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Technical ReportTechnical Report,Cozamin Mine

Zacatecas, Mexico

Prepared for:

Capstone Mining Corp.Suite 900, 999 West Hastings Street

Vancouver, BC CANADA V6C 2W2

Prepared by:

Project No. 2CC031.000

Effective Date: March 31, 2009

Technical Report Cozamin Mine

Zacatecas, Mexico

Capstone Mining Corp. Suite 900 – 999 West Hastings Street

Vancouver, BC V6C 2W2

SRK Consulting (Canada) Inc. Suite 2200, 1066 West Hastings Street

Vancouver, B.C. V6E 3X2

Tel: 604.681.4196 Fax: 604.687.5532 E-mail: [email protected] Web site: www.srk.com

SRK Project Number 2CC031.000

Effective Date: March 31, 2009

This report was written by the following Qualified Persons:

Gordon Doerksen, P.Eng. Jenna Hardy, P.Geo.

Robert Sim, P.Geo. Jeff Woods, CP

Cover PhotosTop: La Guadalupana Ramp Portal Middle: Lead Flotation Cell Bottom: Crushing Circuit

SRK Consulting Technical Report - Cozamin Mine, Mexico Page i

GED/HA Capstone_Cozamin_Technical_Report_2CC031 000_GD_20090527.doc, May. 27, 09 March 31, 2009

Table of Contents

1 Executive Summary ..................................................................................................... 1

2 Introduction ................................................................................................................ 11

3 Reliance on Other Experts ........................................................................................ 12

4 Property Description and Location .......................................................................... 13

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ........ 17

6 History ......................................................................................................................... 19

7 Geological Setting ...................................................................................................... 21

8 Deposit Types ............................................................................................................ 27

9 Mineralization ............................................................................................................. 28

10 Exploration ................................................................................................................. 30

11 Drilling ......................................................................................................................... 34

12 Sampling Method and Approach .............................................................................. 5012.1 Diamond Drill Core Sampling ............................................................................................ 50

12.1.1 Drill Site Control .................................................................................................................... 5012.1.2 Core Shack Control ............................................................................................................... 5012.1.3 Survey control ....................................................................................................................... 51

12.2 Underground Chip Sampling ............................................................................................. 51

13 Sample Preparation, Analyses and Security ........................................................... 5313.1 Sampling Personnel .......................................................................................................... 5313.2 Drill Core Sample Preparation and Analytical Procedures ................................................ 5313.3 Underground Channel Sample Preparation and Analytical Procedures ........................... 56

14 Data Verification......................................................................................................... 5714.1 Database Validation .......................................................................................................... 5714.2 Site Visit Validation ........................................................................................................... 5814.3 Conclusions ....................................................................................................................... 58

15 Adjacent Properties ................................................................................................... 59

16 Mineral Processing and Metallurgical Testing ........................................................ 6016.1 Metallurgical Testing ......................................................................................................... 60

16.1.1 Process Mineralogy ............................................................................................................... 6016.1.2 Metallurgical Testing ............................................................................................................. 61

16.2 Process Description and Flow Sheets .............................................................................. 6116.2.1 Crushing and Screening ........................................................................................................ 61

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16.2.2 Milling .................................................................................................................................... 6316.2.3 Flotation ................................................................................................................................. 6516.2.4 Concentrate Dewatering and Filtration .................................................................................. 6716.2.5 Tailings Handling ................................................................................................................... 69

16.3 Production ......................................................................................................................... 7116.3.1 Production Data Collection and Compilation ......................................................................... 7116.3.2 Production Data ..................................................................................................................... 71

16.4 Conclusions ....................................................................................................................... 75

17 Mineral Resource and Mineral Reserve Estimates .................................................. 7617.1 Mineral Resource Estimate ............................................................................................... 76

17.1.1 Introduction ............................................................................................................................ 76

17.2 Geologic Model, Domains and Coding .............................................................................. 7617.3 Available Data ................................................................................................................... 7717.4 Compositing ...................................................................................................................... 8017.5 Exploratory Data Analysis ................................................................................................. 80

17.5.1 Basic Statistics by Domain .................................................................................................... 8117.5.2 Contact Profiles ..................................................................................................................... 8117.5.3 Conclusions and Modeling Implications ................................................................................ 81

17.6 Bulk Density Data .............................................................................................................. 8217.7 Evaluation of Outlier Grades ............................................................................................. 8317.8 Variography ....................................................................................................................... 8317.9 Model Setup and Limits ..................................................................................................... 8517.10 Interpolation Parameters ........................................................................................ 8517.11 Validation ................................................................................................................ 86

17.11.1 Visual Inspection ............................................................................................................... 8617.11.2 Model Checks for Change of Support ............................................................................... 8717.11.3 Comparison of Interpolation Methods ............................................................................... 8817.11.4 Swath Plots (Drift Analysis) ............................................................................................... 90

17.12 Resource Classification .......................................................................................... 9117.13 Mineral Resource Estimate .................................................................................... 9517.14 Mineral Reserve Estimate ...................................................................................... 97

17.14.1 Assumptions ...................................................................................................................... 9817.14.2 Reserve Estimate ............................................................................................................ 10417.14.3 Risks to the Mineral Reserve Estimate ........................................................................... 10617.14.4 Mineral Reserve Classification ........................................................................................ 107

18 Other Relevant Data and Information..................................................................... 10818.1 Historical Production Data ............................................................................................... 10818.2 Reconciliation of Production Data ................................................................................... 109

19 Additional Requirements for Technical Reports on Development Properties and Production Properties ............................................................................................. 11019.1 Mining Operations ........................................................................................................... 110

19.1.1 Deposit context .................................................................................................................... 11019.1.2 Mining Methods ................................................................................................................... 11019.1.3 Mine Development .............................................................................................................. 11119.1.4 Mine Access ........................................................................................................................ 11119.1.5 Material Handling ................................................................................................................ 11119.1.6 Mine Ventilation ................................................................................................................... 11219.1.7 Mobile Equipment ................................................................................................................ 11219.1.8 Life of Mine Development and Production Schedule .......................................................... 112

19.2 Recoverability .................................................................................................................. 11319.3 Markets ........................................................................................................................... 113

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19.4 Contracts ......................................................................................................................... 11319.5 Environmental Considerations ........................................................................................ 114

19.5.1 Environmental Assessment and Permitting ........................................................................ 11419.5.2 Closure Plan ........................................................................................................................ 118

19.6 Taxes 11919.7 Capital Cost Estimates .................................................................................................... 12019.8 Operating Cost Estimates ............................................................................................... 12019.9 Economic Analysis .......................................................................................................... 121

19.9.1 Economic Results ................................................................................................................ 12219.9.2 Sensitivity Analysis .............................................................................................................. 124

19.10 Payback ................................................................................................................ 12519.11 Mine Life ............................................................................................................... 125

20 Interpretation and Conclusions .............................................................................. 12620.1 Risks 12620.2 Opportunities ................................................................................................................... 127

21 Recommendations ................................................................................................... 128

22 Acronyms and Abbreviations ................................................................................. 130

23 Illustrations ............................................................................................................... 131

24 References ................................................................................................................ 132

25 Date and Signature Page ......................................................................................... 133

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List of Tables

Table 1.1: Annual Tonnes and Grade Processed ............................................................................. 3Table 1.2: Annual Concentrate Grades and Process Recoveries ..................................................... 3Table 1.3: San Roberto Mineral Resource Estimate Summary ......................................................... 4Table 1.4: San Rafael Mineral Resource Estimate Summary ........................................................... 5Table 1.5: Mineral Reserve Estimate by Classification (Dec 31. 2008) ............................................ 5Table 1.6: Annual LOM Plan Production ........................................................................................... 6Table 1.7: Economic Analysis Assumptions ...................................................................................... 7Table 1.8: LOM Economic Analysis Summary .................................................................................. 7Table 1.9: Sensitivity Analysis Results .............................................................................................. 8Table 2.1: Qualified Person and Author Site Visit Dates ................................................................. 11Table 4.1: Cozamin Claim Status March 2009 ................................................................................ 14Table 11.1: Drilling History by Contractor ........................................................................................ 34Table 11.2: Phase I and II Surface Drilling Results ......................................................................... 35Table 11.3: Phase III Underground Drilling Results ........................................................................ 36Table 11.4: Results from the Phase IV Surface Drilling Program ................................................... 40Table 11.5: Results from the Phase V Underground Drilling Program ............................................ 41Table 11.6: Results from the Phase VI Surface Drilling Program ................................................... 43Table 11.7: Results from the Phase VI Underground Drilling Program ........................................... 47Table 13.1: Primary and check laboratories used for Cozamin drill samples .................................. 53Table 13.2: Laboratory methods and elements routinely analyzed ................................................. 54Table 13.3: Analytical methods used when re-analyzing samples with over limit results ............... 54Table 13.4: Primary and check laboratories used for Cozamin channel samples ........................... 56Table 16.1: Actual Monthly Concentrate Grade and Flotation Recovery ........................................ 72Table 16.2: Reagent Consumption .................................................................................................. 74Table 17.1: Statistical Summary of Sample Assay Data ................................................................. 80Table 17.2: Statistical Summary of Composited Sample Data inside Minzone Domain ................. 81Table 17.3: Outlier Grade Controls ................................................................................................. 83Table 17.4: Variogram Parameters ................................................................................................. 84Table 17.5: Block Model Limits ....................................................................................................... 85Table 17.6: Interpolation Parameters for San Roberto Area ........................................................... 86Table 17.7: Interpolation Parameters for San Rafael Area ............................................................. 86Table 17.8: San Roberto Mineral Resource Estimate Summary ..................................................... 96Table 17.9: San Rafael Mineral Resource Estimate Summary ....................................................... 97Table 17.10: Mineral Reserve Estimate NSR Parameters .............................................................. 99Table 17.11: Mineral Reserve Estimates by Stope Block (Dec. 31, 2008) .................................... 105Table 17.12: Mineral Reserve Estimate (Dec. 31, 2008) .............................................................. 107Table 18.1: Historical Plant Production Data ................................................................................. 108Table 18.2 : Mine vs. Mill Reconciliation ....................................................................................... 109Table 18.3: Resource Model vs. Mill Reconciliation ...................................................................... 109Table 19.1: Major Underground Mobile Equipment (Cozamin and contractors) ........................... 112Table 19.2: LOM Operating Schedule ........................................................................................... 112Table 19.3: LOM Flotation Recovery Estimate .............................................................................. 113Table 19.4: Metal and Concentrate Purchase Contracts .............................................................. 113Table 19.5: Capital Cost Summary Estimate ................................................................................ 120Table 19.6: Operating Cost Summary Estimate ............................................................................ 121Table 19.7 LOM Plan with Annual Operating Costs ...................................................................... 121Table 19.8: LOM Main Economic Analysis Assumptions (all years) ............................................. 122Table 19.9: LOM Economic Analysis Summary (2009 – 2017) ..................................................... 122

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Table 19.10: Pre-tax NPV by Discount Rate ................................................................................. 123Table 19.11 Sensitivity Analysis ................................................................................................... 124

List of Figures

Figure 4.1: Cozamin Mine Location in Central Mexico ...................................................................... 13Figure 4.2: Cozamin Concession Area ............................................................................................ 15Figure 4.3: Concession Map of the Cozamin Project Area ................................................................ 16Figure 5.1: Surface Layout of the Cozamin Mine Facilities ............................................................. 18Figure 5.2: Image of Cozamin looking south with Zacatecas in the background (from Google Earth)18Figure 7.1: Regional Geology around the Cozamin Property .......................................................... 22Figure 7.2: Plan showing the distribution of mineralized veins near Zacatecas .............................. 24Figure 7.3: Cross Faults, Level 8 Cozamin Mine ............................................................................ 25Figure 10.1: Longitudinal Section showing pierce points of drilling at Mala Noche Vein ................ 33Figure 11.1: Longitudinal Section view showing the area of the San Roberto Resource ................ 49Figure 16.1 Cozamin Crushing and Screening Flow Sheet .......................................................... 62Figure 16.2: Cozamin Milling Process Flow Sheet .......................................................................... 64Figure 16.3: Cozamin Flotation Process Flow Sheet ...................................................................... 66Figure 16.4: Cozamin Concentrate Handling Process Flow Sheet ................................................. 68Figure 16.5: Cozamin Tailings Handling Process Flow Sheet ....................................................... 70Figure 16.6: Concentrator Monthly Tonnage ................................................................................... 73Figure 16.7: Concentrator Monthly Metal Recoveries ..................................................................... 73Figure 16.8: Concentrator Monthly Concentrate Grades ................................................................ 74Figure 17.1: Drill Hole Distribution ................................................................................................... 78Figure 17.2: Grade and Tonnage Curves of Herco Copper at San Roberto ................................... 87Figure 17.3 Grade and Tonnage Curves of Herco Zinc at San Rafael .......................................... 88Figure 17.4: GT Comparison OK, ID and NN Copper at San Roberto ............................................ 89Figure 17.5: GT Comparison OK, ID and NN Zinc at San Raphael ................................................ 89Figure 17.6: Copper Swath Plot ...................................................................................................... 90Figure 17.7: Zinc Swath Plot ........................................................................................................... 91Figure 17.8: Long Section Showing Resource Classifications ........................................................ 94Figure 17.9: Mineral Reserve Estimation Blocks and Shapes ...................................................... 101Figure 17.10: Illustrative Steps Taken to Define Cut and Fill Mining Shapes ............................... 102Figure 17.11: Steps to Create Long Hole Mining Shape ............................................................... 103Figure 19.1: Economic Analysis Graph ......................................................................................... 123Figure 19.2: NSR Value by Metal .................................................................................................. 124Figure 19.3: Sensitivity Analysis Graph ......................................................................................... 125

List of Appendices

APPENDIX A: Simplified Economic Model

SRK Consulting Technical Report - Cozamin Mine, Mexico Page 1


1 Executive Summary

Introduction

This Technical Report was compiled by SRK Consulting (Canada) Inc. for Capstone Mining Corp.

to provide updated mineral resource and reserve estimates and update the latest operational

conditions and summarize the current life of mine plan. This report was written by Robert Sim,

P.Geo., Jenna Hardy, P.Geo., Jeff Woods, CP and Gordon Doerksen, P. Eng., all Qualified Persons

as defined by NI 43-101.

Location and Ownership

The Cozamin Cu-Zn-Pb-Ag mine and processing plant are located two kilometres northwest of the

city of Zacatecas, Mexico at approximately 22º 48’ N latitude and 102 º 35’ W longitudes. The mine

site is accessible via a short all-weather gravel road from Zacatecas. Infrastructure in the region is

well established and the mine is connected to the regional electrical power grid.

Capstone Mining Corp. owns the Cozamin operation and operates it through its wholly owned

subsidiary, Capstone Gold SA de CV. The Cozamin property is made up of 33 mining concessions

covering 2,898 hectares of area.

Zacatecas has been an active mining region since the 16th century and the near-surface portions of

the mine have been exploited by past operators.

Geology and Exploration

The Zacatecas Mining District covers a belt of epithermal and mesothermal vein deposits that

contain silver, gold and base metals.

Since 2004, Capstone has undertaken exploration and definition drilling totalling 366 diamond drill

holes and 105,261 m. The dominant mineralized vein on the Cozamin property is called the

Mala Noche. This vein has been traced for 5.5 km, strikes approximately east-west and dips on

average at 60º to the north. The Mala Noche vein system occupies a system of anastomosing faults

that are principally comprised of the Mala Noche and El Abra faults along with other less significant

faults. The mineralized bodies within the Mala Noche appear to be strongest where the disparate

faults coalesce into a single fault zone. Although not all of the fault system is mineralized at any

given location, there have been no other significant mineralized fault zones discovered to date.

Results from the exploration and mine development to date indicate that some of the strongest

mineralization in the San Roberto mine rakes to the west at approximately -50º within the vein. Post

mineralization offsets of the Mala Noche vein are minimal and occur along high angle, normal faults

that strike northeast.



The Mala Noche vein in the San Roberto mine workings shows contained sulphides to occur as

disseminations, bands and masses. Pyrite is the dominant vein sulphide and typically comprises

approximately 15 % of the Mala Noche vein in the San Roberto mine.

Pyrrhotite is the second most common sulphide mineral but is present only in the intermediate and

deeper levels of the San Roberto mine and commonly occurs as an envelope to, or intermixed with,

strong chalcopyrite mineralization. Chalcopyrite is the dominant copper sulphide at Cozamin. Like

pyrrhotite, it is more common at the intermediate and deeper levels of the mine and occurs as

disseminations, veinlets and replacement masses. These masses appear to be fractured and

brecciated at intermediate levels in the mine. Minor bornite occurs as disseminated grains in some of

the higher grade zones.

Sphalerite is the most common zinc sulphide mineral and occurs as disseminations and coarse

crystalline masses. Galena is less common than sphalerite but is generally associated with it.

Argentite is the most common silver mineral. It has been identified microscopically occurring as

inclusions in chalcopyrite, pyrite and likely sphalerite and galena.

The main gangue minerals are quartz, chlorite and calcite.

The distribution of metal value in the Cozamin reserves are found predominantly in copper (84 %),

followed by zinc (7 %) and silver (7 %) with minor contribution from lead (2 %). Note that the

distribution of metal values is based on sales at the reserve metal prices including the Silverstone

agreement.

Operating Results

The Cozamin Mine commenced operation in June 2006 and since that time has maintained

continuous production and shown continual improvement. Tables 1.1 and 1.2 show annual

summaries of mine and mill performance.

