(321)	0
Gross Income	$000s	679,353	0	0	81,389	92,462	74,803	76,789	83,221	130,873	104,255	35,561	0
					512	501	623	586	652	499	619	640
Operating Costs	1.0	282,223		0	31,816	35,254	35,232	35,874	37,793	45,465	44,931	15,857	0
Operating Profit	$000s	397,130	0	0	49,573	57,207	39,571	40,915	45,428	85,408	59,324	19,704	0
					512.8371	501.6717	623.7741	587.0940	652.8779	500.3112	619.9327	640.7885
Cash Flow
Operating Margin	$000s	397,130	0	0	49,573	57,207	39,571	40,915	45,428	85,408	59,324	19,704	0
Project Capital	$000s	(87,813)	0	(48,068)	(6,272)	(3,606)	(11,555)	(10,982)	(5,380)	(553)	(1,221)	(176)	0
Mine Closure	$000s	(5,000)											(5,000)
Employee Profit Share	$000s	0	0	0	0	0	0	0	0	0	0	0	0
Income Tax	$000s	(100,839)	0	0	(13,242)	(15,424)	(9,786)	(9,860)	(11,053)	(23,030)	(15,168)	(3,277)	0
Working Capital	$000s	(1,836)	0	0	(1,836)	0	0	0	0	0	0	0	0
Post-Tax Cash Flow	$000s	201,642	0	(48,068)	28,223	38,178	18,230	20,073	28,996	61,825	42,935	16,252	(5,000)

Santa Elena Project
Estimate of Income Tax					>> Begin							End <<
Description	Units	Total	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
		Average	-2	-1	1	2	3	4	5	6	7	8	9
Cumulative	$000s	-	0	(48,068)	(19,845)	18,333	36,563	56,635	85,631	147,456	190,390	206,642	201,642
Present Value	5%	145,566		(45,779)	25,599	32,979	14,998	15,727	21,637	43,938	29,060	10,476	(3,070)
NPV	$000s	-	0	(45,779)	(20,180)	12,800	27,797	43,525	65,162	109,099	138,159	148,635	145,566
IRR	%	60%
Payback	months	18				2,015.52
"	years	1.5
Peak Funding	$000s	(48,068)
PROJECT CAPITAL
Project Capital
infrastructure	$000s	0		0	0	0	0	0	0	0	0	0	0
Mining Equipment	$000s	7,488		2,344	1,345	710	158	2,564	92	92	92	92	0
Development - UG	$000s	34,765		3,481	2,969	2,693	10,857	8,150	5,193	378	1,046	0	0
Process Plant	$000s	41,928		41,928	0	0	0	0	0	0	0	0	0
Electrical & Svcs - UG	$000s	2,355		72	1,249	95	432	160	95	84	84	84	0
Health & safety - UG	$000s	143		0	143	0	0	0	0	0	0	0	0
Construction - UG	$000s	323		0	323	0	0	0	0	0	0	0	0
Salvage value													0
Indirect	$000s	810		243	243	108	108	108	0	0	0	0	0
subtotal	$000s	87,813	0	48,068	6,272	3,606	11,555	10,982	5,380	553	1,221	176	0
Contingncy	$000s	9,305		5,071	680	382	1,222	1,173	570	60	129	19	0
Total Capital	1.0	87,813	0	48,068	6,272	3,606	11,555	10,982	5,380	553	1,221	176	0
Working Capital	$000s	1,836		0	1,836	0	0	0	0	0	0	0	0
ESTIMATE OF TAX
Income Tax
NSR	$000s	684,931	0	0	82,014	93,201	75,453	77,586	83,936	131,766	105,091	35,883	0
Freight & Marketing	$000s	(5,579)	0	0	(626)	(739)	(650)	(797)	(714)	(893)	(837)	(321)	0
Operating Costs	$000s	(282,223)	0	0	(31,816)	(35,254)	(35,232)	(35,874)	(37,793)	(45,465)	(44,931)	(15,857)	0
Operating Profit	$000s	397,130	0	0	49,573	57,207	39,571	40,915	45,428	85,408	59,324	19,704	0
Depreciation & Amort.	$000s	(84,897)	0	0	(5,434)	(5,795)	(6,950)	(8,048)	(8,586)	(8,642)	(8,764)	(8,781)	(8,781)
Interest Expense	$000s	0	100% Equity Assumed. No interest or fees.
Fees	$000s	0
Net Income	$000s	312,233	0	0	44,139	51,413	32,621	32,867	36,842	76,766	50,560	10,923	(8,781)
Loss Carry-Forward
Additions	$000s	23,897	0	0	0	0	0	0	0	0	0	0	8,781
Opening Balance	$000s	71,151	0	0	0	0	0	0	0	0	0	0	8,781
Losses Used	$000s	0	0	0	0	0	0	0	0	0	0	0	0
Closing Balance	$000s	71,151	0	0	0	0	0	0	0	0	0	0	8,781
Loss Carry-Forward	$000s	23,897	0	0	0	0	0	0	0	0	0	0	8,781
Employee Profit Share	0.0%	0	0	0	0	0	0	0	0	0	0	0	0

