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

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

AMENDED DATE: 04 MARCH, 2014

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, MBA, Senior Mineral Economist - Tetra Tech.

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NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

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	5
1.4.1  Underground Production	1
1.5  Recovery Methods	2
1.6  Project Infrastructure	3
1.7  Capital and Operating Costs	4
1.8  Economic Analysis	6
1.9  Expansion Pre-Feasibility Recommendations	8
2.0  INTRODUCTION AND TERMS OF REFERENCE	10
2.1  Report Authors and EBA Quality Control	11
2.2  Site Visits	11
2.3  Effective Date of the Report	11
3.0  RELIANCE OF OTHER EXPERTS	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	108
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	130
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	135
16.6.3  Heap Leach Mining Schedule	135
16.7  Ventilation	135
16.7.1  Ventilation Requirements	135
16.7.2  Ventilation Model	136
16.7.2.1  Conceptual ventilation design	136
16.7.2.2  Intake airway / Escape way Raise	136
16.7.2.3  Fresh Air Raise Head Without separate man way	138
16.7.2.4  Raised Bore for main exhaust ventilation airway	139
16.7.2.5  Vent Tubing Head For Tunnels and Ore drifts	139
16.7.2.6  Setup of ventilation model	140
16.7.3  Interpretation of Ventilation Conceptual Design	145
16.8  Open Pit Mining Method	145
16.9  Spent Ore Mining And Rehandling Methods	146
16.10  Tailings Disposal	146
17.0  RECOVERY METHODS	147
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	164
21.1.2  Underground Development	165
21.1.3  Underground Electrical Installations	166
21.1.4  Construction Consumables and Labour	167
21.1.5  Processing Capital Costs	168
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	177
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
26.5  Breakdown of costs for future work	188
27.0  REFERENCES	189
TABLES
Table 1.1: Update Mineral Reserve and Resource Estimates (April 30, 2013)	2
Table 1.2: Previous Reserve and Resource Estimates (January 2012) for comparative purposes	2
Table 1.3: Santa Elena Reserve Estimation Criteria	4
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

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Table 1.8: Base Case Economic Analysis Results	9
Table 1.9: Comparison between the Base Case and the Spot Price for the Pre-Tax Financial Results	10
Table 1.10: Breakdown of future costs related to recommendations	11
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	777
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

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Table 14.10: Ellipse and Variogram Parameters based on Experimental Variogram Analysis for the Santa Elena Model	86
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	105
Table 15.6: Santa Elena Open Pit Mineral Reserves (Effective December 31, 2012)	106
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

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Table 21.7: Operating Costs	169
Table 21.8: Long Hole Mining Cost Summary	170
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
Table 26.1: Breakdown of future costs related to recommendations	188
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

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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
Figure 13.5: Confirmatory Extraction Kinetics	66
Figure 13.6: Confirmatory Silver Extraction Kinetics	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: IRR Sensitivity Analysis	181

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Figure 24.1: Santa Elena Expansion Project Development Schedule	184
APPENDICES
Appendix A Qualified Person Certificates

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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	Tetra Tech EBA Ltd.
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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1.0   EXECUTIVE SUMMARY

1.1   Introduction

Tetra Tech EBA Inc. (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.  Effective January 1st, 2014, the legal entity EBA Engineering Consultants Ltd. (EBA) became Tetra Tech EBA Inc.

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.

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��

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

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.

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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.

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.

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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.

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.

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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.

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.

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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.

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%
Post-Tax Financial Results
Post-tax NPV DCF @ 5%4	$145,566
Post-tax IRR5	60%

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%
Pre Tax DCF NPV @ 5.0% in millions**	$223.7	$108.7
Pre tax 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.

EBA Engineering has estimated the costs for completion of the above recommended studies as summarised in the table 1.10 below:

 

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Table 1.10: Breakdown of future costs related to recommendations

Recommendation	Future work	Estimated cost
26.1 Geology and Underground Resources	An infill drilling programme consisting of roughly 5,000m of drilling	$1.25 million
26.2 Geotechnical Recommendations	Geotechnical design study for underground	$ 50,000
26.3 Recommendations on Mining Methods and Costs	In house study on options for mining methods	$ 50,000
26.4 Environmental, closure and Asset retirement	Review of existing bond for closure held on the property	$ 25,000
Total future costs related to recommendations		$ 1.375 million

2.0   INTRODUCTION AND TERMS OF REFERENCE

Tetra Tech EBA Inc. (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.  Effective January 1st, 2014, the legal entity EBA Engineering Consultants Ltd. (EBA) became Tetra Tech EBA Ltd.