Since the start of operations, the mill has undergone numerous upgrades, expansions and operating

optimizations. The mine has seen improved access, ventilation and an increase in its mobile

equipment fleet. A shift away from cut and fill mining to predominantly Long Hole (“LH’) open

stoping methods has enabled higher mine production rates.

The life of mine (“LOM”) plan production rate is 1,015,000 tonnes/year and is supported by the

operating results in latter half of 2008 and the first quarter of 2009.



Table 1.1: Annual Tonnes and Grade Processed

Period Ktonnes Ag

(g/t) Cu(%)

Pb(%)

Zn(%)

2006 (Jun.-Dec.) 185.5 70 1.42 0.65 1.77

2007 597.6 70 1.69 0.57 1.37

2008 833.2 63 1.62 0.55 1.31

2009 (1st Quarter) 248.3 56 1.96 0.33 0.81

Total / Average 1,864.6 65 1.67 0.54 1.31

Table 1.2: Annual Concentrate Grades and Process Recoveries

Month Concentrate Grade (%) Recovery (%)

Cu Pb Zn Ag Cu Pb Zn

2006 (Jun.-Dec.) 25 60 46 70 90 71 56

2007 22 63 39 69 86 50 44

2008 23 63 41 71 88 65 49

2009 (1st Quarter) 24 67 45 71 91 63 54

Average 23 63 42 70 88 60 49

Mineral Resource and Reserve Estimates

The mineral resource model has been developed using the MineSight® (v4.50) with a nominal block

size measuring 10 m x 3 m x 3 m, with the long axis oriented parallel to the E-W strike of the

deposit. Grade estimates are made using ordinary kriging with parameters derived from the

geostatistical properties present in the underlying database. Bulk densities are estimated into model

blocks using the inverse distance (ID2) interpolation method. Resources are classified in accordance

with the CIM definition standards for mineral resources.

The mineral resource estimates are shown in Tables 1.3 and 1.4 respectively for the San Roberto and

San Raphael deposits. The mineral reserve estimate for San Roberto is shown in Table 1.5. No

mineral reserves were defined for the San Raphael deposit, as there is insufficient geotechnical and

metallurgical testing, and analysis to support conversion to reserves.



Table 1.3: San Roberto Mineral Resource Estimate Summary

Cut-off Grade (Cu %)

Ktonnes Cu(%)

Zn(%)

Pb(%)

Ag(g/t)

Au (g/t)

SG(t/m

3)

Measured

0.50 2,287 2.00 1.10 0.43 79.2 0.068 2.93

1.00 1,908 2.24 1.03 0.42 84.1 0.064 2.93

1.15 1,749 2.35 1.01 0.41 86.0 0.063 2.93

1.50 1,373 2.63 0.96 0.37 90.0 0.061 2.94

2.00 947 3.04 0.93 0.31 95.1 0.057 2.94

2.50 616 3.48 0.94 0.24 99.0 0.056 2.94

3.00 386 3.92 0.93 0.18 101.3 0.056 2.95

Indicated

0.50 12,303 1.35 1.26 0.30 54.8 0.074 2.91

1.00 7,296 1.76 1.21 0.27 61.2 0.065 2.92

1.15 6,077 1.90 1.20 0.25 63.0 0.063 2.92

1.50 3,963 2.22 1.16 0.23 66.9 0.060 2.93

2.00 2,091 2.67 1.12 0.21 72.8 0.056 2.96

2.50 1,030 3.13 1.04 0.21 79.5 0.052 3.00

3.00 490 3.59 0.94 0.21 87.1 0.050 3.04

Measured + Indicated

0.50 14,590 1.45 1.23 0.32 58.6 0.073 2.91

1.00 9,204 1.86 1.17 0.30 65.9 0.065 2.92

1.15 7,826 2.00 1.15 0.29 68.2 0.063 2.93

1.50 5,336 2.33 1.11 0.26 72.8 0.060 2.93

2.00 3,038 2.79 1.06 0.24 79.7 0.056 2.95

2.50 1,646 3.26 1.00 0.22 86.8 0.054 2.98

3.00 876 3.74 0.94 0.20 93.4 0.053 3.00

Inferred

0.50 4,782 0.95 1.06 0.21 42.4 0.073 2.81

1.00 1,623 1.42 0.98 0.19 49.0 0.063 2.84

1.15 1,100 1.58 0.95 0.17 52.5 0.065 2.85

1.50 504 1.93 1.03 0.18 62.4 0.071 2.92

2.00 181 2.32 1.04 0.16 72.3 0.075 2.98

2.50 41 2.75 0.88 0.15 73.4 0.076 3.13

3.00 5 3.13 0.62 0.10 77.4 0.089 3.26

(1) Mineral Resources do not have demonstrated economic viability.

(2) The “base case” cut-off grade of 1.15 %Cu is highlighted in table.



Table 1.4: San Rafael Mineral Resource Estimate Summary

Cut-off Grade (Zn %)

Ktonnes Cu(%)

Zn(%)

Pb(%)

Ag(g/t)

Au (g/t)

SG(t/m

3)

Indicated

2.0 3,431 2.97 0.21 0.40 33.8 0.441 2.75

2.5 2,407 3.29 0.22 0.43 36.0 0.469 2.75

3.0 1,467 3.64 0.23 0.47 38.3 0.482 2.76

3.5 720 4.07 0.25 0.50 41.4 0.489 2.78

4.0 328 4.48 0.24 0.52 44.3 0.462 2.80

4.5 135 4.87 0.24 0.56 47.5 0.468 2.81

5.0 41 5.22 0.25 0.61 51.3 0.518 2.82

Inferred

2.0 2,642 2.61 0.09 0.37 24.0 0.436 2.65

2.5 1,161 3.11 0.12 0.47 30.2 0.514 2.68

3.0 556 3.55 0.14 0.57 35.8 0.609 2.68

3.5 256 3.92 0.14 0.65 39.5 0.675 2.67

4.0 83 4.32 0.15 0.72 41.9 0.714 2.66

4.5 19 4.76 0.17 0.81 45.4 0.709 2.65

5.0 3 5.16 0.20 0.73 51.4 0.855 2.66


(2) The “base case” cut-off grade of 3 %Zn is highlighted in table.

Table 1.5: Mineral Reserve Estimate by Classification (Dec 31. 2008)

Classification KTonnes Ag

(g/t) Cu(%)

Pb(%)

Zn(%)

Proven 1,606 76.91 2.02 0.44 0.97

Probable 6,491 55.38 1.57 0.26 1.13

Total 8,097 59.65 1.66 0.29 1.10

The mineral reserve estimate utilized metal prices of $1.50 per pound copper, $0.50 per pound zinc,$0.45 per pound lead, and $4.00 per ounce silver.

LOM Operating Plan

The LOM operating plan was reviewed by SRK and deemed to be appropriate. The ore is planned to

be extracted using three mining methods; cut and fill using waste rock fill, longhole open stoping and

Avoca. Each method has been assigned to different mining blocks depending on the physical

characteristics of the orebody.

Development mining and equipment usage was estimated based on the mine schedule. Capital

development is conducted using a Mexican-based contractor. All other mining at Cozamin is done

using Capstone employees.

The mine extends for a strike length of over 1 km and reserves extend to a depth of 600 m. Access

to the underground workings is obtained from two service and haulage ramps and a hoisting shaft.

An annual summary of the tonnes, grade, payable metal and cash costs is shown in Table 1.6.



Table 1.6: Annual LOM Plan Production

Parameter 2009 2010 2011 2012 2013 2014 2015 2016 2017 Total

Mining

Development (m) 7,835 6,077 5,058 3,835 5,863 7,450 6,210 4,302 660 47,290

Milling

Tonnes (000s) 1,015 1,015 1,015 1,015 1,015 1,015 1,015 734 258 8,097

Copper grade (%) 1.70 1.87 1.86 1.88 1.51 1.51 1.52 1.47 1.28 1.66

Zinc grade (%) 1.22 1.12 1.07 0.98 1.04 1.02 1.11 1.18 1.34 1.10

Lead grade (%) 0.51 0.45 0.38 0.34 0.19 0.15 0.17 0.16 0.16 0.29

Silver grade (g/t) 72 72 70 68 53 51 48 43 38 60

Payable Metals

Copper (Mlbs) 33.1 36.6 36.3 36.8 29.4 29.5 29.8 20.8 6.4 258.6

Zinc (Mlbs) 14.8 13.6 12.9 11.9 12.6 12.4 13.5 10.3 4.1 106.1

Lead (Mlbs) 6.5 5.7 4.9 4.3 2.4 2.0 2.1 1.4 0.5 29.9

Silver (Moz) 1.6 1.6 1.6 1.5 1.2 1.1 1.0 0.7 0.2 10.4

Cash costs (US$/lb payable Cu)

Production (on site) costs

1.09 0.99 0.99 0.98 1.23 1.22 1.21 1.25 1.44 1.11

By-product Credits for Zn, Pb & Ag

0.41 0.35 0.33 0.30 0.32 0.31 0.31 0.32 0.38 0.33

Off site cost of Cu concentrate

0.32 0.32 0.32 0.32 0.32 0.32 0.32 0.32 0.32 0.32

Total Cash Cost 1.00 0.96 0.99 1.00 1.22 1.24 1.22 1.26 1.38 1.10

Economic Analysis

A pre-tax economic model of the operating plan was generated using the assumptions shown in

Table 1.7. The estimate of the pre-tax operating results is shown in Table 1.8. Only ongoing capital

(2009 and beyond) was taken into account in the model and, as a result, the project has a very

favourable pre-tax net present value (NPV) at an 8 % discount rate of $172 M.



Table 1.7: Economic Analysis Assumptions

Item Unit Value

METAL PRICES

Copper $/lb 2.00

Zinc $/lb 0.70

Lead $/lb 0.60

Silver $/oz 4.00

FLOTATION RECOVERY

Copper in Cu concentrate % 91

Zinc in Zn concentrate % 65

Lead in Pb concentrate % 60

Silver in all concentrates % 74

OFF-SITE COSTS

TC/RC, Transport, Payables, Penalties, Price Participation

$/t As per current contracts

OPEX

Unit mining cost $/t milled 18.03

Unit processing cost $/t milled 12.99

Unit G&A cost $/t milled 4.49

Unit cost total M$ 35.51

CAPEX

LOM Capital M$ 17.5

Table 1.8: LOM Economic Analysis Summary

Item Unit Value

LOM PRODUCTION

Ore Mined Mt 8.1

Mill head grade – copper Cu % 1.66

Mill head grade – zinc Zn % 1.10

Mill head grade – lead Pb % 0.29

Mill head grade – silver Ag g/t 60

METAL PRODUCTION

Copper in Cu concentrate t Cu 122,000

Zinc in Zn concentrate t Zn 58,000

Lead in Pb concentrate t Pb 14,000

Silver in all concentrates oz Ag 11,468,000

REVENUE

Total NSR revenue (before royalty) M$ 537

Royalty (@3 %) M$ 16


COST

Total OPEX M$ 288

Capex (inc. sustaining) M$ 18

ECONOMIC RESULTS (EBITDA)

NPV0 %DR M$ 216

NPV8 %DR M$ 172



A sensitivity analysis was performed individually on metal price, metal grade, capital cost and

operating cost. The project is sensitive equally to metal price and grade fluctuations with an $82 M

(48 %) increase in pre-tax NPV8% as a result of a 20 % increase in metal price or grade. Conversely,

the project NPV drops by $82 M for a 20 % decrease in metal price or grade. As most of the project

capital has already been spent, the project is not sensitive to capital.

The project is somewhat sensitive to operating costs. A 20 % increase in operating costs leads to a

26 % ($44 M) drop in pre-tax NPV8%. See Table 1.9 for sensitivity results.

Table 1.9: Sensitivity Analysis Results

VariablePre-tax NPV8 % (M$)

-20 % 0 % +20 %

Capital Cost 175 172 169

Operating Cost 217 172 128

Metal Price 90 172 254

Grade 90 172 254

Conclusions and Recommendations

The Cozamin project has been successfully developed into viable mining and milling operation that,

based on the assumptions made, shows a positive return on the mining and processing of current

reserves and has the potential to expand its life if some of the current resources can be converted into

reserves. There is no guarantee that an increase in reserves will be achieved and will depend upon

further exploration, metallurgical, geotechnical and hydrogeology assessments as well as market

conditions such as metal prices and smelter terms.

The main risks to the project are:

Water Supply and Management – Long-term water supply for the 3,000 tpd production rate is

not totally established and mine personnel are looking at solutions which include greater control

of the site water balance, securing underground water rights, drilling a deep ground water well

and improvements to fresh water diversion structures.

Mining Control - The mine must continue to ensure accurate drilling and blasting practices are

maintained to minimize dilution, minimize secondary breaking and optimize extraction.

Adequate back-up stopes must be available to give the mine production flexibility should

dilution become a problem in a particular stope.

External Factors – Exchange rates, off-site costs and, in particular, base metal prices all have the

potential to seriously affect the economic results of the mine. Negative variance to these items

from the assumptions made in the economic model would reduce the profitability of the mine

and the mineral resource and reserve estimates.



The main opportunities for the project are:

Improved ore handling system for the hoisting shaft – The LH mining method has the potential

to produce oversized muck that can be a bottleneck at the shaft grizzly and loading pocket.

Improvements in drilling and blasting practices can help alleviate the problem as well as an

improvement in the underground truck dump/grizzly/rock breaker set-up.

Timely updates of the resource model will allow for better mine planning and scheduling.

Previous planning has been conducted using primarily the channel sample results – essentially a

2 dimensional approach to defining mining limits. Monthly updates of a block model using all

available sampling and mapping information will greatly improve the mine design/planning

process.

Continued improvement in metal recovery and concentrate grade as demonstrated in year to date

operating statistics.

Maximizing mill throughput on a sustained basis to reduce unit costs. Mill has operated in excess

of 3,500 tpd for periods of days which is 20 % greater than LOM throughput.

Mine life may be extended by exploration on the 4 km of the Male Noche vein outside of the

existing resource area, or converting the San Roberto inferred resources to reserves or acquiring

additional claims which cover the down dip extension of the Male Noche vein to the east .

Review of 31 drill holes omitted from resource model due to apparent survey issues may result

in an increase in resources. For example, drill hole U62 (16.2 m 3.3 % Cu) intersected

significantly thicker mineralization than surrounding channel sample data. If, through

resurveying or re-drilling (if required), this hole is reintroduced into the database, it may result in

an increase in the overall mineral resource.

The main recommendations identified by the QP authors of this report are summarized as:

Refine the water balance to determine needs and potential long-term sources;

Improve the characterization of ARD/ML of tailings and waste rock with further sampling, and

testing to support storage options decisions;

Mine ventilation measurement and control needs to be improved so ventilation system is

optimized in terms of overall air volume and the ventilation of each individual mining area;

Surveying of the mined-out LH stopes is highly advisable to help drilling and blasting practices

to minimize dilution and optimize extraction as well for reconciliation of planned vs. actual

stope shapes.

Review of a series of 31 drill holes which exhibit irregular results compared to surrounding data

suggesting errors in the recording of these holes in the database. Re-drilling of some holes may

be required after this review, but other holes may be added to the database if specific errors are



found and remedied. Work practices must be altered so that data collected from all future drilling

programs is properly audited on a timely basis.

Monthly updates of the block model using all current sampling and mapping information in

order to provide a timely and accurate basis for mine planning purposes.



additional claims which cover the down dip extension of the Male Noche vein to the east.

A review of 31 drill holes omitted from resource model due to apparent survey issues may result

in an increase in resources. For example, drill hole U62 (16.2 m 3.3 % Cu) intersected


resurveying or re-drilling (if required), this hole is reintroduced into the database, it may result in


All of the recommendations above are part of the on-going operation of the mine and do not require

a special budget for their implementation.



2 Introduction

This technical report was compiled by SRK Consulting (Canada) Inc. for Capstone Mining Corp.

“Capstone” to describe new mineral resource and reserve estimates for the Cozamin Mine, Mexico.

The report also provides an update of the Cozamin operations and life of mine (“LOM”) plan.

Contributions were made to this report by several Qualified Persons as noted in Table 2.1. Each QP

is solely responsible for his/her own work.

Table 2.1: Qualified Person and Author Site Visit Dates

Name Role Scope of Inspection Date of Most Recent

Site Visit

Gordon Doerksen, P.Eng. QP Underground mine, concession boundaries and general site

February 10-13, 2009

Jenna Hardy, P.Geo. QP General site February 5-13, 2009

Robert Sim, P.Geo. QP Underground mine, Sample prep. facility March 24 – 26, 2009

Jeff Woods, CP QP Processing facility and assay lab February 10-13, 2009

The sources of data and information used in this report are cited in Section 24 - References.

All units in this report are based on the Système International d'Unités (International System of Units

or “SI”), except for some units which are deemed industry standards such as troy ounces (oz) for

precious metals and pounds (“lb”) for base metals.

All currency values are in US dollars (“$”) unless otherwise noted.

This report uses many abbreviations and acronyms common in the mining industry, most of which

are defined in the body of the text. Further explanations can be found in Section 22.

SRK’s opinion contained herein is based on information provided to SRK by Capstone and its

consultants throughout the course of SRK’s investigations, which in turn reflect various technical

and economic conditions at the time of writing. The information so provided has been taken in good

faith by SRK and while SRK has checked and/or verified the information wherever possible, SRK

cannot be held responsible for errors or omissions in the source data. Given the nature of the mining

business, these conditions can change significantly over relatively short periods of time and

consequently the outcomes can be more or less favourable than projected.