Santa Elena Project
Estimate of Income Tax					>> Begin							End <<
Description	Units	Total	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
		Average	-2	-1	1	2	3	4	5	6	7	8	9
Taxable Income	$000s	336,130	0	0	44,139	51,413	32,621	32,867	36,842	76,766	50,560	10,923	0
Taxable Income	30%	(100,839)	0	0	(13,242)	(15,424)	(9,786)	(9,860)	(11,053)	(23,030)	(15,168)	(3,277)	0
Depreciation
Period	1				1	2	3	4	5	6	7	8	9
Depreciable Costs	2	0
infrastructure	3	0	0	0	0	0	0	0	0	0	0	0	0
Mining Equipment	4	7,488	0	2,344	1,345	710	158	2,564	92	92	92	92	0
Development - UG	5	34,765	0	3,481	2,969	2,693	10,857	8,150	5,193	378	1,046	0	0
Process Plant	6	41,928	0	41,928	0	0	0	0	0	0	0	0	0
Electrical & Svcs - UG	7	2,355	0	72	1,249	95	432	160	95	84	84	84	0
Health & safety - UG	8	143	0	0	143	0	0	0	0	0	0	0	0
Construction - UG	9	323	0	0	323	0	0	0	0	0	0	0	0
Indirect	10	810	0	243	243	108	108	108	0	0	0	0	0
Depreciable Costs	12	87,813			54,340	3,606	11,555	10,982	5,380	553	1,221	176	0
Straight-Line (10-yr Basis)	1				1	2	3	4	5	6	7	8	9
1		100%			10.0%	10.0%	10.0%	10.0%	10.0%	10.0%	10.0%	10.0%	10.0%
2		100%				10.0%	10.0%	10.0%	10.0%	10.0%	10.0%	10.0%	10.0%
3		100%					10.0%	10.0%	10.0%	10.0%	10.0%	10.0%	10.0%
4		90%						10.0%	10.0%	10.0%	10.0%	10.0%	10.0%
5		80%							10.0%	10.0%	10.0%	10.0%	10.0%
6		70%								10.0%	10.0%	10.0%	10.0%
7		60%									10.0%	10.0%	10.0%
8		50%										10.0%	10.0%
9		40%											10.0%
10		30%
11		20%
12		10%
Straight-Line (10-yr Basis)	1				1	2	3	4	5	6	7	8	9
1		54,340			5,434	5,434	5,434	5,434	5,434	5,434	5,434	5,434	5,434
2		3,606	0	0	0	361	361	361	361	361	361	361	361
3		11,555	0	0	0	0	1,155	1,155	1,155	1,155	1,155	1,155	1,155
4		9,884	0	0	0	0	0	1,098	1,098	1,098	1,098	1,098	1,098
5		4,304	0	0	0	0	0	0	538	538	538	538	538
6		387	0	0	0	0	0	0	0	55	55	55	55
7		733	0	0	0	0	0	0	0	0	122	122	122
8		88	0	0	0	0	0	0	0	0	0	18	18
9		0	0	0	0	0	0	0	0	0	0	0	0
10		0	0	0	0	0	0	0	0	0	0	0	0
11		0	0	0	0	0	0	0	0	0	0	0	0
12		0	0	0	0	0	0	0	0	0	0	0	0
Depreciation		84,897	0	0	5,434	5,795	6,950	8,048	8,586	8,642	8,764	8,781	8,781