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).

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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, MBA, 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:

•   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

The contributors to this report have not relied upon other experts for information used to complete this report.

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.

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

Metallurgical test work was 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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13.2.2   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.

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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.

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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

The stope configuration were summarised 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.

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.

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.

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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 (Table 21.2), 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.  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. 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: IRR 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.

26.5   Breakdown of costs for future work

EBA Engineering has estimated the costs for completion of the above recommended studies as summarised in the table 26.1 below:

Table 26.1: Breakdown of future costs related to recommendations

Recommendation	Future work	Estimated cost
26.1 Geology and Underground Resources	An infill drilling programme consisting of roughly 5,000m of drilling	$1.25 million
26.2 Geotechnical Recommendations	Geotechnical design study for underground	$ 50,000
26.3 Recommendations on Mining Methods and Costs	In house study on options for mining methods	$ 50,000
26.4 Environmental, closure and Asset retirement	Review of existing bond for closure held on the property	$ 25,000
Total future costs related to recommendations		$ 1.375 million

Note that since Santa Elena is an operating mine, much of the above work is ongoing or is being undertaken in house with current in house equipment and expertise.

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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.

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.

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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.

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

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

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.

Effective Date: April 30, 2013

Amended Date: 04 March, 2014

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, 26.4 and 26.5.

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.

Dated this 04 March, 2014,

Original signed and sealed by

“N. Eric Fier”

________________________________

N. Eric Fier, CPG, P.Eng., President and Chief Operating Officer for SilverCrest Mines Inc.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

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 Tetra Tech EBA Inc., 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

Effective Date: April 30th, 2013

Amended Date: 04 March. 2014

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, 26.1 and 27 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.

Dated this 04 of March, 2014

Original signed and sealed by

“P. James F. Barr”

________________________________

P. James F. Barr, P.Geo

Senior Geologist, Tetra Tech EBA Inc.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

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 Tetra Tech EBA Inc. 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.

Effective Date: April 30, 2013

Amended Date: 04 March, 2014

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.

Dated this 04 March, 2014,

Original signed and sealed by

“C. Michael Tansey”

________________________________

C. Michael Tansey, P.Eng.

Senior Mining Engineer, Tetra Tech EBA Inc.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

Certificate of qualified person – John Fox

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.C , V6B 1E8

2)   I hold a Bachelors of Science in Applied Mineral Sciences and Extractive Metallurgy from the University of Leeds, 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.

Effective Date: April 30, 2013

Amended Date: 04 March, 2014

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.

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.

Dated this 04 March, 2014,

Original signed and sealed by

“John R.W. Fox”

________________________________

John R.W. Fox, P.Eng.

Consulting Metallurgical Engineer

Laurion Consulting Inc.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

I, Carlos Chaparro, do hereby declare that:

1)   I currently reside in Vancouver, British Columbia, Canada, and am currently employed as Senior Geotechnical Engineer for Tetra Tech EBA Inc., 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.

Effective Date: April 30, 2013

Amended Date: March 04, 2014

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 4th, 5th and 6th of September,  2011 and 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 which is the subject of the technical report.

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.

Dated this 04 day of March, 2014

Original signed and sealed by

“Carlos Chaparro”

 Carlos Chaparro, P.Eng

Senior Geotechnical Engineer, Tetra Tech EBA Inc.

NATIONAL INSTRUMENT 43-101 TECHNICAL REPORT

 EFFECTIVE DATE: APRIL30, 2013 | AMENDED DATE: MARCH 04, 2014

I, Nick Michael, BS, MBA, do hereby declare that:

1)   I am a Principal Mineral Economist with Tetra Tech, Inc. with a business address at 350 Indiana Street, Suite 500, Golden, Colorado 80401, USA.

2)   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.

3)   I am a Registered Member in good standing (#4104304) with the Society for Mining, Metallurgy and Exploration, Inc (SME).

4)   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.

5)   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.

Effective Date: April 30, 2013

Amended Date: March 04, 2014

6)   I am responsible for section 22 and 25.3 of this Technical Report.

7)   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.

8)   I have never visited nor have I had involvement with the Santa Elena property prior to this report.

9)   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.

10)   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.

11)   Prior to this report, I have had no prior involvement on this property.

12)   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.

Dated this 04 March, 2014,

Original signed and sealed by

“Nick Michael”

________________________________

Nick Michael, BS, MBA

Mineral Economist, Tetra Tech Inc.