This report includes technical information, which requires subsequent calculations to derive

subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding

and consequently introduce a margin of error. Where these occur, SRK does not consider them to be

material.



3 Reliance on Other Experts

This report relies on non-QP, Mr. José de Jesús Espino Zapata, Gerente Administrativo, Capstone

S.A de C.V., for Mexican taxation information.



4 Property Description and Location

The Cozamin poly-metallic base metal project is located in the Morelos Municipality of the

Zacatecas Mining District near the south-eastern boundary of the Sierra Madre Occidental

Physiographic Province in north-central Mexico (Figure 4.1). The mine and processing facilities are

located near coordinates 22º 48’ N latitude and 102º 35’ W longitude on 1:250,000 Zacatecas

topographic map sheet (F13-6).

Figure 4.1: Cozamin Mine Location in Central Mexico

The Cozamin project consists of 35 mining concessions covering approximately 2,898 ha (Table 4.1

and Figures 4.2 and 4.3). One additional concession is pending approval for registration with the

Mexican Mines Department.

Mexican mining law requires that the boundaries of a mineral concession be established by a

registered Mexican Mineral Concession Surveyor. These boundaries have been checked by

Capstone surveyors in the mine’s Topography Department, and were determined to be correct.

Capstone purchased 110 ha of surface land rights in 3 parcels at Cozamin from the Ejido Hacienda

Nueva and the Ejido La Pimienta (Figure 4.2).



Table 4.1: Cozamin Claim Status March 2009

Description / Name Title Number Name on Title Claim Classification Claim Area (ha)

001 Plateros 188806 Capstone Exploitation 9.0000

002 Santa lucia 195187 Capstone Exploitation 18.7267

003 San Nicolás 200150 Capstone Exploitation 5.3697

004 San Jacinto Frac. 1 202437 Capstone Exploitation 78.7955

005 San Jacinto Frac. 2 202438 Capstone Exploitation 17.7846

006 Sta. Barbara Frac. 4 202628 Capstone Exploitation 0.4585


008 Gabriela II 203364 Capstone Exploitation 18.9438

009 Plateros Dos 208838 Capstone Exploitation 50.0000

010 La Liga 217237 Capstone Exploitation 20.1817

011 San Bonifacio 217858 Capstone Exploitation 40.8518


013 La Secadora 219630 Capstone Exploitation 9.0000

014 La Providencia 223954 Capstone Exploitation 60.0000

015 Unificación Carlos 224657 Capstone Exploitation 542.5265

016 Orlando 225620 Capstone Exploitation 11.7899

017 San Luis I 223325 Capstone Exploitation 290.6121

018 San Luis II 224466 Capstone Exploitation 133.8409

019 San Luis II Frac. I 224467 Capstone Exploitation 2.1713

020 San Luis II Frac. II 224468 Capstone Exploitation 2.4654

021 Acueducto 224469 Capstone Exploitation 13.5590

022 Acueducto Frac. I 224470 Capstone Exploitation 9.5980

023 La Parroquia 224471 Capstone Exploitation 1.2601

024 La Gloria 224474 Capstone Exploitation 4.1372

025 La Sierpe 224503 Capstone Exploitation 4.2638

026 La Sierpe Frac. I 224504 Capstone Exploitation 0.0108

027 San Judas 226699 Capstone Exploitation 14.5989

028 El Lucero 226834 Capstone Exploitation 145.3505

029 Lorena * 227712 Capstone Exploitation 318.5825

030 Sara 228086 Capstone Exploitation 231.9436

031 El Ranchito 228343 Capstone Exploitation 11.2997

032 El Ranchito f. 1 228344 Capstone Exploitation 0.6189

033 La Veta 228345 Capstone Exploitation 1.4533

034 Anabel 229238 Capstone Exploitation 310.771

(035) Cecilia 230921 Capstone Exploitation 425.6022

Totals 2,898.1308

(36) ximena E.27627 Not yet granted



The project is 100 % owned by Capstone subject to a 3 % net smelter royalty payable to Grupo Bacis

S.A. de C.V. (“Bacis”), a Mexican resource company.

The Cozamin property requires land rental and government fee payments on the mining concessions.

In January 2009 the taxes totalled MX$ 107,234. A similar amount is expected for July 2009.

Figure 4.2 shows a plan view of the Cozamin mining concessions and the surface rights owned or

rented by Cozamin.

Figure 4.2: Cozamin Concession Area



5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

The Cozamin Mine is located in the Western Sierra Madre Physiographic Province near the

boundary with the Mesa Central Province (Central Plateau Province). The Zacatecas area is

characterized by rounded NW trending mountains with the Sierra Veta Grande to the north and the

Sierra de Zacatecas to the south. Elevations on the property vary from 2,400 m to 2,600 m above sea

level (“masl”).

The Zacatecas area is located between forested and sub-tropical regions to the southwest and desert

conditions to the northeast. The climate in the region is semi-arid. Vegetation consists of natural

grasses, mesquite or huizache and crasicaule bushes. Standing bodies of water are dammed as most

streams are intermittent.

The climate in the region is semi-arid with maximum temperatures of approximately 30ºC during the

summer season and minimum temperatures in the winter season producing freezing conditions and

occasional snow. The rainy season extends from June until September. The average annual

precipitation is approximately 500 mm.

The Cozamin project is located 3.5 km to the NNE of the city of Zacatecas, the Zacatecas state

capital. The municipality of Zacatecas has a population of approximately 130,000 people. Other

communities in the immediate vicinity of the project include the following: Hacienda Nueva

(3km W), Morelos (5 km NW) and Veta Grande (5 km N).

Cozamin is accessible via paved roads to the project area boundary where good, all-weather roads

provide access to the mine and most of the surrounding area.

The project area falls within the Hacienda Nueva and La Pimienta Ejido concessions.

The Cozamin Project has excellent surrounding infrastructure including schools, hospitals, railroads,

and electrical power.

The Cozamin Mine is connected to the national power grid with current approval to draw 4.5 MW.

Generators, (both operating and back-up) on site have a capacity of 1.0 MW. The company is

currently constructing a new power line that will allow it to draw up to 7.5 MW from the grid. The

ultimate capacity of the current tailings pond at Cozamin is an additional 10 M tonnes.



Figure 5.1: Surface Layout of the Cozamin Mine Facilities

Figure 5.2: Image of Cozamin looking south with Zacatecas in the background (from Google Earth)

Mill

Tailings Facility

Offices San Ernesto Portal

La Guadalupana Portal

Zacatecas



6 History

In pre-Hispanic times, the area was inhabited by Huichol Indians who mined native silver from the

oxidized zone of argentiferous vein deposits in the Zacatecas Mining District. In 1546, Juan de

Tolosa, guided by a local Indian, arrived in Zacatecas (then Lomas de Bracho) to examine

argentiferous occurrences. In 1548, production commenced at 3 mines: the Albarrada mine on the

Veta Grande system, and the San Bernabe mine and Los Tajos del Panuco on the Mala Noche vein

system. The initial operations worked only the oxides for silver and some gold, and later the

sulphide zones were worked for base and precious metals.

During the Mexican Revolution (1910-1917), mining was essentially halted with flooding and

cave-ins limiting access. Foreign companies worked the mines for base metals from 1936 to 1948

but the lack of electric power, labour problems and low metal prices resulted in closure of

unprofitable mines. From 1972, Consejo de Recursos Minerales (CRM) worked mines in the

El Bote, La Purisima and La Valencia zones.

A number of old workings are located throughout the project area, but accurate records of early

production are not available. Historic production from the Zacatecas district is estimated by the

Consejo de Recursos Minerales (“CRM”) (1992) to be 750,000,000 ounces of silver from 20 million

tonnes grading over 900 g/t Ag and approximately 2.5 g/t Au. Lead, zinc and copper have also been

recovered but the production and grades were not estimated.

Minera Cozamin was established in 1982 by Jack Zaniewicki who consolidated concession holdings

over the Mala Noche vein and operated the San Roberto Mine and plant at 250 t/d until

October 1996. During this period, Industrias Peñoles S.A. de C.V. (“Peñoles) undertook exploration

in the district but did not buy any significant concessions.

In all, it is estimated that 1.2 million tonnes of ore were mined and processed at Cozamin prior to

October 1996.

In October 1996, Zaniewicki sold Cozamin for US$6,800,000 to Minera Argenta, a subsidiary of

Grupo Bacis. Bacis expanded the mill to a 750 tpd flotation plant, and processed 250,000 tonnes of

ore grading 1.2 % Cu, 90 g/t Ag, 0.5 g/t Au, 1.8 % Zn and 0.6 % Pb from 1997 to the end of 1999.

Bacis developed resources principally by drifting and raising on the Mala Noche vein within the

San Roberto (Cozamin) mine. Diamond drilling was only used as an exploration tool to identify

areas with mineralization peripheral to the developed mine workings.

At the end of 1999, Bacis had historic (not 43-101 compliant) resources at San Roberto in all

categories that totalled about 6 million tonnes grading about 1% Cu, 0.9% Pb, 3.2% Zn and silver in

the range of 85 g/t to 105 g/t.



In 1999, Grupo Bacis closed the mine. The principal factors that resulted in the mine closure were

low metal prices and under capitalization. Additionally, poor grade control in the mine and poor

recovery in the plant contributed to the closure.

Diamond drill holes completed by Peñoles and Bacis suggested that the average grade of copper in

the mine might increase with depth but, unfortunately, Grupo Bacis did not have the time and capital

to develop the mine to deeper levels.

Capstone commenced field work at Cozamin in March 2004 and drilling from surface started one

month later. By July 2004, Capstone management decided to significantly expand the initial 5,000

m surface drilling program. The drilling program was expanded again in September 2004.

In November, Capstone management decided to dewater the mine and continue to explore the

deposit by drilling from underground. The company acquired an engineering group headed by

John Wright and Robert Barnes to manage the underground development for the exploration drilling

and the economic appraisal of the deposit. Drilling from surface was completed in March 2005 after

completion of 37 holes totalling 17,968 m. Dewatering of the mine was completed in January 2005.

The hoist, shaft and mine infrastructure were sufficiently rehabilitated in January and February 2005

to allow for the development of the first cross cuts for underground drilling which commenced at the

end of March 2005.

Initial metallurgical studies by SGS Lakefield and PRA were received in January and a feasibility

report for the project was completed by Robert Rodger in February 2006.

The first phase of underground drilling was completed in April, 2006 and totalled 17,736 m in

114 drill holes.

An independent study of geological resources at Cozamin by Gary Giroux was completed in

October 2005. A second resource estimate study was completed in June 2006

The mine and plant were commissioned in June and July 2006 and reached an average daily

production rate of 1,000 t/d in September 2006. A decision to double the mine and plant capacity

was made in October 2006 and this was completed in July 2007.

The second phase of underground drilling was completed in July 2007 and totalled 21,261 m in

69 holes. In addition, a program of 5 surface exploration holes totalling 4,825 m was completed in

April 2007.

Additional drilling in 2008 brought the total drilling from surface and underground at Cozamin to

105,779 m in 365 holes.

Daily production at Cozamin was increased to 3,000 t/d in late 2008.



7 Geological Setting

The Zacatecas Mining District covers a belt of epithermal and mesothermal vein deposits that

contain silver, gold and base metals (copper, lead and zinc). The district is in the southern Sierra

Madre Occidental Physiographic Province near the boundary with the Mesa Central Physiographic

Province in north-central Mexico. The dominant structural features that localize mineralization are of

Tertiary age, and are interpreted by Ponce and Clark (1988) to be related to the development of a

volcanic centre and to northerly trending basin-and-range structures.

The Zacatecas Mining District occurs in a structurally complex setting, associated with siliceous

subvolcanic and volcanic rocks underlain by sedimentary and meta- sedimentary rocks (Figure 7.1).

The geologic units of the Zacatecas area include Triassic metamorphic rocks of the Zacatecas

Formation and overlying basic volcanic rocks of the Upper Jurassic or Lower Cretaceous Chilitos

Formation. The Tertiary rocks consists mainly of a Red Conglomerate unit deposited in Paleocene

and/or Eocene times, and overlying rhyolitic tuff and intercalated flows that were deposited from

Eocene to Oligocene times. Some Tertiary rhyolite bodies cut the Mesozoic and Tertiary units and

formed flow domes.

Zacatecas Formation

The Zacatecas Formation represents the oldest rocks in the district and appears to be equivalent to

the Pimienta Metasediments of Ponce and Clark (1988). The Zacatecas Formation, a marine Upper

Triassic unit, consists of sericite schists, phyllites, slates, quartzites, metasandstone, flint,

metaconglomerate and recrystallized limestone. The unit hosts the El Bote and Pimienta vein

systems to the west of the city of Zacatecas.

Chilitos Formation

The Upper Jurassic to Lower Cretaceous Chilitos Formation is composed of andesitic to basaltic

volcanic rocks with pillow structures and some limestone lenses. The units are referred to as

greenstone of the Zacatecas area and as the Zacatecas Microdiorite by Ponce and Clark (1988).



Mala Noche

City of Zacatecas

0 1000 2000 3000 5000

METERS

N 2520000

N 2525000

N 2530000

E745000

E750000

E755000

Zaragoza

Shaft San Roberto

ShaftSan Rafael

Shaft

Alluvion

Rhyolite Tuff

Andesite + Shale

Andesite Tuff and Flows

Phyllite

Rhyolite Dikes and Domes

Figure 7.1: Regional Geology around the Cozamin Property

Zacatecas Red Conglomerate

The red conglomerate contains fragments of Chilitos and Zacatecas Formation rocks and is probably

of Early Tertiary (Paleocene-Eocene) age. The unit is deposited south of the La Cantera fault in the

structural zone occupied by the city of Zacatecas.

Tertiary Volcanic and Volcaniclastic Rocks

Tertiary volcanic rocks are generally associated with and south of the Zacatecas caldera. They are

described by the CRM (1992) as rhyolitic tuffs with flow intercalations of rhyolite composition that

were extruded during the Oligocene to Eocene. The rhyolitic rocks are reported to have moderate to

high silica content and high potassium content.

A very small group of epiclastic deposits (“Ted”) occur in a road cut near the Bufa flow dome and

small areas of chemical sediments (“Tcs”) are present in the western flank of the Zacatecas caldera

(Ponce and Clark, 1988).



Rhyolitic Subvolcanic Bodies

Ponce and Clark (1988) suggest that subvolcanic intrusive phases include silicic subvolcanic bodies,

lava-flow domes, intrusive tuffs, ignimbrite bodies, pipes and autoclastic breccias. The rhyolitic

subvolcanic bodies, generally dikes and subvolcanic bodies, are structurally controlled by radial or

concentric faults and fractures of the caldera structure. The subvolcanic rhyolitic bodies are

concentrated in the central part of the Zacatecas district in a northwest-southeast trending zone.

The host rocks for the Mala Noche vein are intercalated carbonaceous meta- sedimentary rocks and

andesitic volcanic rocks ranging in age from Triassic to Cretaceous, and Tertiary rhyolite intrusive

rocks and flows (Figure 7.1). Mineralization in the Mala Noche vein appears to have been episodic.

A copper-silver dominant phase was one of the last stages of mineralization. Economically, this is

the most important phase of mineralization at Cozamin. In general, this copper-silver phase was

emplaced into an envelope of pre- existing vein hosting moderate to strong zinc and lead

mineralization and moderate silver mineralization. Thus, the host lithology to the vein does not

appear to have influenced the strength of the copper-silver phase of mineralization which is typically

enveloped by earlier vein material.

The close association of the western third of the Mala Noche vein with rhyolite flow domes and the

strength of contained copper mineralization in this sector of the vein support the hypothesis that the

copper mineralization in the San Roberto mine at Cozamin is relatively close to an intrusive center.

Figure 7.2 shows the relationship of the San Roberto mine to the significant rhyolite flow domes that

have been mapped in the area.

Alternatively, faulting may be localized along the contact of the phyllites with the more competent

andesites and lutites. One kilometre to the south of the Mala Noche, mineralization in the Paroquia

mine is hosted by gneissic rocks that are mapped by the Consejo Recursos Minerales as Upper

Jurassic, Zacatecas Formation.

Intercalated and gradational andesite tuffs, flows and carbonaceous lutites are widely distributed over

the property. Within this marine sequence of sediments and volcanics, there are some pillowed

flows of basaltic composition. The Consejo Recursos Minerales has mapped this sequence as

Cretaceous Chilitos Formation.

Rhyolite flows and dikes are spatially associated with the San Roberto mine. Cerro La Sierpe

(500 m NNW of the shaft), Cerro San Gil (1.5 km WNW of the shaft) and Cerro El Grillo (750 m

SSW of the shaft) are all rhyolite flow domes that, together, surround the western third of the Mala

Noche vein. It is this sector of the Mala Noche vein where the only significant copper mineralization

has been found to date. Rhyolite dikes are difficult to distinguish from the massive rhyolite flows.

However, some of the best quartz stockworking at Cozamin occurs within massive rhyolite that does

not display the fluidal textures and polymictic clasts that are common in most of the rhyolite bodies.