Period		2013	2014	2015	2016	2017	2018	2019	2020	2021	Total	Averages
From Underground Schedule
Ramp development	m	742	0	542	1758	0	584	617	0	0	3,501	438
Waste f/w development - to capital	m	436	506	476	1573	2221	1593	155	429	0	6,953	869
Total waste development - to capital	m	1178	506	1018	3331	2221	2177	772	429	0	10,454	1,307
Fresh air raises - to capital		89	-	65	211	-	70	74	-	-	420	53
Exhaust vent raise -  to capital		248				198					198	198
											�� -
Total development m		1178	2158	3240	3914	3250	3007	772	429		16,770	2,396
											-
Total stope development as waste											-
Stope development as ore	m		1,652	2,222	583	1,029	830	-	-	-	6,316	790
Stope dev tonnes	tonnes		71,863	66,343	12,016	149,270	183,731	-	-	-	483,223	60,403
Stope dev grade AU	g/t		2.58	2.38	1.27	1.07	1.05	2.32	1.70		12	2
Stope dev grade AG	g/t		138.33	134.99	91.90	105.63	87.18	115.59	119.64	-	793	99
											-
Tonnes mined in stopes	tonnes		55,844	326,069	523,504	496,818	524,628	684,604	580,407	245,225	3,437,100	429,638
Stope grade AU	g/t		2.58	2.38	1.27	1.07	1.05	2.32	1.70	1.52	14	2
Stope grade AG	g/t		138	135	92	106	87	116	120	131	924	116
Grams AU	tonnes		144,123	774,910	665,435	532,656	552,647	1,590,241	986,625	373,884	5,620,520	702,565
Stope cost	tonnes		$990,600	$5,670,597	$9,220,707	$8,684,122	$11,265,203	$20,588,378	$20,198,107	$8,588,776	85,206,490	10,650,811
Underground mining
Total ore from underground	tonnes	-	127,707	392,412	535,520	646,088	708,359	684,604	580,407	245,225	3,920,323.35	490,040
Days underground mining			150	360	360	360	360	360	360	240	2,550.00
Daily production rate	tpd		851	1,090	1,488	1,795	1,968	1,902	1,612	1,022	11,727.02
AU Grade	g/t	-	2.58	2.38	1.27	1.07	1.05	2.32	1.70	1.52	13.90	1.7
AG Grade	g/t		138	135	92	106	87	116	120	131	924.07
AU Ounces mined	Oz.	0	10,598	29,986	21,888	22,273	23,993	51,133	31,724	12,022	203,617.54	25,452
AU Ounces recovered	Oz.	0	9,750	27,587	20,137	20,491	22,074	47,042	29,186	11,060	187,328.14	23,416
AG Ounces mined	Oz.		568,033	1,703,300	1,582,451	2,194,325	1,985,631	2,544,511	2,232,704	1,031,485	13,842,439.93	1,730,305
AG Ounces recovered	Oz.		383,422	1,149,727	1,068,155	1,481,169	1,340,301	1,717,545	1,507,075	696,252	9,343,646.95
											-
Mining cost	$	$0	$990,600	$5,670,597	$9,220,707	$8,684,122	$11,265,203	$20,588,378	$20,198,107	$8,588,776	85,206,490.00	10,650,811
Ore dev cost	$	$0	$3,799,600	$5,110,600	$1,340,900	$2,366,700	$1,618,500	$0	$0	$0	14,236,300.00	1,779,538
											-
Processing cost	$	$0	$2,745,707	$8,436,860	$11,513,678	$13,890,901	$15,229,718	$14,718,991	$12,478,760	$5,272,337	84,286,951.93	10,535,869
											-
G and A	$		$3,650,759	$3,650,759	$3,650,759	$3,650,759	$3,650,759	$3,650,759	$3,650,759	$1,825,379	27,380,690.21	3,422,586
G and A / tonne			$3.6	$3.6	$3.6	$3.6	$3.6	$3.6	$3.6	$7.2	3.75	4.07
											-
Total operating cost underground	$	$0	$11,186,666	$22,868,816	$25,726,044	$28,592,482	$31,764,180	$38,958,128	$36,327,625	$15,686,492	211,110,432.14	26,388,804
Underground operating cost per tonne			$88	$58	$48	$44	$45	$57	$63	$64	53.85	58

Period		2013	2014	2015	2016	2017	2018	2019	2020	2021	Total	Averages
Reprocess ore
												2,844,530
Heap leach ore	Tonnes		335,426.16	615,587.90	472,480.07	361,911.60	299,641.02	323,395.76	427,592.58	8,494.92	2,844,530.00	355,566
Heap leach grade AU	g/t		0.66	0.66	0.66	0.66	0.66	0.66	0.66	0.66	5.28
Heap leach grade AG	g/t		33.3	33.3	33.3	33.3	33.3	33.3	33.3	33.3	266.40
Mineable ounces AU	Oz.		7,118	13,064	10,027	7,680	6,359	6,863	9,074	180	60,366.23	7,546
Recoverable ounces AU	Oz.		6,549	12,019	9,225	7,066	5,850	6,314	8,348	166	55,536.93	6,942
Mineable ounces Ag	Oz.		359,154	659,134	505,903	387,513	320,837	346,273	457,840	9,096	3,045,750.77	380,719
Recoverable ounces Ag	Oz.		242,428.98	444,915.66	341,484.59	261,571.32	216,565.31	233,734.02	309,042.19	6,139.70	2,055,881.77	256,985
											-
Mining cost	$		$207,964	$381,664	$292,938	$224,385	$185,777	$200,505	$265,107	$5,267	1,763,608.60	220,451
											-
Processing cost	$		$6,540,810	$12,003,964	$9,213,361	$7,057,276	$5,843,000	$6,306,217	$8,338,055	$165,651	55,468,335.00	6,933,542
											-
Total operating cost Heap	$	$0	$6,748,774	$12,385,629	$9,506,299	$7,281,661	$6,028,777	$6,506,723	$8,603,163	$170,918	57,231,943.60	20
Open pit mining
Open pit tonnes		881,843	544,867								1,426,710.00
Open pit waste 0.5:1 for 2014		1,317,600	272,433								1,590,033.42
Open pit Au grade			1								1.23
AU mineable oz from Open pit			21,603	-	-	-					21,602.55
Au recoverable oz from open pit			19,874	-	-	-					19,874.35
Open pit Ag grade			61								60.91
Ag mineable oz from Open pit			1,065,773								1,065,773.50
Ag recoverable oz from open pit			719,397								719,397.11
Mining cost ore			$1,580,113	-	-	-					1,580,113.00
Mining cost waste			$585,732	-	-	-					585,731.54
											-
Processing cost			$11,714,631	-	-	-					11,714,630.86
Total operating cost open pit			$13,880,475	-	-	-					25.475