Figure 7.2: Plan showing the distribution of mineralized veins near Zacatecas

In the underground workings at Cozamin, the Mala Noche vein has been shown to occupy a system

of anastomosing faults that is principally comprised of the Mala Noche and Elabra faults along with

other less significant faults. Although not all of the fault system is mineralized at any given location,

there have been no other significant mineralized fault zones discovered to date.



Post mineralization offsets of the Mala Noche vein occur along high angle, normal faults that mostly

strike northeast. These offsets are both dip slip and strike slip. The offsets are generally less than

10 m and often there is drag peripheral to the fault which minimizes mining dilution as the vein can

be followed across the fault. In some locations, it is apparent that the cross cutting faults had an

effect on the emplacement and grade of mineralization prior to post mineralization movement on the

fault.

The principal cross faults in the San Roberto mine area displayed on Level 8 and are presented in

Figure 7.3.

Figure 7.3: Cross Faults, Level 8 Cozamin Mine

The Mala Noche is the principal fault associated with mineralization at Cozamin. In the San

Roberto Mine, the Mala Noche strikes WNW (N70-80W) and the dip varies from 38° to 90° to the

north. There is a clear association of higher copper grades with steeper dips of the Mala Noche fault.

Where the Mala Noche is weakly mineralized, it appears that the principal alteration in this fault is

mostly quartz-pyrite.

The El Abra fault is closely associated with the Mala Noche with which it forms an anastamosing set

in strike and dip. Grades in the San Roberto mine are strongest where the two faults coalesce. The

dominant alteration associated with the El Abra fault is silica-calcite-pyrite. On Level 8 immediately

east of the shaft, the drift roof had to be stabilized where the El Abra fault meets the Mala Noche

fault/vein.



The Rosita fault is also sub-parallel to the Mala Noche but mostly lies in the hanging wall. The

principal alteration associated with the Rosita fault is coarse crystalline calcite suggesting that this

fault is possibly post mineralization and quite open.

The San Ernesto fault is best known in the San Ernesto shaft which was sunk 60 m on the fault in the

hanging wall to the Mala Noche at the west end of the San Roberto Mine. The fault strikes WNW

and dips at about 60° to the NNE. Mineralization encountered in the fault to date has been zinc and

lead dominant. This fault and associated mineralization may be related to lenses of hanging wall

zinc found in the western sector of the San Roberto mine.

The Margarita Fault is located about 100 m west of the shaft on Level 8. The fault strikes NNE and

dips at 70° to the WSW. Movement on the fault appears to be minimal as indicated by the mapping

to date. Minor argillic alteration is associated with the fault.

The Josefina fault is found on Level 8 about 50 m west of the shaft. The fault strikes SE and dips at

about 55° degrees to the NE. Movement on the fault appears to be dextral with a displacement of

about 5 m. Minor argillic alteration is found in the fault zone.

The Lorena fault is located about 25 m west of the shaft on Level 8. This fault strikes NE and dips at

about 70o to the SE. Post mineralization movement on the Lorena fault appears to be less than 2

metres and only weak argillic alteration is found within the fault. The intersection of the Lorena and

Josefina faults on Level 8 resulted in poor roof stability in the area of a prior electrical substation

35 m west of the shaft.

On Level 8, the Anabel Fault is found 155 m east of the shaft. The fault strikes NNE and dips E at

about 60o. Movement on the fault appears to be dextral strike slip with possibly some normal dip

slip displacement. The projection of the Mala Noche vein is offset about 10 m horizontally along

this fault. However, there has been significant drag on the west side of the fault resulting in minimal

displacement of the vein across the fault plane. Mineralization west of this fault is strongly

diminished. Alteration in the Anabel fault is principally silicification.

The Lupita fault is located 255 m E of the shaft on Level 8. The fault strikes NE and dips at about

65° to the SE. Displacement on the fault appears to be minimal and only minor silicification is

associated with the fault.

The Karla fault is located 465 m east of the shaft on Level 8. This fault has been mapped only on

Level 8. Its strike is NE and the fault dips 65o SE. Apparent horizontal offset on the fault is about

3 m as a result of normal dip slip or possible dextral strike slip displacement. There is no significant

drag or alteration associated with this fault.



8 Deposit Types

All mineralization at Cozamin occurs in veins. The main stage of copper–dominant stage of

mineralization at Cozamin can be classified as intermediate sulphidation, high temperature

epithermal or, more likely, mesothermal mineralization. The copper-dominant stage of

mineralization appears to cut across earlier epithermal zinc-dominant mineralization. The epithermal

veins display well banded quartz veins and open space fillings and quartz druse vug linings. The

higher temperature veins have significantly less vugs, and the veins can be massive pyrrhotite-pyrite-

chalcopyrite.

This transition from epithermal zinc dominant mineralization to copper-dominant mesothermal

mineralization is thought to be the result of an evolving, telescoping hydrothermal system that was

epithermal in its early stages and became mesothermal as the hydrothermal migrated upwards.

Chalcopyrite-pyrite-pyrrhotite mineralization can be seen to cut earlier sphalerite-galena-pyrite

mineralization in drill core and workings.

Zones of massive pyrrhotite along with apparent retrograded calcsilicates suggest mineral deposition

in a mesothermal environment that was superimposed on epithermal alteration and mineralization.

This telescoping hydrothermal system is closely associated with the district’s largest center of

rhyolite flow domes that may be the upward expression of a felsic stock.



9 Mineralization

The dominant mineralized vein on the Cozamin project is the Mala Noche. This vein has been

traced for 5.5 km on surface on the property. It strikes approximately east-west and dips on average

at 60º to the north. There are at least 18 shafts that provide access to the historical workings at

Cozamin (Figure 7.3). The largest of these is the San Roberto mine which has a strike length of 1.4

km. Mineralization peripheral to these workings was the principal target of Capstone’s exploration

at Cozamin.

The Mala Noche vein system occupies a system of anastomosing faults. The mineralized bodies

within the Mala Noche appear to be strongest where the disparate faults coalesce into a single fault

zone. Results from the exploration and mine development to date indicate that some of the strongest

mineralization in the San Roberto mine rakes to the west at approximately -50º within the vein. Post

mineralization offsets of the Mala Noche vein are minimal and occur along high angle, normal faults

that strike northeast.

Moderate propyllitic wall rock alteration is generally limited to 3 m into the hanging wall and

footwall. Gangue minerals in the Mala Noche vein consist of quartz, silica, calcite, chlorite, epidote

and minor disseminated sericite. The quartz occurs as coarse grained druse coarse crystalline

masses, and a stockwork of quartz veinlets.

Mineralization in the Mala Noche vein at Cozamin appears to have been episodic. Early epithermal

mineralization and alteration (represented by sulphide pseudomorphs of carbonates and possibly

barite and well-banded quartz veins and vug linings of quartz druse) have been overprinted by higher

temperature pyrite-pyrrhotite-chalcopyrite dominant mineralization in a telescoped, intrusive related

hydrothermal system. The Mala Noche vein in the San Roberto mine workings shows contained

sulphides to occur as disseminations, bands and masses. Considering the limited exposure of the

copper-silver phase of mineralization in the current depths of the mine workings, conclusions about

mineralization styles at this point in time are preliminary.

Pyrite is the dominant vein sulphide and typically comprises approximately 15 % of the Mala Noche

vein in the San Roberto mine. It occurs as fine disseminations and veinlets, coarse crystalline

replacements, and pseudomorphs of possible epithermal carbonates such as barite and calcite.

Pyrrhotite is the second most common sulphide mineral but is present only in the intermediate and

deeper levels of the San Roberto mine. It occurs as replacement masses, pseudomorphs of platey

masses and acicular replacements probably after amphibole. Pyrrhotite commonly occurs as an

envelope to, or intermixed with, strong chalcopyrite mineralization.

Chalcopyrite is the only copper sulphide recognized megascopically at Cozamin. Like pyrrhotite, it

is more common at the intermediate and deeper levels of the mine.



It occurs as disseminations, veinlets and replacement masses. These masses appear to be fractured

and brecciated at intermediate levels in the mine.

Sphalerite is the dominant economic sulphide in the upper levels in the San Roberto mine. Most of

the sphalerite is marmatitic. It occurs as disseminations and coarse crystalline masses and is

commonly marginal to the chalcopyrite-dominant portion of the vein.

Galena is less common than sphalerite but is generally associated with it. Where it is abundant, it

occurs as coarse crystalline replacement masses. Both coarse and fine crystalline masses of galena

are argentiferous.

Arsenopyrite typically occurs as minor, microscopic inclusions in pyrite.

Argentite is the most common silver mineral. It has been identified microscopically occurring as

inclusions in chalcopyrite and pyrite. Assays indicate that silver is also probably present in

sphalerite and galena.

Bismuth and silver selenides occur as inclusions predominantly in chalcopyrite and pyrite.

The main gangue minerals are quartz and calcite with rhodochrosite, barite and gypsum also

reported.



10 Exploration

The area with the largest historic workings on the Cozamin project, the San Roberto mine, was

selected as the principal exploration target. Widely spaced exploration drill holes by prior operators

at Cozamin suggested that mineralization extended at least 100 m below Level 9; the deepest level

developed in the San Roberto mine which was allowed to flood at the end of 1997. These intercepts

indicated that mineralization in the Mala Noche vein had significantly higher copper concentrations

beneath the historic mine workings.

In 2004, Capstone decided to drill test the Mala Noche vein beneath the historic workings of the San

Roberto mine. The initial three drill sections, comprised of two drill holes each, all intersected

significant economic mineralization over true widths varying from 3.2 m to 14.9 m. These three drill

sections were distributed over 550 m of strike extent beneath the historic workings. At that point,

Capstone decided to drill single hole sections every 100 m beneath the San Roberto workings. These

holes targeted the Mala Noche at approximately 2,150 metres above sea level (masl) which is 65 m

below the historic workings. This strategy resulted in the first 20 exploration holes being distributed

over a strike length of 1.4 km. Of these first 20 drill holes, 17 intersected significant mineralization

that averaged 6.64 m in true width and had weighted grade averages of 2.61% Cu, 91.25 g/t Ag and

1.38% Zn.

These significantly higher copper grades and undiminished silver grades are associated with

significant amounts of pyrrhotite. This reinforced the company’s conviction that the historic

workings at San Roberto are located just above the upper reaches of a large copper-silver

mineralized system of mesothermal character. Subsequent exploration drilling showed that the

copper-silver dominant phase of mineralization extends below 1,865 masl which is 350 m below the

historic workings.

A summary of the completed exploration phases is presented below:

Preliminary Work

Exploration by Capstone on the Cozamin project commenced with engineering examinations by

Capstone directors Peter Kuhn, P.Eng. and Jack Marr, P.Geo. Site examinations were conducted by

Peter Christopher, P. Eng. in 2003 and 2004. Christopher collected two rock chip samples from the

Virginias mine decline and 24 splits of half core from mineralized intervals in diamond drill holes

previously drilled by Bacis. These samples were submitted to Acme Laboratories in Vancouver for

copper, lead, zinc, gold, and silver assays and ICP analyses. The assay results confirmed Bacis

records and the Phase I drilling program commenced in March 2004 under the supervision of

qualified person (QP) Hugh Willson, Capstone’s Vice President of Exploration. Preliminary

underground sampling was not completed because most of the mineralized underground workings

were flooded.



Phase I

Budget $US1,000,000 (undertaken March 2004 - August 2004)

Mapped 5.5 km of the surface trend of the Mala Noche vein system.

Completed CSAMT (8 line kilometres) and NSAMT (16 line kilometres) with magnetic survey

(26 line kilometres) over the Mala Noche vein system (Zonge Engineering and Research

Organization).

Completed a 7,484.44 m surface NQ-diameter diamond drill program (holes CG-04-01 to

CG-04-19).

Completed an independent review (Hawthorne, 2004) of the existing plant and mill to determine

cost of rehabilitation and expansion.

Phase II

Budget $US2,500,000 (undertaken September 2004 - March 2005)

Further evaluation of geophysical results.

Completed a 10,483.27 m surface NQ-diameter diamond drill program (holes CG-04-20 to

CG-04-37) that mainly tested the Mala Noche vein at elevations between the 1,900 m and 2,050

m level below old workings in the San Roberto mine.

Completion of preliminary metallurgical test by SGS Lakefield.

Phase III

Budget $US4,537,500 (undertaken April 2005 - April 2006)

Metallurgical study completed by Process Research Associates Ltd.

Clifton Associates Ltd. of Guadalajara, Jalisco, Mexico and Nimbus. Management Ltd. of

Vancouver submitted an environmental impact assessment (MIA), an impact study for land use

(ETJ) and a risk assessment (ER) to the Mexican federal regulatory agency in charge of

environmental issues.

Completed a 17,687.70 m underground definition NQ-diameter diamond drill program (holes

CG-U01 to CG-U114).

Initial resource estimate prepared in October 2005 by Giroux Consultants Ltd. based on the

37 surface drill holes, 66 underground drill holes and 48 underground channel samples.

Feasibility study completed in March 2006 by RJR Mineral Services.

Updated resource prepared in July 2006 by Giroux Consultants Ltd. incorporating assay results

from all surface and underground diamond drill holes, and 768 additional channel samples from

the initial 2005 estimate.



Phase IV and V

Combined budget $US6,000,000 (undertaken October 2006 - July 2007)

Completed a 4,824.56 m surface PQ/NQ-diameter diamond drill program (holes CG-06-38 to

CG-06-39 and CG-07-40 to CG-07-42) that tested the Mala Noche vein at elevations between

approximately 600-700 m below surface of the San Roberto mine.

Completed a 21,441.10 m underground NQ-diameter diamond drill program (holes CG-06-U115

to CG-06-U124, and CG-07-U125 to CG-07-U183). These holes were designed to infill and

extend the 2006 estimated resources.

Significant results of these drill programs are presented in Section 11.

Approximately 4,200 m of underground NQ-diameter diamond drilling was completed from July 31

to December 13, 2007. Fifteen holes were drilled to infill the resource as well as to extend some

holes that did not appear to completely cut the entire mineralized zone as defined in previous drill

programs.

In 2008, a total of 39,430 m of surface and underground diamond drilling was completed at Cozamin

thereby warranting this revised and updated 43-101 compliant report. Figure 10.1 shows a plan of

the Mala Noche vein and the holes drilled from surface. Figure 10.2 presents a longitudinal section

of the San Roberto mine with the pierce points for drill holes cutting the Mala Noche vein.

Phase VI

Combined budget $US5,000,000 (undertaken November 2007 – December 2008)

Completed 30,430m of HQ diameter diamond drilling from surface (holes CG-08-43 to CG-08-

150)

Completed 9,000m of NQ diameter diamond drilling from underground (holes CG-07-U184 to

CG-08-U205 and also some prior holes were extended)

The underground drilling was targeted on areas in the San Roberto deposit that needed infill drilling

to attain a higher level of confidence and a higher resource classification. Some holes were used to

extend the resources to greater depth.

The surface drilling was focussed on the San Rafael zone to the east of the San Roberto deposit and

also used to augment the underground infill drilling and to test the continuity of mineralization in the

Zaragoza area at the west end of the San Roberto deposit.

A longitudinal section of the Mala Noche vein showing the pierce points of drill holes drilled from

both surface and underground is presented in Figure 10.1.

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11 Drilling

In all, 150 surface and 216 underground exploration holes have been drilled at Cozamin in the period

from April 2004 through December 2008 which is the cut off for the current resource and reserve

estimates. A summary of the drilling companies and down hole survey tools used for the drilling is

presented in Table 11.1.

Table 11.1: Drilling History by Contractor

Contractor Year Phase Holes Drilled

MetresDrilled

Downhole Survey Instrument

SURFACE

Britton Brothers 2004-2005 I 37 17,967.71 Eastman Single Shot

Major Drilling 2006-2007 V 5 4,824.56 FLEXIT SensIT

Major Drilling 2008 VI 107 30,430 Reflex EZ-Shot

UNDERGROUND

Canrock Drilling 2005-2006 II 78 10,031.24 Reflex EZ-Shot

Explor 2005 II 1 305.70 Reflex EZ-Shot

Tecmin 2005-2006 III 35 7,417.37 Reflex EZ-Shot

Tecmin 2006-2007 IV 180 25,259.70 Reflex EZ-Shot

Tecmin 2008 VI 22 8,999.55 Reflex EZ-Shot

Drill holes are located using a total station TOPCON instrument, model GTS-236W. Down hole

survey readings were recorded using either an Eastman Single Shot, FLEXIT SensIT or Reflex EZ-

Shot instrument as indicated in Table 11.1. Survey readings are generally taken every 50-150 m for

surface holes and every 50-100 m for underground holes. Survey results have been corrected for

magnetic declination (+8º).

Dip variations in surface holes are not more than 5.3º, with an average value of 1.1°. The maximum

down hole dip variation in the underground holes is 15.4º with an average variation of 1.3°.

Descriptions of each drilling phase completed on the Cozamin project and a summary of significant

intersections are presented below. True widths are estimated by correcting for strike and dip of the

vein with regard to the bearing and inclination of the drill hole. Where no assay is recorded an

intercept was not calculated.

Phases I-III

The Phase I and II surface exploration drill programs totalled 17,967.71 m of NQ-diameter diamond

core in 37 holes that were drilled in 2004 and the first quarter of 2005. The Phase III underground

definition drill program consisted of 114 holes that totalled 17,736.31 m of NQ-diameter diamond

core that were drilled in 2005 and the first half of 2006.



Significant intersections are presented in Tables 11.2 and 11.3. True widths are estimated by

correcting for strike and dip of the vein with regard to the bearing and inclination of the drill hole.

The reader is referred to Capstone news releases dated 2005 and 2006 for a complete list of drill

results.

Table 11.2: Phase I and II Surface Drilling Results

HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

S1 M.N. Vein 289.00 293.00 4.00 2.92 0.034 108.24 1.67 0.70 2.76

S2 M.N. Vein 306.80 311.55 4.75 3.02 0.222 125.85 2.78 0.12 0.59

S3 M.N. Vein 301.35 321.80 20.45 17.26 0.031 73.69 3.15 0.12 0.44

S4 M.N. Vein 341.35 351.83 10.48 8.11 0.161 34.78 1.65 0.00 0.74

S5 M.N. Vein 353.87 355.90 2.03 1.82 0.194 203.97 3.00 0.33 2.29

S6 M.N. Vein 412.73 428.39 15.66 14.02 0.032 43.54 1.31 0.02 0.26

S7 M.N. Vein 377.00 380.00 3.00 2.36 0.130 177.02 1.66 1.97 12.70

S7 fw 418.65 420.05 1.40 1.10 0.011 21.68 1.05 0.00 1.79

S8 M.N. Vein 337.65 339.75 2.10 1.84 0.071 139.76 2.47 0.03 2.19

S9 M.N. Vein 307.00 307.70 0.70 0.50 0.068 40.46 0.82 0.03 0.06

S10 M.N. Vein 483.30 501.40 18.10 12.34 0.013 77.69 2.10 0.15 1.50

S11 M.N. Vein 367.56 373.85 6.29 4.21 0.082 118.76 2.28 0.10 2.56

S12 hw 169.16 171.25 2.09 1.54 1.200 162.87 0.10 0.99 7.98

S12 hw 176.10 177.00 0.90 0.66 0.148 79.27 1.33 0.45 10.75

S12 M.N. Vein 291.50 296.75 5.25 3.95 0.137 135.82 2.69 0.39 1.67

S13 M.N. Vein 318.60 333.45 14.85 9.51 0.103 103.49 1.17 0.97 1.44

S14 hw 40.85 42.85 2.00 1.37 0.000 8.16 0.06 0.05 4.94

S14 M.N. Vein 246.16 247.65 1.49 1.02 0.039 4.88 0.08 0.00 0.42

S15 M.N. Vein 294.00 302.75 8.75 6.83 0.011 68.59 3.45 0.03 1.63

S16 M.N. Vein 317.35 321.85 4.50 3.16 0.026 63.84 1.65 0.03 0.37

S17 hw 321.79 323.59 1.80 1.10 0.112 19.11 0.07 2.53 9.32

S17 M.N. Vein 348.60 355.69 7.09 4.01 0.233 36.68 0.42 0.47 2.79

S18 hw 331.00 335.55 4.55 1.82 0.120 44.07 1.00 0.95 1.72

S18 M.N. Vein 344.40 361.40 17.00 6.81 0.024 78.39 3.48 0.07 0.35

S19 M.N. Vein 191.75 198.30 6.55 2.76 0.162 85.78 1.49 0.53 2.01

S20 M.N. Vein 312.70 335.70 23.00 16.57 0.366 73.11 0.71 0.44 3.68

S21 M.N. Vein 466.00 472.90 6.90 3.58 0.021 14.37 0.33 0.03 7.01

S22 M.N. Vein 450.65 452.30 1.65 1.24 0.121 30.19 1.67 0.01 0.06

S23 M.N. Vein 134.95 141.80 6.85 5.31 0.448 6.12 0.01 0.30 0.75

S24 M.N. Vein 416.95 417.85 0.90 0.54 0.404 18.20 0.09 0.31 9.37

S24 fw 449.90 451.30 1.40 0.83 0.023 42.47 0.16 0.38 5.80

S25 M.N. Vein 268.10 269.05 0.95 0.89 0.145 49.57 0.36 0.12 6.86

S25 fw 281.70 282.00 0.30 0.28 4.600 92.20 0.04 6.73 8.28

S26 M.N. Vein 647.00 666.00 19.00 8.49 0.014 21.93 1.11 0.02 0.14

S27 hw 345.00 345.70 0.70 0.49 0.172 27.64 0.16 0.71 5.43

S27 M.N. Vein 585.80 593.74 7.94 5.85 0.243 62.13 1.81 0.03 0.80

S28 hw 628.20 628.90 0.70 0.52 0.013 39.17 1.79 0.01 0.06

S28 M.N. Vein 639.30 644.40 5.10 3.79 0.011 40.75 1.99 0.01 0.11

S29 M.N. Vein 510.20 513.24 3.04 2.82 0.020 189.90 6.63 0.08 1.17

S30 M.N. Vein 599.00 601.80 2.80 2.52 0.003 18.30 1.08 0.01 0.05



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

S1 M.N. Vein 289.00 293.00 4.00 2.92 0.034 108.24 1.67 0.70 2.76

S31 M.N. Vein 463.20 464.00 0.80 0.65 0.098 57.50 0.19 3.33 2.74

S32 M.N. Vein 192.61 194.41 1.80 1.52 0.089 15.12 0.12 0.35 1.12

S33 M.N. Vein 664.64 667.74 3.10 2.37 0.004 19.41 0.28 0.04 5.22

S34 hw 104.60 106.70 2.10 1.76 1.216 64.30 0.10 6.34 6.68

S34 M.N. Vein 565.90 572.50 6.60 5.53 0.010 39.95 1.70 0.02 0.08

S35san

ernesto 251.25 253.50 2.25 1.60 0.053 106.77 1.27 0.07 3.81

S35 M.N. Vein 478.50 482.00 3.50 2.48 0.002 0.26 0.00 0.00 0.00

S36 M.N. Vein 458.50 459.45 0.95 0.67 0.450 40.02 0.01 0.06 1.41

S37 hw 450.70 452.85 2.15 1.37 0.134 12.88 0.08 0.72 2.75

S37 hw 455.30 456.30 1.00 0.64 0.502 21.45 0.04 3.85 1.72

S37 M.N. Vein 508.20 509.20 1.00 0.63 1.550 25.65 0.16 2.22 3.14

Table 11.3: Phase III Underground Drilling Results

HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

U01 hw 62.80 63.30 0.50 0.19 0.007 36.10 1.13 0.00 1.65

U01 M.N. Vein 67.90 69.10 1.20 0.46 0.059 29.03 0.73 0.15 0.69

U02 M.N. Vein 44.10 50.80 6.70 6.60 0.049 90.78 2.17 0.35 0.60

U03 M.N. Vein 51.60 65.85 14.25 11.62 0.035 59.64 1.02 0.14 1.79

U04 M.N. Vein 62.00 68.40 6.40 3.53 0.176 82.37 1.17 1.66 0.88

U05 M.N. Vein 42.10 58.30 16.20 13.80 0.194 102.36 1.14 2.77 1.11

U06 M.N. Vein 49.25 58.30 9.05 7.38 0.061 101.00 1.78 1.68 1.18

U07 hw 32.00 35.70 3.70 3.13 0.141 106.15 0.66 7.85 7.57

U07 M.N. Vein 42.78 51.50 8.72 7.39 0.066 49.69 1.10 0.17 0.61

U07 fw 52.90 54.40 1.50 1.27 0.047 32.78 0.62 0.01 1.45

U08 M.N. Vein 39.70 53.70 14.00 9.23 0.122 51.55 0.73 2.39 2.93

U09 M.N. Vein 50.80 54.80 4.00 3.44 0.020 51.95 1.13 0.02 0.26

U10 M.N. Vein 48.75 57.30 8.55 5.91 0.033 62.44 2.10 0.10 0.54

U11 M.N. Vein 52.55 67.06 14.51 11.86 0.052 57.88 2.13 0.02 0.23

U12 M.N. Vein 57.50 63.30 5.80 3.66 0.020 22.95 0.90 0.01 0.03

U12 fw 68.60 69.40 0.80 0.51 0.016 28.00 0.79 0.01 8.59

U13 M.N. Vein 55.00 57.50 2.50 1.95 0.023 2.49 0.02 0.03 0.27

U14 M.N. Vein 61.90 67.06 5.16 2.60 0.023 3.39 0.01 0.04 0.25

U15 M.N. Vein 49.10 51.00 1.90 1.52 0.044 81.87 2.39 0.23 0.72

U16 hw 34.75 35.70 0.95 0.47 0.381 39.35 0.10 2.81 9.87

U16 M.N. Vein 58.20 59.70 1.50 0.75 0.152 54.63 0.52 0.31 1.75

U17 M.N. Vein 35.70 52.40 16.70 13.62 0.065 103.07 1.09 2.37 0.52

U17 fw 60.70 61.90 1.20 0.98 0.379 32.80 0.48 3.32 6.72

U18 hw 34.10 35.20 1.10 0.74 0.517 34.24 0.08 5.63 10.36

U18 M.N. Vein 39.70 53.00 13.30 9.00 0.045 96.95 1.39 0.81 0.60

U19 M.N. Vein 39.40 62.10 22.70 14.29 0.092 103.39 1.33 3.00 1.62



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

U01 hw 62.80 63.30 0.50 0.19 0.007 36.10 1.13 0.00 1.65

U20 M.N. Vein 45.90 75.60 29.70 22.52 0.095 98.38 1.06 1.15 1.41

U21 M.N. Vein 50.90 65.00 14.10 6.97 0.039 49.52 0.72 0.17 1.89

U22 M.N. Vein 71.00 72.90 1.90 0.97 0.012 58.01 2.88 0.01 0.08

U23 M.N. Vein 45.50 56.40 10.90 7.86 0.295 3.17 0.01 0.03 0.80

U24 M.N. Vein 53.75 64.00 10.25 5.02 0.117 6.37 0.04 0.05 1.02

U25 M.N. Vein 55.75 63.25 7.50 5.03 0.265 90.94 1.26 0.16 0.58

U26 M.N. Vein 51.10 54.35 3.25 2.79 0.024 70.15 1.31 0.02 2.30

U27 M.N. Vein 42.09 49.33 7.24 7.23 0.083 78.59 1.32 0.10 4.07

U28 M.N. Vein 56.87 59.87 3.00 2.14 0.057 43.59 0.60 0.14 1.56

U29 M.N. Vein 57.90 61.25 3.35 2.41 0.057 105.24 1.61 0.59 1.69

U30 M.N. Vein 63.25 70.00 6.75 2.91 0.074 55.75 1.07 0.07 1.11

U31 hw 40.75 43.50 2.75 1.75 0.058 8.13 0.02 0.32 5.81

U31 M.N. Vein 55.50 61.50 6.00 3.82 0.206 67.40 0.83 0.77 6.18

U32 M.N. Vein 84.50 102.00 17.50 2.50 0.067 127.14 2.61 0.59 3.43

U33 M.N. Vein 65.00 70.00 5.00 2.48 0.089 82.67 0.93 0.57 3.02

U34 M.N. Vein 102.00 132.50 30.50 23.84 0.075 116.01 1.97 0.45 0.98

U35 M.N. Vein 144.00 169.00 25.00 17.03 0.039 81.16 1.97 1.10 0.67

U36 M.N. Vein 116.50 129.25 12.75 6.79 0.042 124.65 3.67 0.44 3.26

U37 M.N. Vein 161.75 164.75 3.00 1.06 0.011 74.07 4.33 0.01 0.49

U38 M.N. Vein 131.75 143.00 11.25 8.72 0.085 90.72 2.34 2.28 1.26

U39 hw 147.00 148.00 1.00 0.44 0.026 24.90 0.43 0.21 4.59

U39 M.N. Vein 155.00 175.85 20.85 9.18 0.045 96.14 2.82 0.17 1.37

U39 fw 183.50 184.50 1.00 0.44 0.033 65.05 3.26 0.01 0.12

U40 hw 138.25 139.10 0.85 0.49 0.002 5.61 0.07 0.01 10.51

U40 M.N. Vein 142.75 149.25 6.50 3.78 0.048 72.41 2.32 0.06 2.42

U41 M.N. Vein 114.00 114.50 0.50 0.45 0.039 38.25 1.04 0.03 0.70

U42 M.N. Vein 123.50 132.75 9.25 7.41 0.057 50.77 1.47 0.52 1.52

U43 M.N. Vein 129.25 134.25 5.00 3.60 0.026 81.22 2.67 0.11 0.53

U44 M.N. Vein 101.50 115.25 13.75 10.57 0.012 38.76 1.48 0.01 0.05

U45 M.N. Vein 114.00 127.50 13.50 10.06 0.066 104.32 2.83 0.33 1.29

U46 M.N. Vein 113.50 128.75 15.25 13.45 0.065 175.92 3.38 3.42 0.99

U47 M.N. Vein 103.50 112.75 9.25 7.32 0.029 50.29 1.87 0.01 0.07

U48 hw 120.40 121.25 0.85 0.58 0.008 22.98 0.56 0.01 2.85

U48 M.N. Vein 134.00 155.00 21.00 14.41 0.040 79.55 1.52 1.13 0.83

U49 M.N. Vein 94.00 105.00 11.00 10.45 0.485 15.85 0.03 0.30 1.39

U50 M.N. Vein 108.50 123.50 15.00 11.04 0.052 84.11 1.89 0.20 0.79

U51 M.N. Vein 93.75 96.75 3.00 2.85 0.057 54.28 1.27 0.11 0.08

U51 fw 104.50 105.50 1.00 0.95 0.018 32.10 1.14 0.02 0.06

U52 M.N. Vein 115.50 118.00 2.50 1.78 0.027 67.40 1.74 0.37 2.99

U52 fw 120.00 122.75 2.75 1.96 0.009 11.02 0.54 0.01 8.27

U52 fw 127.50 128.15 0.65 0.46 0.014 64.99 0.23 0.90 11.80

U53 M.N. Vein 166.50 179.75 13.25 3.40 0.120 75.68 1.50 0.67 2.84



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

U01 hw 62.80 63.30 0.50 0.19 0.007 36.10 1.13 0.00 1.65

U54 hw 130.25 132.00 1.75 0.64 0.133 68.16 0.86 1.20 2.70

U54 M.N. Vein 161.00 184.50 23.50 7.45 0.012 51.21 2.79 0.01 0.09

U55 M.N. Vein 77.50 79.50 2.00 1.18 1.258 59.78 0.37 1.25 4.00

U56 M.N. Vein 53.00 59.00 6.00 5.52 0.094 87.43 1.15 0.33 3.00

U57 M.N. Vein 117.50 118.50 1.00 0.84 0.006 3.30 0.16 0.00 8.15

U58 M.N. Vein 67.25 80.00 12.75 7.93 0.095 90.73 2.21 0.15 3.18

U59 M.N. Vein 60.00 63.00 3.00 2.72 0.116 67.78 1.42 0.60 4.08

U60 M.N. Vein 76.00 78.75 2.75 1.87 0.088 188.35 2.14 6.59 5.93

U61 hw 134.75 138.00 3.25 1.13 0.066 206.47 1.76 10.38 3.63

U61 M.N. Vein 185.50 192.50 7.00 2.25 0.014 59.72 3.52 0.01 0.14

U62 M.N. Vein 119.00 157.50 38.50 24.12 0.028 74.12 2.71 0.13 0.71

U63 hw 90.50 91.50 1.00 0.92 0.026 7.85 0.25 0.01 9.57

U63 M.N. Vein 106.50 109.75 3.25 2.98 0.090 17.99 0.19 0.08 2.62

U64 M.N. Vein 126.00 135.50 9.50 7.19 0.117 21.08 0.59 0.03 5.62

U65 no

intersection

U66 M.N. Vein 157.00 159.25 2.25 1.25 0.012 45.61 1.74 0.01 0.13

U67 M.N. Vein 144.50 148.50 4.00 2.17 0.069 80.71 2.65 0.41 0.44

U68 M.N. Vein 124.00 130.00 6.00 4.13 0.118 108.12 2.92 0.07 1.34

U69 M.N. Vein 176.00 180.00 4.00 0.89 0.042 84.86 4.80 0.01 0.28

U70 hw 95.00 98.00 3.00 2.43 0.057 60.18 1.21 1.14 1.23

U70 M.N. Vein 116.00 123.50 7.50 6.08 0.039 114.23 4.48 0.22 0.16

U71 M.N. Vein 170.50 175.50 5.00 1.98 0.022 145.18 7.71 0.02 0.66

U71 fw 179.50 180.50 1.00 0.40 0.147 50.55 2.78 0.02 1.38

U72 M.N. Vein 167.50 174.50 7.00 3.01 0.042 190.76 5.37 0.40 1.38

U73 hw 80.50 82.00 1.50 1.04 0.322 118.20 1.43 10.61 2.13

U73 M.N. Vein 99.00 126.00 27.00 19.37 0.030 60.73 1.95 0.06 0.27

U74 M.N. Vein 50.50 53.50 3.00 2.89 0.067 126.37 2.68 0.16 2.80

U75 M.N. Vein 154.50 166.00 11.50 6.14 0.019 108.28 5.89 0.03 1.65

U76 M.N. Vein 72.75 91.50 18.75 11.40 0.103 94.30 2.69 0.10 0.76

U77 M.N. Vein 252.75 257.00 4.25 2.12 0.027 45.89 1.14 0.14 1.89

U78 hw 73.00 75.00 2.00 1.77 0.224 54.23 0.93 3.67 5.12

U78 M.N. Vein 93.50 107.00 13.50 11.97 0.042 87.07 3.01 0.59 1.59

U79 M.N. Vein 160.50 165.50 5.00 3.14 0.018 63.31 1.66 0.05 2.11

U80 M.N. Vein 196.00 214.50 18.50 3.33 0.041 55.76 2.83 0.02 1.38

U81 M.N. Vein 107.00 136.00 29.00 16.40 0.028 34.70 0.85 0.30 0.21

U82 M.N. Vein 187.50 197.00 9.50 2.62 0.022 75.75 2.75 0.04 0.60

U83 hw 140.50 143.50 3.00 0.23 0.220 86.78 1.70 6.55 1.15

U83 hw 156.75 164.50 7.75 0.59 0.071 32.23 0.30 0.92 3.43

U83 M.N. Vein 175.00 189.00 14.00 1.33 0.029 92.97 3.29 0.14 0.67

U84 M.N. Vein 146.50 158.00 11.50 1.73 0.020 65.37 3.02 0.02 2.28

U85 M.N. Vein 176.75 187.00 10.25 5.70 0.074 64.82 1.76 0.04 1.80



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

U01 hw 62.80 63.30 0.50 0.19 0.007 36.10 1.13 0.00 1.65

U86 M.N. Vein 188.50 190.50 2.00 0.93 0.040 111.85 3.70 0.03 0.15

U87 M.N. Vein 159.00 169.00 10.00 6.71 0.189 84.76 2.09 0.16 1.91