Period	2013	2014	2015	2016	2017	2018	2019	2020	2021	Total	Averages
Summary
Total tonnes from underground	-	127,707	392,412	535,520	646,088	708,359	684,604	580,407	245,225	3,920,323
Total tonnes from old heap leach	-	335,426	615,588	472,480	361,912	299,641	323,396	427,593	8,495	2,844,530
Total tonnes from open pit	881,843	544,867	-	-	-	-	-	-	-	544,867
Total tonnes processed		1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	1,008,000	253,720	7,309,720	913,715
Tonnes per day		2,800	2,800	2,800	2,800	2,800	2,800	2,800	705	20,305	2,538
Total AU ounces mined	-	39,319	43,050	31,915	29,954	30,352	57,996	40,799	12,202	285,586	35,698
Total AU ounces recovered		36,173	39,606	29,361	27,557	27,924	53,357	37,535	11,226	262,739	32,842
Total AU ounces sold		36,101	39,527	29,303	27,502	27,868	53,250	37,460	11,204	262,214	32,777
Total Ag ounces mined		1,992,961	2,362,434	2,088,354	2,581,838	2,306,469	2,890,783	2,690,545	1,040,581	17,953,964	2,244,246
Total Ag ounces recovered		1,345,248	1,594,643	1,409,639	1,742,741	1,556,867	1,951,279	1,816,118	702,392	12,118,926	1,514,866
Total AG ounces sold		1,343,231	1,592,251	1,407,525	1,740,126	1,554,531	1,948,352	1,813,393	701,339	12,100,747	1,512,593
Total AU EQ Ounces mined		77,803	88,670	72,241	79,810	74,891	113,818	92,754	32,296	632,284	79,035
Total AU EQ Ounces recovered		62,150	70,399	56,582	61,210	57,988	91,036	72,605	24,790	496,760	62,095
Total AU EQ Ounces sold		62,039	70,274	56,483	61,105	57,887	90,873	72,477	24,747	495,884	61,985
Total AG EQ Ounces mined		4,029,103	4,591,822	3,741,077	4,133,004	3,878,279	5,894,158	4,803,329	1,672,483	32,743,256	4,092,907
Total AG EQ Ounces recovered		3,218,500	3,645,680	2,930,144	3,169,813	3,002,932	4,714,383	3,759,879	1,283,742	25,725,074	3,215,634
Total AG EQ Ounces sold		3,212,735	3,639,186	2,924,989	3,164,345	2,997,705	4,705,930	3,753,267	1,281,526	25,679,684	3,209,960
Total operating cost	0	31,815,916	35,254,445	35,232,343	35,874,143	37,792,957	45,464,850	44,930,788	15,857,410	282,222,851	35,277,856
Total operating cost per tonne processed	-	32	35	35	36	37	45	45	62	38.61	41
Total operating cost per ounce AU EQ recovered		512	501	623	586	652	499	619	640	568
Total operating cost per ounce AG EQ recovered		10	10	12	11	13	10	12	12	11
Total operating cost per ounce AU EQ sold		513	502	624	587	653	500	620	641	569
Total operating cost per ounce AG EQ sold		10	10	12	11	13	10	12	12	11
Working capital
Working capital - 3 weeks of op costs	$0.00	$1,835,533.59	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00		-
Working capital adjustment		$1,835,533.59	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$0.00	$1,835,533.59	229,442