U88 M.N. Vein 181.50 194.00 12.50 3.87 0.078 92.53 2.66 0.09 1.53

U89 M.N. Vein 146.50 148.50 2.00 1.28 0.091 52.08 1.67 0.06 0.13

U90 M.N. Vein 49.50 51.50 2.00 1.90 0.103 34.60 0.55 0.03 1.33

U91 hw 105.00 106.00 1.00 0.31 0.190 16.60 0.32 0.78 7.45

U91 hw 131.50 135.00 3.50 1.10 0.171 72.56 2.23 0.04 0.11

U91 hw 146.50 147.50 1.00 0.27 0.011 46.90 1.24 0.03 0.07

U91 M.N. Vein 166.00 170.25 4.25 1.17 0.018 95.15 4.46 0.02 0.37

U92 M.N. Vein 53.00 66.00 13.00 11.25 0.172 43.85 0.53 0.55 2.56

U93 M.N. Vein 58.00 70.00 12.00 3.75 0.124 120.13 1.13 0.52 1.79

U93 fw 86.00 88.00 2.00 0.63 0.103 69.48 1.67 0.02 0.34

U94 hw 258.50 259.50 1.00 0.21 0.007 32.55 1.77 0.01 0.08

U94 M.N. Vein 286.00 306.00 20.00 4.08 0.018 103.75 5.33 0.11 1.37

U95 M.N. Vein 256.00 261.00 5.00 2.93 0.021 73.78 2.70 0.42 0.23

U96 M.N. Vein 248.50 260.50 12.00 6.34 0.032 70.67 3.12 0.07 1.10

U97 M.N. Vein 235.00 241.25 6.25 3.63 0.024 31.36 1.13 0.03 0.31

U97 fw 247.00 248.25 1.25 0.73 0.011 28.58 1.40 0.00 0.04

U98 hw 193.00 194.00 1.00 0.30 0.068 63.30 0.68 3.31 4.02

U98 M.N. Vein 247.25 279.25 32.00 9.64 0.011 54.44 2.04 0.43 1.70

U99 M.N. Vein 204.50 221.00 16.50 12.57 0.018 65.02 2.64 0.09 0.63

U100 M.N. Vein 215.00 240.25 25.25 14.57 0.023 53.70 1.41 1.13 0.78

U101 M.N. Vein 225.75 253.25 27.50 8.36 0.048 104.74 4.28 0.50 1.23

U102 M.N. Vein 269.50 278.00 8.50 4.14 0.031 82.36 2.78 0.05 1.18

U103 hw 139.00 149.50 10.50 5.39 0.082 212.05 2.77 8.18 3.96

U103 hw 161.50 162.25 0.75 0.39 0.048 64.23 0.97 0.04 0.12

U103 M.N. Vein 175.25 182.50 7.25 3.72 0.015 113.37 4.67 0.05 0.30

U104 hw 171.00 171.75 0.75 0.21 0.080 130.67 1.92 13.22 6.71

U104 hw 187.00 188.00 1.00 0.28 0.014 77.45 1.28 4.16 11.50

U104 M.N. Vein 198.50 202.00 3.50 0.97 0.045 111.33 1.76 2.89 7.23

U105 M.N. Vein 154.00 170.50 16.50 7.39 0.036 141.50 2.29 0.15 7.64

U106 hw 256.00 259.00 3.00 1.04 0.020 241.40 3.28 1.12 3.85

U106 hw 267.50 268.50 1.00 0.35 0.011 36.80 2.15 0.01 0.08

U106 M.N. Vein 304.00 314.50 10.50 3.71 0.057 114.41 4.50 0.23 1.19

U107 M.N. Vein 123.50 135.00 11.50 8.78 0.076 91.54 1.63 2.76 0.86

U108 hw 131.50 132.50 1.00 0.42 0.014 40.90 1.52 0.30 2.05

U108 M.N. Vein 139.50 156.25 16.75 7.07 0.032 114.90 3.20 1.55 0.28

U109 hw 316.25 317.50 1.25 0.23 0.033 96.80 6.07 0.04 0.22

U109 M.N. Vein 358.00 364.00 6.00 1.11 0.009 60.64 2.63 0.03 1.77

U110 M.N. Vein 125.00 133.50 8.50 2.03 0.051 66.00 1.96 1.21 0.85

U111 hw 92.00 93.50 1.50 0.51 0.083 44.00 0.15 17.23 8.97

U111 hw 122.50 125.50 3.00 1.03 0.008 97.67 3.50 0.02 0.21



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

U01 hw 62.80 63.30 0.50 0.19 0.007 36.10 1.13 0.00 1.65

U111 M.N. Vein 135.50 147.00 11.50 3.97 0.010 71.19 2.68 0.01 0.22

U112 hw 231.50 235.50 4.00 2.18 0.054 66.34 0.73 2.42 2.84

U112 hw 268.00 269.25 1.25 0.68 0.021 101.92 3.88 0.03 0.29

U112 hw 291.50 293.50 2.00 1.08 0.020 141.03 8.35 0.03 0.29

U112 hw 296.50 299.00 2.50 1.35 0.010 23.14 1.33 0.01 0.05

U112 hw 307.00 308.00 1.00 0.54 0.011 40.75 1.79 0.06 1.09

U112 M.N. Vein 319.50 323.00 3.50 1.89 0.022 118.91 5.45 0.10 0.62

U113 hw 282.50 284.00 1.50 0.46 0.002 30.10 1.88 0.01 0.06

U113 hw 290.50 292.00 1.50 0.46 0.003 26.37 1.45 0.01 0.04

U113 M.N. Vein 343.25 349.00 5.75 1.78 0.017 78.85 3.95 0.02 1.11

U114 hw 281.50 284.00 2.50 0.88 0.000 33.36 2.02 0.01 0.06

U114 M.N. Vein 291.00 294.00 3.00 1.06 0.000 101.25 4.72 0.03 0.17

Drill holes CG-05-U01, CG-05-U14, CG-05-U15, CG-05-U16, CG-05-U23, CG-05-U24,

CG-05-U28, CG-05-U41, CG-05-U49, CG-05-U57, CG-05-U63, CG-05-U66, CG-05-U89,

CG-05-U90 and CG-05- U92 were anomalous but did not intersect any significant mineralization.

Drill hole CG-05-U65 was abandoned because of poor ground conditions and did not reach the planned

depth to intersect the vein.

Phases IV and V

Surface drilling commenced in October 2006 and was completed in April 2007. The five surface holes

were drilled with PQ-diameter rods to approximately 300 m, reduced to HQ rods to about 700 m and

then reduced again to NQ rods. A total of 4,824.5 m were drilled and 304 samples assayed copper,

silver, lead, zinc and gold at ALS Chemex in Vancouver.

Significant intersections are presented in Tables 11.4 and 11.5. True widths are estimated by correcting

for strike and dip of the vein with regard to the bearing and inclination of the drill hole.

Table 11.4: Results from the Phase IV Surface Drilling Program

HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

S38 M.N. Vein 768.75 773.50 4.75 3.56 0.087 52.08 1.34 0.03 0.06

S39 M.N. Vein 735.00 737.00 2.00 1.18 0.078 37.00 0.76 0.05 0.07

S39 M.N. Vein 743.50 745.50 2.00 1.18 0.540 69.75 0.19 0.21 4.22

S40 M.N. Vein 667.50 668.50 1.00 0.67 0.025 22.50 0.39 0.05 0.08

S41 M.N. Vein 798.50 799.50 1.00 0.73 0.025 219.00 0.01 0.16 0.17

S42 M.N. Vein 976.00 977.00 1.00 0.31 0.025 37.50 1.34 0.01 0.04

S43 M.N. Vein 196.85 202.85 6.00 4.78 1.166 95.71 0.46 1.10 6.13



Drill holes CG-06-39, CG-07-40 and CG-07-42 were anomalous but did not intersect any significant

mineralization.

Underground NQ-diameter drilling commenced in November 2006 and was completed in July 2007.

Sixty nine holes were drilled for a total of 21,441.10 m. Two thousand, two hundred and seventy seven

samples assayed for copper, silver, lead, zinc and gold at ALS Chemex in Vancouver.

Table 11.5: Results from the Phase V Underground Drilling Program

HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

U115 hw 64.00 64.75 0.75 0.50 0.097 182.33 2.08 0.27 18.53

U115 hw 142.00 142.75 0.75 0.48 0.107 38.00 1.30 0.01 0.23

U115 M.N. Vein 148.50 152.00 3.50 2.24 0.241 59.14 0.67 0.17 0.51

U116 M.N. Vein 133.90 135.40 1.50 1.47 0.025 40.67 0.56 0.04 0.20

U117 M.N. Vein 168.00 171.50 3.50 2.82 0.155 5.96 0.01 0.07 0.26

U118 hw 399.45 401.35 1.90 0.80 0.025 30.79 1.60 0.02 0.22

U118 M.N. Vein 413.10 416.30 3.20 1.34 0.063 68.88 1.83 0.03 1.94

U119 M.N. Vein 392.00 395.75 3.75 1.75 0.053 27.67 2.12 0.01 0.46

U120 M.N. Vein 310.50 329.50 19.00 12.75 0.026 54.20 3.62 0.04 0.59

U121 hw 282.20 283.50 1.30 0.66 0.025 72.23 0.57 5.09 3.51

U121 M.N. Vein 317.00 323.00 6.00 3.16 0.025 109.75 3.62 0.32 3.07

U122 hw 374.50 375.50 1.00 0.41 0.025 58.50 1.38 0.18 2.80

U122 M.N. Vein 409.00 417.00 8.00 3.31 0.025 53.93 3.84 0.01 0.13

U123 M.N. Vein 390.15 392.00 1.85 1.11 0.105 69.73 4.08 0.01 0.45

U124 M.N. Vein 304.20 309.50 5.30 3.19 0.039 125.99 3.83 0.22 3.42

U125 M.N. Vein 286.40 293.25 6.85 3.89 0.049 111.95 2.63 0.29 2.08

U126 M.N. Vein 201.00 203.80 2.80 2.27 0.168 32.20 1.18 0.07 0.47

U127 M.N. Vein 249.50 253.00 3.50 2.69 0.123 102.00 6.89 0.04 0.47

U128 M.N. Vein 171.50 172.50 1.00 0.99 0.053 15.00 0.20 0.13 5.08

U129 M.N. Vein 285.50 291.50 6.00 4.50 0.166 89.92 2.26 0.10 3.23

U130 M.N. Vein 333.00 334.00 1.00 0.83 0.025 55.00 1.98 0.02 2.40

U131 M.N. Vein 188.20 192.25 4.05 3.14 0.133 153.17 2.86 0.09 0.52

U132 M.N. Vein 327.25 335.50 8.25 5.04 0.025 30.00 0.90 0.05 4.58

U133 M.N. Vein 319.55 323.70 4.15 3.88 0.025 7.88 0.04 0.15 0.41

U134 M.N. Vein 232.50 248.00 15.50 10.49 0.057 27.88 1.16 0.01 0.31

U135 M.N. Vein 342.40 346.25 3.85 2.80 0.025 10.46 0.05 0.05 1.31

U136 M.N. Vein 346.40 347.00 0.60 0.21 0.025 84.00 4.69 0.03 0.18

U136 fw 385.00 385.40 0.40 0.14 0.025 26.00 0.97 0.07 0.46

U137 M.N. Vein 263.70 266.20 2.50 1.59 0.025 7.90 0.21 0.01 0.15

U138 M.N. Vein 346.00 348.00 2.00 1.16 0.025 10.88 0.81 0.01 0.03

U139 hw 190.00 192.50 2.50 2.25 0.247 52.80 0.54 0.33 4.64

U139 M.N. Vein 219.50 222.25 2.75 2.47 0.058 75.00 3.68 0.02 0.18

U140 stringer 361.00 363.50 2.50 1.53 0.025 2.50 0.06 0.00 0.01

U141 M.N. Vein 400.75 403.00 2.25 0.76 0.025 32.78 2.52 0.01 0.10

U142 M.N. Vein 214.50 218.65 4.15 2.87 0.029 6.93 0.04 0.04 1.23

U143 Other Vein 339.00 343.00 4.00 2.23 0.025 5.34 0.09 0.01 0.01

U144 M.N. Vein 243.25 250.50 7.25 5.29 0.034 47.07 2.94 0.01 0.42



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

U115 hw 64.00 64.75 0.75 0.50 0.097 182.33 2.08 0.27 18.53

U145 M.N. Vein 258.00 259.75 1.75 1.57 0.025 129.29 6.91 0.07 0.31

U146 M.N. Vein 180.50 181.00 0.50 0.44 0.025 30.00 1.37 0.03 0.14

U146 stringer 186.50 187.00 0.50 0.44 0.160 2.50 0.14 0.08 7.07

U147 M.N. Vein 355.50 358.00 2.50 1.57 0.375 520.00 5.64 0.42 3.18

U148 M.N. Vein 397.00 403.00 6.00 2.18 0.032 22.59 1.63 0.01 0.08

U149 hw 204.00 205.50 1.50 1.00 0.025 776.67 3.98 18.81 7.79

U149 M.N. Vein 246.00 249.55 3.55 2.26 0.068 30.21 1.78 0.01 0.51

U150 M.N. Vein 338.50 340.50 2.00 1.78 0.034 70.25 1.49 0.03 0.29

U151 hw 196.00 199.00 3.00 2.01 0.122 70.17 0.90 2.32 3.95

U151 hw 227.35 229.00 1.65 1.10 0.108 77.70 0.89 7.34 3.37

U151 M.N. Vein 232.95 252.00 19.05 12.20 0.061 137.35 4.00 0.79 1.30

U152 M.N. Vein 350.75 361.00 10.25 6.02 0.029 30.44 2.00 0.01 0.18

U153 hw 278.35 280.60 2.25 1.27 0.069 46.40 2.73 0.00 0.08

U153 M.N. Vein 290.25 313.00 22.75 12.84 0.041 57.69 3.55 0.03 0.21

U154 M.N. Vein 159.85 161.75 1.90 1.63 0.025 67.08 0.65 0.16 1.18

U155 M.N. Vein 134.50 138.00 3.50 3.48 0.389 113.86 1.57 0.86 1.74

U156 M.N. Vein 145.80 149.50 3.70 3.36 0.027 88.51 1.98 0.09 1.16

U157 M.N. Vein 364.00 367.50 3.50 2.73 0.094 51.00 1.30 0.12 0.80

U158 M.N. Vein 170.25 174.50 4.25 3.75 0.025 51.53 2.16 0.01 2.70

U159 M.N. Vein 201.50 202.50 1.00 0.81 0.258 65.00 2.45 0.19 1.15

U160 hw 63.00 64.50 1.50 0.41 0.095 18.33 0.27 0.00 13.85

U160 M.N. Vein 363.60 365.30 1.70 0.81 0.025 66.76 2.14 0.02 0.39

U161 M.N. Vein 173.35 176.50 3.15 2.05 0.228 175.94 4.79 0.57 1.65

U162 M.N. Vein 209.65 210.00 0.35 0.24 0.025 43.00 1.62 0.00 0.10

U163 M.N. Vein 349.85 351.00 1.15 0.73 0.025 2.50 0.00 0.01 0.02

U163 fw 359.00 361.00 2.00 1.27 0.025 2.50 0.12 0.01 9.42

U164 M.N. Vein 381.50 382.00 0.50 0.25 0.025 88.00 0.80 0.01 0.03

U165 hw 117.00 119.75 2.75 0.64 0.025 121.18 1.78 0.16 1.17

U165 M.N. Vein 258.50 260.50 2.00 0.81 0.025 32.88 1.19 0.03 0.37

U166 M.N. Vein 173.50 177.00 3.50 2.94 0.025 56.21 2.29 0.01 2.45

U167 M.N. Vein 191.00 203.50 12.50 10.68 0.034 93.24 4.12 0.05 0.48

U168 M.N. Vein 285.65 288.50 2.85 2.14 0.168 19.54 0.16 0.02 2.55

U169 M.N. Vein 267.00 270.50 3.50 1.19 0.000 70.70 2.45 0.08 0.33

U170 M.N. Vein 192.15 196.00 3.85 1.81 0.000 43.48 1.68 0.06 0.50

U171 M.N. Vein 303.00 309.00 6.00 5.40 0.030 55.69 1.62 1.25 3.85

U172 hw 348.50 349.00 0.50 0.36 0.080 23.00 1.51 0.01 0.06

U172 M.N. Vein 364.50 365.00 0.50 0.39 0.025 71.00 2.24 0.10 5.40

U173 M.N. Vein 220.25 224.00 3.75 1.49 0.025 30.07 0.81 0.03 1.81

U174 hw 239.75 240.50 0.75 0.61 0.025 24.67 0.79 0.04 1.23

U174 M.N. Vein 265.10 267.00 1.90 1.54 0.248 153.95 4.68 0.02 0.60

U175 M.N. Vein 323.50 323.70 0.20 0.16 0.120 314.00 0.54 13.60 5.40

U176 M.N. Vein 202.85 212.00 9.15 8.47 0.039 56.87 2.38 0.48 1.04

U177 hw 241.00 242.35 1.35 0.89 0.125 198.56 1.50 5.94 8.93

U177 M.N. Vein 290.70 292.75 2.05 1.37 0.030 42.15 2.23 0.03 0.40

U178 M.N. Vein 195.35 200.70 5.35 4.47 0.160 338.81 6.59 0.44 0.67

U179 M.N. Vein 173.00 177.50 4.50 2.57 0.031 120.56 2.14 0.15 1.00



HOLEID

StructureFrom(m)

To(m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

U115 hw 64.00 64.75 0.75 0.50 0.097 182.33 2.08 0.27 18.53

U179 fw 191.00 193.50 2.50 1.43 0.036 39.00 1.41 0.02 2.47

U179 fw 204.00 204.50 0.50 0.29 0.025 37.00 1.37 0.02 1.09

U180 M.N. Vein 224.35 226.70 2.35 1.13 0.037 131.19 2.47 0.03 0.67

U181 M.N. Vein 301.40 303.00 1.60 1.55 0.025 50.81 0.63 0.49 3.78

U182 M.N. Vein 176.85 179.25 2.40 1.89 0.025 67.44 1.18 0.02 7.66

U183 M.N. Vein 348.95 355.35 6.40 5.49 0.065 67.27 1.50 0.25 1.27

Phase VI

Surface drilling commenced in January 2008 and was completed in October 2008. The 105 holes were

drilled with HQ rods and, where necessary, reduced to NQ rods. A total of 29,642.90 m were drilled

from surface and 4,497 samples assayed for copper, silver, lead, zinc and gold. Samples were

prepared at ALS Chemex in Hermosillo Sonora, Mexico and assayed by ALS Chemex in Vancouver.

Duplicate samples taken for QA/QC were sent to SGS in Toronto.

Underground NQ-diameter drilling commenced in November 2006 and was completed in July 2007.

Sixty nine holes were drilled for a total of 21,441.10 m. Assaying of 2277 samples for copper, silver,

lead, zinc and gold was carried out by ALS Chemex in Vancouver.

Significant intersections are presented in Tables 11.6 and 11.7.

Table 11.6: Results from the Phase VI Surface Drilling Program

HOLEID

StructureFrom(m)

To (m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

S44 M.N. Vein 173.70 180.00 6.30 5.64 0.373 81.08 0.15 0.18 2.43

S44 M.N. Vein 192.15 192.90 0.75 0.66 1.160 113.33 0.07 1.69 3.54

S45 M.N. Vein 253.90 255.65 1.75 1.06 0.234 64.43 0.25 0.26 2.38

S45 M.N. Vein 258.75 266.00 7.25 4.39 0.069 99.10 1.09 0.15 2.68

S46 hw 237.25 238.50 1.25 0.67 0.280 57.80 0.70 0.08 2.46

S46 M.N. Vein 265.60 271.75 6.15 3.27 0.051 67.07 0.70 0.04 3.64

S47 no intersection

S48 hw 350.50 351.65 1.15 0.46 1.102 26.52 0.06 0.46 16.83

S48 M.N. Vein 395.00 396.25 1.25 0.51 1.352 214.60 1.92 0.28 5.06

S49 M.N. Vein 368.00 372.00 4.00 1.50 0.164 20.63 0.12 0.07 0.22

S50 M.N. Vein 242.00 244.50 2.50 2.07 0.260 93.90 1.78 0.04 0.12

S51 M.N. Vein 311.50 314.50 3.00 1.67 0.027 62.67 1.10 0.06 1.31

S52 hw 130.70 131.25 0.55 0.47 5.276 30.91 0.15 1.07 8.56

S52 M.N. Vein 135.50 140.00 4.50 3.83 0.480 12.04 0.05 0.08 1.61

S53 M.N. Vein 226.55 233.00 6.45 4.67 0.253 65.15 0.39 0.37 6.60

S54 M.N. Vein 111.25 114.00 2.75 2.28 0.355 22.91 0.19 0.15 1.61

S55 M.N. Vein 297.40 297.90 0.50 0.35 0.000 16.50 0.11 0.10 6.02

S56 M.N. Vein 307.20 312.75 5.55 4.19 0.187 53.62 0.31 0.24 7.32

S57 M.N. Vein 160.00 163.50 3.50 2.12 0.550 2.14 0.01 0.08 0.50

S58 M.N. Vein 366.25 368.75 2.50 1.72 0.646 126.80 0.48 1.05 5.16



HOLEID

StructureFrom(m)

To (m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

S44 M.N. Vein 173.70 180.00 6.30 5.64 0.373 81.08 0.15 0.18 2.43

S59 M.N. Vein 93.50 95.40 1.90 1.03 0.317 15.79 0.03 0.10 3.26

S60 hw 52.00 54.00 2.00 1.70 0.401 94.88 0.09 0.36 2.19

S60 M.N. Vein 92.00 92.50 0.50 0.43 0.485 39.50 0.17 2.61 10.39

S61 hw 66.00 67.00 1.00 0.87 3.040 28.00 0.02 0.16 1.72

S61 M.N. Vein 162.50 167.00 4.50 3.87 0.237 8.11 0.01 0.30 1.35

S62 M.N. Vein 45.25 53.00 7.75 4.04 0.036 69.19 0.78 0.29 2.47

S63 hw 69.50 72.50 3.00 2.05 16.663 39.17 0.19 0.61 1.02

S63 M.N. Vein 97.00 104.25 7.25 4.96 0.180 9.14 0.01 0.59 2.30

S64 M.N. Vein 62.25 64.50 2.25 1.92 0.117 79.33 0.61 2.36 2.55

S65 M.N. Vein 210.50 211.50 1.00 0.71 0.118 18.75 0.09 0.55 4.50

S66 hw 16.50 17.50 1.00 0.75 0.348 36.25 0.19 0.57 22.96

S66 M.N. Vein 54.00 65.50 11.50 8.65 41.258 26.37 0.16 0.16 5.91

S67 M.N. Vein 260.25 261.75 1.50 1.33 0.087 23.00 0.04 1.76 1.76

S69 hw 60.35 62.50 2.15 1.87 0.537 13.14 0.01 0.74 1.80

S69 M.N. Vein 68.00 93.88 25.88 22.45 0.298 14.89 0.02 0.27 1.29

S70 M.N. Vein 166.75 168.75 2.00 1.76 0.416 8.88 0.02 0.05 1.21

S71 M.N. Vein 188.25 189.25 1.00 0.77 0.270 24.25 0.25 0.03 7.95

S72 hw 264.50 265.50 1.00 0.72 0.428 38.50 0.16 0.33 10.32

S72 M.N. Vein 272.00 277.00 5.00 3.59 0.241 30.90 0.16 0.14 5.08

S73 M.N. Vein 58.00 62.00 4.00 3.55 1.839 36.38 0.06 0.74 13.00

S74 M.N. Vein 227.15 235.75 8.60 6.74 0.439 52.48 0.23 1.13 6.48

S75 M.N. Vein 210.40 216.90 6.50 4.98 1.033 60.31 0.16 1.53 6.22

S76 M.N. Vein 219.50 223.75 4.25 3.54 0.033 56.05 1.36 0.04 1.56

S77 M.N. Vein 178.50 182.25 3.75 3.38 0.588 93.93 0.24 1.58 10.14

S78 hw 242.00 249.00 7.00 6.42 3.943 19.96 0.08 0.10 2.63

S78 M.N. Vein 252.60 254.90 2.30 2.11 0.112 34.17 0.25 0.42 7.51

S79 hw 138.50 139.50 1.00 0.49 0.245 53.50 0.57 2.37 1.43

S79 M.N. Vein 209.50 214.50 5.00 3.80 0.603 139.65 1.57 0.30 0.86

S80 M.N. Vein 210.15 216.00 5.85 4.44 0.515 57.14 0.26 0.89 5.18

S81 M.N. Vein 290.50 304.65 14.15 9.84 0.583 72.18 0.37 1.16 3.69

S82 M.N. Vein 221.50 225.00 3.50 2.48 0.051 73.79 2.09 0.04 0.44

S83 M.N. Vein 277.00 279.80 2.80 2.24 0.095 17.23 0.08 0.18 2.53

S84 M.N. Vein 394.75 399.00 4.25 2.81 0.512 40.35 0.38 1.03 5.43

S85 M.N. Vein 320.40 324.25 3.85 3.03 0.229 35.95 0.18 0.46 6.92

S85 M.N. Vein 336.15 340.50 4.35 3.42 1.309 78.62 0.45 0.76 10.11

S86 M.N. Vein 248.35 257.50 9.15 7.10 0.050 4.39 0.09 0.01 0.78

S87 hw 204.30 205.50 1.20 1.06 2.125 8.08 0.03 0.41 4.85

S87 M.N. Vein 214.90 217.10 2.20 1.95 0.080 2.00 0.01 0.01 0.25

S88 M.N. Vein 237.75 242.00 4.25 3.38 0.475 52.59 0.25 1.63 4.06

S88 fw 245.00 247.00 2.00 1.59 0.170 28.63 0.15 0.33 3.28

S89 hw 135.50 137.50 2.00 1.54 0.263 79.00 0.42 2.24 4.62

S89 M.N. Vein 306.50 314.50 8.00 6.17 0.271 210.44 3.27 0.05 1.23

S90 M.N. Vein 384.00 385.40 1.40 1.02 0.422 109.04 0.39 0.46 5.73

S91 M.N. Vein 297.00 305.50 8.50 5.11 0.244 92.42 1.25 0.08 1.98

S92 M.N. Vein 216.75 220.50 3.75 3.25 0.128 35.20 1.12 0.01 0.29



HOLEID

StructureFrom(m)

To (m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

S44 M.N. Vein 173.70 180.00 6.30 5.64 0.373 81.08 0.15 0.18 2.43

S93 M.N. Vein 300.00 306.20 6.20 3.51 0.745 65.71 0.51 1.41 5.10

S94 M.N. Vein 414.70 419.00 4.30 2.77 0.541 45.65 0.41 0.80 5.32

S95 M.N. Vein 291.50 297.50 6.00 3.87 0.097 116.54 2.18 0.03 1.08

S96 M.N. Vein 376.50 378.25 1.75 0.77 0.144 73.57 0.93 0.08 3.59

S97 M.N. Vein 261.00 275.25 14.25 12.08 0.072 47.44 1.14 0.05 0.30

S98 M.N. Vein 182.00 190.00 8.00 6.06 1.388 157.38 0.37 0.63 5.27

S99 hw 282.00 285.00 3.00 2.35 0.457 19.67 0.06 0.86 6.06

S99 M.N. Vein 408.00 411.50 3.50 2.74 0.115 12.57 0.03 0.02 1.03

S100 M.N. Vein 212.25 224.25 12.00 7.79 0.424 42.85 0.25 0.37 2.79

S101 M.N. Vein 222.00 223.75 1.75 1.46 0.247 40.57 0.46 0.17 1.45

S102 M.N. Vein 284.50 288.00 3.50 1.41 0.349 77.57 0.60 0.10 2.65

S103 M.N. Vein 276.20 279.50 3.30 2.10 0.066 31.79 0.44 0.05 0.60

S104 M.N. Vein 314.50 321.00 6.50 3.13 0.228 70.29 1.49 0.03 1.09

S104 fw 328.50 334.00 5.50 2.65 1.148 14.09 0.03 1.42 3.63

S105 M.N. Vein 313.00 323.00 10.00 6.04 0.384 46.35 0.54 0.10 2.81

S105 fw 340.60 343.00 2.40 1.45 0.600 144.58 1.81 0.26 6.07

S106 M.N. Vein 334.50 351.00 16.50 15.03 0.063 77.02 2.42 0.13 3.69

S107 M.N. Vein 172.00 177.00 5.00 4.21 2.981 44.30 0.14 0.86 5.36

S108 hw 327.50 333.40 5.90 2.50 1.865 7.27 0.01 0.32 4.70

S108 hw 335.75 339.50 3.75 1.59 0.127 6.33 0.02 0.07 2.67

S108 M.N. Vein 352.50 357.50 5.00 2.12 0.129 31.20 0.33 0.05 2.57

S108 M.N. Vein 365.00 380.50 15.50 6.56 0.559 68.61 0.82 0.45 5.46

S109 M.N. Vein 206.25 208.75 2.50 1.60 0.540 42.80 0.27 0.67 2.47

S110 M.N. Vein 341.00 343.00 2.00 1.13 0.050 8.50 0.08 0.01 3.22

S111 M.N. Vein 280.75 283.80 3.05 1.57 0.424 41.74 0.38 2.21 7.76

S112 hw 429.50 431.50 2.00 0.97 0.170 48.25 0.37 0.15 4.69

S112 M.N. Vein 453.00 457.25 4.25 2.05 0.158 27.29 0.09 0.57 3.79

S113 M.N. Vein 334.30 353.50 19.20 10.94 0.459 37.74 0.25 0.71 3.44

S114 M.N. Vein 25.00 49.50 24.50 23.87 0.201 62.21 0.72 0.32 1.67

S115 hw 382.00 386.00 4.00 1.37 0.568 34.38 0.04 1.49 6.82

S115 M.N. Vein 406.50 417.10 10.60 3.64 0.425 20.39 0.24 0.94 2.84

S116 M.N. Vein 49.50 55.20 5.70 4.98 0.720 50.30 0.43 0.22 6.61

S117 M.N. Vein 301.65 310.75 9.10 3.56 0.653 49.73 0.44 0.43 4.26

S118 hw 56.50 59.00 2.50 0.64 1.080 23.60 0.51 0.05 1.13

S118 M.N. Vein 97.50 110.00 12.50 3.22 0.158 8.16 0.07 0.02 3.40

S119 hw 445.50 447.50 2.00 0.81 0.088 21.50 0.09 0.70 2.99

S119 M.N. Vein 480.00 485.50 5.50 2.23 0.265 46.05 0.09 0.69 3.83

S120 M.N. Vein 31.50 33.60 2.10 1.65 0.425 63.81 0.11 0.35 4.55

S121 hw 223.50 226.50 3.00 2.09 0.358 30.42 0.27 0.16 3.69

S121 M.N. Vein 236.20 239.25 3.05 2.13 0.702 48.39 0.33 0.11 3.64

S122 M.N. Vein 54.00 55.75 1.75 1.40 0.714 56.43 0.13 3.17 12.44

S123 M.N. Vein 51.25 58.50 7.25 4.67 1.060 46.67 0.04 2.23 10.17

S124 M.N. Vein 324.90 329.30 4.40 2.56 0.513 88.34 0.37 0.22 4.22

S125 M.N. Vein 36.00 45.50 9.50 8.55 0.230 7.37 0.02 0.06 0.48

S126 M.N. Vein 43.75 52.00 8.25 7.15 0.074 126.97 1.03 0.21 2.99



HOLEID

StructureFrom(m)

To (m)