					Year -2	Year -1	Yea 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Area	Capital expense item	Freight per unit	Total Capital	Schedule check	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
Total capital life of mine			87,813,477	87,813,477		48,067,874	6,272,455	3,605,930	11,554,820	10,982,336	5,379,848	553,384	1,221,259	175,571	-
Summary	Initial		48,067,874	-
	General infrastructure			-		-
	Mining equipment purchases		2,343,810	2,343,810		2,343,810
	Development - UG		3,480,550	3,480,550		3,480,550
	Process plant		41,928,472	41,928,472		41,928,472
	Electrical and services - UG		72,042	72,042		72,042
	Health and safety - UG		-	-		-
	Construction (incl. labour) - UG		-	-		-
	Indirect		243,000	243,000		243,000
	Sustaining		39,745,603	-
	General infrastructure			-			-	-	-	-	-	-	-	-	-
	Mining equipment purchases		5,144,587	5,144,587			1,344,788	710,030	158,000	2,563,769	92,000	92,000	92,000	92,000	-
	Development		31,284,348	31,284,348			2,968,901	2,692,610	10,856,503	8,150,276	5,192,558	377,813	1,045,688	-	-
	Process plant		-	-			-	-	-	-	-	-	-	-	-
	Electrical and services		2,283,346	2,283,346			1,249,444	95,290	432,317	160,290	95,290	83,571	83,571	83,571	-
	Health and safety		143,442	143,442			143,442	-	-	-	-	-	-	-	-
	Construction (incl. labour)		322,880	322,880			322,880	-	-	-	-	-	-	-	-
	Indirect		567,000	567,000			243,000	108,000	108,000	108,000	-	-	-	-	-
	Contingency		9,304,765	9,304,765		5,070,754	679,656	382,400	1,221,782	1,172,890	569,533	60,395	128,832	18,521	-
	Budget spent as of JAN 2012 TO APRIL 30 2013		24,263,480

Area	Capital expense item	Got quote	Quantity	Unit cost	Sunk costs	Contingency applied to estimate	Contingency amount	Freight per unit
	Long hole drill	Quote	1	1,055,000		5%	52,750
	Scoop trams	Quote	3	588,600		5%	88,290
	Jackleg/stopers	Costmine	4	13,000		25%	13,000
	Jumbo drills (cut and fill stopes) DD321-40 (110kW)	Quote	2	993,000		5%	99,300
	Jumbo accessories	Quote	2	117,100		30%	70,260
General	Primary ventilation	Quote	5	31,542		30%	47,313
infrastructure	Auxiliary fans	Quote	10	12,690		5%	6,345
	Anfo loader	Quote	1	27,300		10%	2,730
	Scissor lift 92 HP	Quote	1	105,000		10%	10,500
	Grader - 110 HP	Quote	1	190,000		10%	19,000
	Kubota/s	Website	3	25,000		10%	7,500
	Crew transport -  Miller	Quote	2	60,000		10%	12,000
	Service truck - Miller	Quote	1	79,000		10%	7,900
	Shotcrete plant	Cost mine	1	16,632		25%	4,158
	Underground diamond drill/Gopher	Cost mine	2	88,585		25%	44,293
	Compressor	Cost mine	1	48,778		25%	12,194
	Pipe fusing machine	Cost mine	1	30,700		25%	7,675
	Ongoing mining related capex	SE current	8	80,000		15%	96,000
	Cement storage silo for backfill	Cost mine	1	72,900		25%	18,225

				Year -2	Year -1	Yea 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Area	Capital expense item	Total Capital	Schedule check	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
Total capital life of mine		87,813,477	87,813,477		48,067,874	6,272,455	3,605,930	11,554,820	10,982,336	5,379,848	553,384	1,221,259	175,571	-
						-
						-
	Long hole drill	1,107,750	1,107,750		1,107,750
	Scoop trams	1,854,090	1,854,090		1,236,060		618,030
	Jackleg/stopers	65,000	65,000			65,000
	Jumbo drills (cut and fill stopes) DD321-40 (110kW)	2,085,300	2,085,300						2,085,300
	Jumbo accessories	304,460	304,460						304,460
	Primary ventilation	205,023	205,023			123,014			82,009
	Auxiliary fans	133,245	133,245			133,245
General	Anfo loader	30,030	30,030			30,030
infrastructure	Scissor lift 92 HP	115,500	115,500			115,500
	Grader - 110 HP	209,000	209,000			209,000
	Kubota/s	82,500	82,500			82,500
	Crew transport -  Miller	132,000	132,000			66,000		66,000
	Service truck - Miller	86,900	86,900			86,900
	Shotcrete plant	20,790	20,790			20,790
	Underground diamond drill/Gopher	221,463	221,463			221,463
	Compressor	60,972	60,972			60,972
	Pipe fusing machine	38,375	38,375			38,375
	Ongoing mining related capex	736,000	736,000			92,000	92,000	92,000	92,000	92,000	92,000	92,000	92,000
	Cement storage silo for backfill	91,125	91,125						91,125
		-	-			-

Area	Capital expense item	Got quote	Quantity	Unit cost	Sunk costs	Contingency applied to estimate	Contingency  amount	Freight per unit
Total capital life of mine
	Development	34,764,898
	Ramp year -2		1,005		5,300,000
	Ramp Year -1		742	2,300		5%	85,330
	Ramp Year 1		-	2,300		15%	-
	Ramp Year 2		542	2,300		15%	186,990
	Ramp Year 3		1,758	2,300		15%	606,510
	Ramp Year 4		-	2,300		15%	-
	Ramp Year 5		584	1,950		15%	170,820
	Waste development year -1		436	2,300		15%	150,420
	Waste development year 1		506	2,300		15%	174,570
	Waste development year 2		476	2,300		15%	164,220
	Waste development year 3		1,573	2,300		25%	904,475
	Waste development year 4		2,221	2,300		25%	1,277,075
	Waste development year 5		1,593	1,950		25%	776,588
Development	Waste development year 6		155	1,950		25%	75,563
	Waste development year 7		429	1,950		25%	209,138
	Waste development year 8		-	1,950		25%	-
	Ventilation shaft/raise pre production		248	2,104		25%	130,489
	Ventilation shaft/raise year 4		198	2,104		25%	104,293
	Ventilation drifts year 1		211	2,300		15%	72,795
	Ventilation drifts sustaining		676	1,950		25%	329,306
	Raise bore underground station 536 level	Cost mine	1	22,499		25%	5,625
	Raise bore underground station 325 level	Cost mine	1	22,499		25%	5,625
	Pit ramp shaft	Cost mine	269	2,104		25%	141,494
	Ventilation cross cuts - pit ramp	Costmine	1	22,499		25%	5,625
	Fresh air raises year 1	Santa Elena financials	89.04	1,300		15%	17,363
	Fresh air raises year 2	Santa Elena financials	-	1,300		15%	-
	Fresh air raises year 3	Santa Elena financials	65.04	1,300		15%	12,683
	Fresh air raises year 4	Santa Elena financials	210.96	1,300		15%	41,137
	Fresh air raises year 5	Santa Elena financials	-	1,300		15%	-
	Underground shops		1	207,000		25%	51,750
	Underground definition drilling	Budget meters	3,000	50		30%	45,000

				Year -2	Year -1	Yea 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Area	Capital expense item	Total Capital	Schedule check	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
	Development
	Ramp year -2
	Ramp Year -1	1,791,930	1,791,930		1,791,930
	Ramp Year 1	-	-			-
	Ramp Year 2	1,433,590	1,433,590				1,433,590
	Ramp Year 3	4,649,910	4,649,910					4,649,910
	Ramp Year 4	-	-						-
	Ramp Year 5	1,309,620	1,309,620							1,309,620
	Waste development year -1	1,493,620	1,493,620		1,493,620
	Waste development year 1	1,338,370	1,338,370			1,338,370
	Waste development year 2	1,259,020	1,259,020				1,259,020
	Waste development year 3	4,522,375	4,522,375					4,522,375
	Waste development year 4	6,385,375	6,385,375						6,385,375
	Waste development year 5	3,882,938	3,882,938							3,882,938
	Waste development year 6	377,813	377,813								377,813
	Waste development year 7	1,045,688	1,045,688									1,045,688
Development	Waste development year 8	-	-										-
	Ventilation shaft/raise pre production	652,447	652,447			652,447
	Ventilation shaft/raise year 4	626,251	626,251						626,251
	Ventilation drifts year 1	558,095	558,095			558,095
	Ventilation drifts sustaining	1,646,531	1,646,531					823,266	823,266
	Raise bore underground station 536 level	28,124	28,124			28,124
	Raise bore underground station 325 level	28,124	28,124					28,124
	Pit ramp shaft	707,470	707,470					707,470
	Ventilation cross cuts - pit ramp	28,124	28,124					28,124
	Fresh air raises year 1	133,115	133,115			133,115
	Fresh air raises year 2	-	-				-
	Fresh air raises year 3	��      97,235	97,235					97,235
	Fresh air raises year 4	315,385	315,385						315,385
	Fresh air raises year 5	-	-							-
	Underground shops	258,750	258,750			258,750
	Underground definition drilling	195,000	195,000		195,000

Area	Capital expense item	Got quote	Quantity	Unit cost	Sunk costs	Contingency applied to estimate	Contingency amount	Freight per unit
Total capital life of mine
Processing plant
	Crushing and screening	Santa Elena	1	7,830,031	2,189,006	0	282,051
	Storage and handling	Santa Elena	1	4,060,339	471,881	0	179,423
	Grinding and classification	Santa Elena	1	9,100,639	1,340,650	0	387,999
	Leaching and CCD circuit	Santa Elena	1	7,456,735	2,717,322	0	236,971
	Recovery (MC plant)	Santa Elena	1	6,504,402	2,267,527	0	211,844
	Reagents	Santa Elena	1	2,499,638	101,571	0	119,903
	Water storage	Santa Elena	1	347,384	8,009	0	16,969
	Tailings	Santa Elena	1	5,588,977	1,235,010	0	217,698
	Auxiliaries	Santa Elena	1	1,514,271	621,636	0	44,632
	EPCM group	Santa Elena	1	6,063,897	4,352,395	0	85,575
	Gen sets	Santa Elena	1	6,191,048	2,276,429	0	195,731
	Stackers	Santa Elena	1	1,018,479	1,018,479	0	0
	Other equipment	Santa Elena	1	363,565	363,565	0	0
	Tailings conveyor	Costmine	1	325,000		0	48,750
	total						-

Electrical and services
	Surface main underground vacuum breaker	Quote	1	25,000	15%	3,750
	Main 5kV line to adit area	Quote	800	50	25%	10,000
	Surface to underground high voltage line initial	Quote	291	50	25%	3,638
	Surface to underground high voltage line sustaining	Quote	750	50	25%	9,375
	Cable to surface ventilation fans	Quote	200	��           50	25%	2,500
	Transformer at surface ventilation	Quote	1	22,000	25%	5,500
	440v underground line/s	Cost mine	1,500	50	30%	22,500
	Surface substation	SVL	1	-	30%	-
	Underground power centres incl. transformer 500kVA	Quote	4	61,000	15%	36,600
	Miscellaneous electrical equipment and lighting	N	1	50,000	35%	17,500
	Surface cap lamp battery charger	Cost mine	2	15,194	25%	7,597
	Underground water line intial	Cost mine	1,800	30	30%	16,200
	Underground water line sustaining	Cost mine	15,000	30	30%	135,000
	Underground dewater line - initial	Cost mine	250	100	30%	7,500
	Underground dewater line - sustaining	Cost mine	500	100	30%	15,000
	Underground dewater pumps centrifugal initial - 60kW	Cost mine	2	26,389	30%	15,833
	Underground dewater pumps centrifugal initial - 45kW	Cost mine	4	26,389	30%	31,667
	Submersible pumps 40 HP	Cost mine	2	22,150	30%	13,290
	Pump valves/ suction pipes etc.	Cost mine	1	5,278	30%	1,583
	Underground dewater pumps centrifugal sustaining	Cost mine	8	18,916	30%	45,398
	Trash pumps for slurry	Cost mine	3	23,720	30%	21,348
	Compressed air line	Cost mine	3,000	50	30%	45,000
	Ventilation tubing	Cost mine	1,800	35	30%	18,900
	Leaky feeder communication	Cost mine	1	76,055	30%	22,817

				Year -2	Year -1	Yea 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Area	Capital expense item	Total Capital	Schedule check	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022
Total capital life of mine		87,813,477	87,813,477		48,067,874	6,272,455	3,605,930	11,554,820	10,982,336	5,379,848	553,384	1,221,259	175,571	-
Processing plant
	Crushing and screening	5,923,076	5,923,076		5,923,076
	Storage and handling	3,767,881	3,767,881		3,767,881
	Grinding and classification	8,147,988	8,147,988		8,147,988
	Leaching and CCD circuit	4,976,384	4,976,384		4,976,384
	Recovery (MC plant)	4,448,719	4,448,719		4,448,719
	Reagents	2,517,970	2,517,970		2,517,970
	Water storage	356,344	356,344		356,344
	Tailings	4,571,665	4,571,665		4,571,665
	Auxiliaries	937,267	937,267		937,267
	EPCM group	1,797,077	1,797,077		1,797,077
	Gen sets	4,110,350	4,110,350		4,110,350
	Stackers	0	0		0
	Other equipment	0	0		0
	Tailings conveyor	373,750	373,750		373,750
	total

Surface main underground vacuum breaker	28,750	28,750		28,750
Main 5kV line to adit area	50,000	50,000		50,000
Surface to underground high voltage line initial	18,188	18,188		18,188
Surface to underground high voltage line sustaining	46,875	46,875			11,719	11,719	11,719	11,719
Cable to surface ventilation fans	12,500	12,500		12,500
Transformer at surface ventilation	27,500	27,500		27,500
440v underground line/s	97,500	97,500		97,500
Surface substation	-	-		-
Underground power centres incl. transformer 500kVA	280,600	280,600		140,300		140,300
Miscellaneous electrical equipment and lighting	67,500	67,500		67,500
Surface cap lamp battery charger	37,984	37,984		37,984
Underground water line intial	70,200	70,200		70,200
Underground water line sustaining	585,000	585,000			83,571	83,571	83,571	83,571	83,571	83,571	83,571
Underground dewater line - initial	32,500	32,500		32,500
Underground dewater line - sustaining	65,000	65,000					65,000
Underground dewater pumps centrifugal initial - 60kW	68,611	68,611	68,611
Underground dewater pumps centrifugal initial - 45kW	137,223	137,223		137,223
Submersible pumps 40 HP	57,590	57,590		57,590
Pump valves/ suction pipes etc.	6,861	6,861	3,431	3,431
Underground dewater pumps centrifugal sustaining	196,726	196,726				196,726
Trash pumps for slurry	92,508	92,508		92,508
Compressed air line	195,000	195,000		195,000
Ventilation tubing	81,900	81,900		81,900
Leaky feeder communication	98,872	98,872		98,872

Area	Capital expense item	Got quote	Quantity	Unit cost	Sunk costs	Contingency applied to estimate	Contingency amount	Freight per unit
	Hardhats	Cost mine	100	15		30%	450
	Gloves	Cost mine	100	5		30%	150
	Hearing protection	Cost mine	100	15		30%	450
Health and safety	Miners lamps	Cost mine	100	200		30%	6,000
	Respirators (boxes)	Cost mine	20	17		30%	102
	Back up alarms	Cost mine	50	120		30%	1,800
	Signs and other safety equipment	Cost mine	1	25,000		30%	7,500
	Fire extinguishers	Cost mine	25	100		30%	750
	Medicals	Cost mine	50	50		30%	750
	Safety glasses	Cost mine	100	5		30%	150
	Self rescuer	Cost mine	100	400		30%	12,000
	Boots	Cost mine	100	100		30%	3,000

	Vent shaft collar concrete	Cost mine	20	142	35%	995
	Vent shaft collar labour	Cost mine	200	25	35%	1,750
	Vent shaft steel	Cost mine	2	99	35%	69
	general underground electrical labour	Cost mine	240	25	35%	2,100
	Fan electrical labour	Cost mine	160	25	35%	1,400
	Underground ventilation doors	Cost mine	2	15,513	35%	10,859
	Ventilation walls materials	Cost mine	50	142	35%	2,487
	Ventilation walls labour	Cost mine	720	25	35%	6,300
	Refuge chambers	Cost mine	2	75,000	35%	52,500
Construction (incl. labour)	Underground machine shop	Cost mine	1	15,000	35%	5,250
	Underground pump stations concrete	Cost mine	50	142	35%	2,487
	Underground pump stations labour	Cost mine	275	25	35%	2,406
	Ladder ways in fresh air raise year 1	Cost mine	244	91	35%	7,758
	Ladder ways in fresh air raise year 2	Cost mine	155	91	35%	4,928
	Ladder ways in fresh air raise year 3	Cost mine	220	91	35%	6,995
	Ladder ways in fresh air raise year 4	Cost mine	366	91	35%	11,634
	Ladder ways in fresh air raise year 5	Cost mine	155	91	35%	4,928
	Ladder ways in fresh air raise year 6	Cost mine	156	91	35%	4,960
	General underground installations labour hrs.	Cost mine	3,600	25	35%	31,500
Indirect
	Staffing, recruiting and expatriation expenses		1	200,000	35%	70,000
	Underground design and engineering consultants		1	400,000	35%	140,000
Contingency					11%	9,304,765

				Year -2	Year -1	Yea 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8	Year 9
Area	Capital expense item	Total Capital	Schedule check	2012	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022

Health and safety
	Hardhats	1,950	1,950	1,950
	Gloves	650	650	650
	Hearing protection	1,950	1,950	1,950
	Miners lamps	26,000	26,000	26,000
	Respirators (boxes)	442	442	442
	Back up alarms	7,800	7,800	7,800
	Signs and other safety equipment	32,500	32,500	32,500
	Fire extinguishers	3,250	3,250	3,250
	Medicals	3,250	3,250	3,250
	Safety glasses	650	650	650
	Self rescuer	52,000	52,000	52,000
	Boots	13,000	13,000	13,000

	Vent shaft collar concrete	3,837	3,837		3,837
	Vent shaft collar labour	6,750	6,750		6,750
	Vent shaft steel	266	266		266
	general underground electrical labour	8,100	8,100		8,100
	Fan electrical labour	5,400	5,400		5,400
Construction (incl. labour)	Underground ventilation doors	41,885	41,885		41,885
	Ventilation walls materials	9,592	9,592		9,592
	Ventilation walls labour	24,300	24,300		24,300
	Refuge chambers	202,500	202,500		202,500
	Underground machine shop	20,250	20,250		20,250
	Underground pump stations concrete	9,592	9,592		9,592
	Underground pump stations labour	9,281	9,281		9,281
	Ladder ways in fresh air raise year 1	29,925	29,925		29,925
	Ladder ways in fresh air raise year 2	19,007	19,007			19,007
	Ladder ways in fresh air raise year 3	26,982	26,982				26,982
	Ladder ways in fresh air raise year 4	44,876	44,876					44,876
	Ladder ways in fresh air raise year 5	19,007	19,007						19,007
	Ladder ways in fresh air raise year 6	19,130	19,130							19,130
	General underground installations labour hrs.	121,500	121,500		121,500
Indirect
	Staffing, recruiting and expatriation expenses	270,000	270,000	135,000	135,000
	Underground design and engineering consultants	540,000	540,000	108,000	108,000	108,000	108,000	108,000
Contingency		9,304,765	9,304,765	5,070,754	679,656	382,400	1,221,782	1,172,890	569,533	60,395	128,832	18,521	-