Length (m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

S44 M.N. Vein 173.70 180.00 6.30 5.64 0.373 81.08 0.15 0.18 2.43

S127 M.N. Vein 185.50 191.50 6.00 4.75 0.640 50.74 0.33 0.49 2.77

S128 M.N. Vein 420.50 425.00 4.50 1.67 0.607 77.22 0.05 0.89 3.01

S129 hw 169.00 171.30 2.30 1.21 0.711 14.74 0.04 0.93 5.96

S129 M.N. Vein 276.00 278.50 2.50 1.31 0.214 30.40 0.24 0.05 1.90

S130 M.N. Vein 283.75 287.00 3.25 2.39 0.102 31.23 0.49 0.06 1.55

S131 hw 203.95 207.50 3.55 2.85 0.331 63.14 0.11 0.38 1.75

S131 M.N. Vein 389.50 399.50 10.00 8.02 0.030 90.05 4.82 0.06 1.60

S132 M.N. Vein 274.00 275.50 1.50 1.33 0.138 93.50 0.27 13.23 1.19

S132 fw 284.00 290.25 6.25 5.54 0.050 61.36 0.52 0.09 0.57

S133 M.N. Vein 357.50 364.25 6.75 3.71 0.044 168.04 8.86 0.03 0.85

S134 M.N. Vein 291.00 299.00 8.00 6.55 0.184 42.97 0.77 0.07 4.00

S135 M.N. Vein 394.00 407.75 13.75 11.52 0.186 95.04 2.93 0.01 0.90

S136 hw 461.50 462.50 1.00 0.65 0.009 90.75 2.00 0.04 0.13

S136 M.N. Vein 480.00 484.25 4.25 2.75 0.010 95.24 2.34 0.48 3.01

S137 ? 172.00 172.25 0.25 0.00 0.216 93.00 0.48 2.07 0.87

S138 M.N. Vein 44.50 48.50 4.00 3.43 0.227 23.56 0.35 0.02 3.98

S139 hw 246.30 248.00 1.70 1.64 0.054 17.21 0.45 0.01 1.42

S139 M.N. Vein 254.75 260.50 5.75 5.54 0.081 4.96 0.03 0.09 0.46

S140 M.N. Vein 469.75 495.25 25.50 15.23 0.093 55.27 2.46 0.12 1.01

S141 M.N. Vein 34.25 39.75 5.50 5.02 0.224 20.55 0.43 0.02 0.57

S142 M.N. Vein 48.00 49.50 1.50 1.31 0.000 25.67 0.70 0.03 0.35

S143 M.N. Vein 44.70 50.75 6.05 5.41 0.483 14.47 0.14 0.04 11.72

S144 M.N. Vein 25.00 30.48 5.48 4.90 Unsampled, no economic mineral

S145 M.N. Vein 42.50 53.74 11.24 9.92 0.021 10.53 0.13 0.02 0.78

S146 M.N. Vein 40.00 44.00 4.00 2.80 0.194 8.31 0.02 0.02 0.43

S147 M.N. Vein 230.25 232.75 2.50 1.63 0.538 45.60 0.11 1.75 2.41

S148 M.N. Vein 299.00 323.25 24.25 15.03 0.616 25.20 0.45 1.82 4.31

S149 M.N. Vein 133.00 137.33 4.33 3.78 0.693 38.12 0.10 0.88 4.47

S150 M.N. Vein 155.40 166.80 11.40 8.34 0.922 43.40 0.11 0.85 3.73

S150 M.N. Vein

fw1 249.00 251.44 2.44 1.78 6.510 318.67 0.05 0.27 2.01

S150 M.N. Vein

fw2 323.75 328.10 4.35 3.18 0.695 36.53 0.82 0.04 1.78



Table 11.7: Results from the Phase VI Underground Drilling Program

HOLEID

StructureFrom(m)

To (m) Length

(m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

U184 M.N. Vein 288.00 289.00 1.00 0.94 0.740 54.50 0.88 0.02 0.35

U185 hw 196.50 197.50 1.00 0.88 0.305 69.75 0.93 0.08 3.07

U185 hw 210.50 215.75 5.25 4.08 0.010 15.38 0.26 0.25 2.97

U185 M.N. Vein 230.25 235.75 5.50 3.65 0.010 4.68 0.20 0.47 2.66

U186 M.N. Vein 154.30 158.75 4.45 3.97 0.010 83.38 1.78 0.09 1.20

U187 M.N. Vein 235.75 241.00 5.25 2.74 0.224 91.44 2.76 0.35 2.33

U188 M.N. Vein 300.50 317.75 17.25 16.89 0.035 3.01 0.04 0.02 0.11

U189 hw 107.50 109.30 1.80 1.80 0.000 19.44 0.25 0.11 5.37

U189 M.N. Vein 300.50 300.75 0.25 0.25 0.000 53.60 0.61 0.46 1.27

U190 M.N. Vein 318.00 319.00 1.00 1.00 0.010 93.50 2.28 0.03 0.30

U191 hw 312.50 314.00 1.50 1.25 0.000 5.73 0.07 0.71 4.80

U191 M.N. Vein 342.50 355.00 12.50 11.04 0.039 60.34 2.45 0.10 0.99

U192 M.N. Vein 359.00 369.00 10.00 7.94 0.074 134.93 2.67 0.13 1.14

U192 fw 376.00 377.50 1.50 1.19 0.000 41.20 1.01 0.02 0.03

U192 fw 384.00 385.00 1.00 0.79 0.010 261.30 4.75 0.02 0.09

U193 M.N. Vein 407.75 408.00 0.25 0.17 0.000 48.00 1.20 0.01 0.05

U194 M.N. Vein 326.00 332.00 6.00 5.58 0.073 126.00 1.86 0.23 1.12

U194 fw 344.20 345.00 0.80 0.74 0.175 160.63 0.89 0.02 0.11

U195 M.N. Vein 272.75 297.75 25.00 6.69 0.034 57.89 2.19 0.52 0.94

U196 M.N. Vein 362.50 365.25 2.75 1.62 0.000 9.20 0.26 0.03 0.01

U197 hw 357.50 358.50 1.00 0.26 0.010 12.00 0.78 0.02 1.25

U197 M.N. Vein 426.25 427.50 1.25 0.33 0.010 51.40 2.32 0.01 0.08

U198 hw 301.00 301.50 0.50 0.16 0.010 125.00 3.51 0.06 0.38

U198 hw 406.50 420.00 13.50 4.37 0.010 26.87 1.59 0.01 0.54

U198 M.N. Vein 431.25 440.75 9.50 3.08 0.010 60.84 2.72 0.05 1.24

U199 M.N. Vein 322.50 331.50 9.00 5.03 0.010 28.33 2.01 0.01 0.09

U200 M.N. Vein 274.00 308.75 34.75 21.68 0.056 41.83 1.82 0.11 3.31

U200 M.N. Vein 315.80 321.40 5.60 3.49 0.116 54.34 2.00 0.01 0.10

U201 M.N. Vein 299.45 312.00 12.55 9.51 0.050 11.46 0.31 0.03 6.38

U202 hw 315.25 324.50 9.25 4.24 0.050 20.24 1.02 0.06 0.86

U202 M.N. Vein 330.25 366.45 36.20 16.59 0.068 74.47 1.75 1.76 4.75

U203 M.N. Vein 406.25 412.50 6.25 3.39 0.176 38.28 1.82 0.03 0.59

U204 M.N. Vein 373.50 388.25 14.75 7.68 0.089 69.95 2.93 0.11 2.86

U205 stringer 356.00 363.20 7.20 1.42 0.054 39.31 1.19 0.08 1.67

U205 stringer 401.75 418.50 16.75 4.91 0.052 35.67 1.33 0.04 0.11

U205 M.N. Vein 446.00 453.50 7.50 2.46 0.106 66.43 1.95 0.12 4.29

U206 hw 364.00 364.50 0.50 0.31 0.050 29.00 1.64 0.04 0.13

U206 M.N. Vein 378.15 381.70 3.55 2.20 0.053 18.30 1.21 0.01 0.06

U207 M.N vein 446.25 457.00 10.75 6.12 0.050 29.33 1.81 0.01 0.57

U207 fw 463.25 465.00 1.75 1.00 0.050 70.57 4.41 0.04 0.26

U208 M.N vein 392.50 404.75 12.25 5.83 0.053 64.84 3.64 0.08 0.19

U208 fw 415.25 416.85 1.60 1.47 0.050 14.75 0.71 0.03 3.61

U210 M.N. Vein 278.00 283.90 5.90 5.75 0.084 140.62 1.66 1.19 6.96

U211 M.N. Vein 347.80 351.50 3.70 2.77 0.069 50.49 1.59 0.02 2.25

U212 M.N. Vein 313.00 329.75 16.75 15.87 0.072 76.76 1.58 0.05 0.58

U213 M.N. Vein 310.50 319.25 8.75 7.22 0.050 26.26 0.50 0.48 1.21

U214 M.N. Vein 244.50 267.75 23.25 19.20 0.045 62.48 1.86 0.70 1.49



HOLEID

StructureFrom(m)

To (m) Length

(m)

True Width

(m)

Au (g/t)

Ag (g/t)

Cu(%)

Pb(%)

Zn(%)

U184 M.N. Vein 288.00 289.00 1.00 0.94 0.740 54.50 0.88 0.02 0.35

U214 stringer 270.25 272.00 1.75 1.44 0.222 41.43 1.33 0.16 0.09

U215 M.N. Vein 434.50 451.50 17.00 11.24 0.052 29.03 1.44 0.03 1.28

U216 M.N. Vein 514.50 516.00 1.50 0.99 0.050 18.33 1.00 0.01 0.12

U217 M.N. Vein 538.50 542.00 3.50 2.31 0.050 27.58 1.14 0.02 2.65

U209 was cancelled at 36 metres, November 18, 2008.

In 2008, a total of about 39,430 m of diamond drilling was completed thereby triggering this

Technical Report.

The total for the property is 105,261 m with 6,583 samples in 366 drill holes.

Figure 11.1 illustrates the company’s mineral resources in the San Roberto Zone (red block) and the

under explored strike extensions of the Mala Noche vein.



12 Sampling Method and Approach

The results of two sampling methods are presented in this report: drill core cutting and underground

chip sampling.

12.1 Diamond Drill Core Sampling

The following are relevant excerpts from Capstone’s work procedure for drill core handling and

sampling. Many of these procedures were observed during the author’s site visits. Capstone

personnel were found to conduct this work in a professional manner that follows accepted industry

standards.

12.1.1 Drill Site Control

Core boxes are delivered to the drill site in such a manner that they are clean and free of grease

at the site. The drill hole number and box numbers are marked clearly on each box by the driller.

The driller places a wood block in the core box at the end of each core interval recovered from

the hole that clearly shows the distance down hole in feet and metres (converted by the drillers).

The drill site geologist checks the order of the core and makes a very quick log of the rock in the

core box prior to transportation.

A clean top is placed on each box and is tightly sealed by a minimum of two heavy rubber straps

made from tire inner tubes (or equivalent) prior to transportation from the drill site.

Capstone employees transport the core from the drill site to the core shack. Transportation is as

gentle as reasonably possible.

12.1.2 Core Shack Control

When the core arrives at the core shack, a geologist oversees the placement of the core boxes on

the logging tables and the removal of the tops.

The core is inspected and cleaned if required. The boxes are clearly marked for hole number and

start and finish depths prior to photographing and logging for rock quality.

The core are initially photographed and then logged for rock quality, lithology, structure,

alteration and mineralization prior to marking out the sample intervals.

Once the sample intervals have been marked on the boxes, the saw line is marked by the

geologist. The saw line for primary samples is marked in the centre of the core with each side

being roughly equivalent for apparent grade.

The drill hole number and sample interval are clearly entered in the sample book. One ticket

stub is stapled in the corresponding interval in the core box by the geologist and the other two

ticket stubs are placed in the sample bag by the sampler. The sample books are archived at

Cozamin.



The sample interval does not exceed 0.5 m in the vein and 2 m in the wallrock. Very high grade

intervals are marked out and sampled separately from lower grade zones. Sample boundaries are

based on mineral proportions and/or texture (e.g. massive versus disseminated). However,

sample intervals are not less than 0.25 m in length.

The sampler must be very careful to exactly cut the sample as indicated by the geologist. Only a

single piece of core should be removed from the core box at a time. Care should be taken to

replace the unsampled portion of core back in the box in the correct orientation.

12.1.3 Survey control

The locations and orientations of the drill holes are checked by the surveyor prior to drilling and

after the completion of each hole. The driller identifies each hole with a wood plug showing the

drill hole number clearly labelled with permanent black marker. Drill hole locations are be

surveyed using the total stations TOPCON surveying instrument.

Down hole orientation surveys are undertaken on completion of each hole. A minimum of two

survey points are collected on shorter holes. On longer holes, survey points are taken

approximately every 75 m starting at the bottom of the hole. Drill hole surveys are completed

using the Reflex EZ-shot instrument. The survey results are reviewed by the geologist upon

entry into the database and any suspect readings, due to the presence of magmatic minerals are

rejected from use.

12.2 Underground Chip Sampling

The following are relevant excerpts from Capstone’s work procedure for underground channel

sampling:

Mark Out and Survey Control

Geologist marks out the vein and the mineralized portion of the vein with spray paint on the

mining face or back to be sampled. Sample lines and intervals are marked by the geologist with

a 4 m line spacing and appropriate sample intervals. Intervals length can be variable and

represent the geology (e.g. variations in sulphide and gangue minerals, proportions, textures and

cross cutting structures).

The mine surveyor picks up the start location of each channel sample using the total station

(TOPCON) survey instrument.

Each channel sample is entered into the database as a separate “hole” with collar location, total

length and bearing direction. All bearings are determined by turning angles from established

survey points.



Sampling Control

Chip samples up to 20 cm wide are collected along the marked sample line. The line number

and sample interval are clearly entered in the sample book. Two stubs are placed in the sample

bag by the sampler. The sample books are archived at Cozamin.

The geologist checks the sample bags and tag books for consistency before releasing the samples

to the laboratory for analysis.



13 Sample Preparation, Analyses and Security

13.1 Sampling Personnel

Capstone employees are responsible for the all on-site sampling of drill core. Analysis of these

samples is done at accredited outside laboratories. Channel samples are prepared by Capstone

employees for analysis at the on-site laboratory. Duplicate quality control samples (coarse crush and

pulp) are prepared by Capstone employees for analysis at an off-site laboratory. Blind samples

comprised of standard reference material are included in the sample streams.

13.2 Drill Core Sample Preparation and Analytical Procedures

Diamond drill core samples are prepared as outlined in Section 12.1. Additional security measures

undertaken include:

Only Capstone employees are allowed in the core shack when unsampled core is laid out waiting

to be cut.

No person other than the geologist responsible for logging is allowed to handle the core prior to

sampling. The geologist takes great care to ensure that core is returned to the box in the same

position and orientation from which it came.

Visitors to the core shack must be accompanied by a Capstone employee.

A minimum of ten consecutive samples are placed in order in a large sack. The sack is sealed

with tape and by a numbered seal that prevents opening the sack without damaging the seal. The

sample number series of the enclosed samples are clearly written on the exterior of the sack. The

batch number, the serial numbers of the seals and the corresponding sample number series are

written on the transmittal form to be sent to the preparation laboratory.

Capstone drill hole and channel samples have been analyzed at the laboratories as shown in

Tables 13.1to 13.3.

Table 13.1: Primary and check laboratories used for Cozamin drill samples

Principal Laboratory Check Laboratory From To Phase

BSi Inspectorate ALS Chemex April 2004 August 2004 I

ALS Chemex BSI Inspectorate Sept 2004 March 2005 II

SGS ALS Chemex April 2005 April 2006 III

ALS Chemex SGS Sept 2006 Present IV and V

In 2004, Phase I drill core samples were prepared at BSi Inspectorate (BSi) in Durango, Mexico and

analyzed at their laboratory in Nevada. BSi is registered ISO 9002 compliant, certificate 37925.

Check samples were analyzed at ALS Chemex in Vancouver, Canada. ALS Chemex has

ISO registration in Canada (ISO 9001:2001 and ISO 17025).



Phase II drill core samples were prepared in Hermosillo by ALS Chemex and shipped to Vancouver,

Canada, for analysis. Check samples were analyzed at BSi Inspectorate in Sparks, Nevada.

Phase III drill core samples were prepared at SGS in Durango, Mexico. The pulps were shipped

directly to Canada for analysis by SGS Toronto. SGS is ISO 9002 registered and ISO 17025

accredited for Specific Tests, SCC No. 456. Check samples were sent to ALS Chemex in Vancouver

for analysis.

Phase IV and V drill core samples were sent to ALS Chemex in Hermosillo, Mexico, for preparation.

The pulps are shipped directly to Canada for analysis by ALS Chemex in Vancouver. Duplicate

samples were sent to SGS Toronto for check analysis.

Table 13.2: Laboratory methods and elements routinely analyzed

Lab Au Ag Cu Pb Zn

BSI Inspectorate

fire assay 2AT, atomic

absorption finish

aqua regia digest, atomic absorption

finish

aqua regia digest, atomic

absorption finish


absorption finish


absorption finish

ALS Chemex fire assay, gravimetric

finish

fire assay, gravimetric finish

and four acid digest, ICP-AES

finish

four acid digest, ICP-AES finish



SGS

multi-acid digest, atomic

absorption finish

multi-acid digest, atomic absorption

finish

four acid digest, ICP-OES finish



Table 13.3: Analytical methods used when re-analyzing samples with over limit results

Lab Ag Over Limit Cu Over Limit Pb Over Limit Zn Over Limit

BSiInspectorate


finish


finish


finish

ALS Chemex four acid digest,

titration

SGSfire assay, atomic absorption finish

four acid digest, sodium peroxide ICP-OES finish

four acid digest, sodium peroxide ICP-

OES finish

four acid digest, sodium peroxide ICP-OES finish

Current Preparation Method

Samples are sent to ALS Chemex in Hermosillo for preparation. Upon receipt, samples are

inspected for any irregularities. Samples are then dried, weighed, crushed. Two hundred and fifty

grams is split and pulverized to at least 85% passing 75 microns. Reject material is retained at ALS

Chemex in Hermosillo in a cold storage facility. Prepared pulps are sent to ALS Chemex in

Vancouver for primary analysis. Check sample pulps are sent to SGS in Toronto for analysis.



Current Analytical Methods

At ALS Chemex, gold and silver were analyzed by fire assay with a gravimetric finish using a 50 g

charge. The detection range for this method is 0.05 ppm to 1,000 ppm Au and 5 ppm to10,000 ppm

Ag.

Silver was also analyzed with copper, lead and zinc using a four-acid digest by inductively coupled

plasma – atomic emission spectroscopy (ICP-AES). The detection ranges with this method are:

1 ppm to 1,500 ppm Ag and 0.001 ppm to 10,000 ppm for Cu, Pb and Zn.

Samples with over limit lead results are re-analyzed using the CON02 method in which the sample

undergoes a four acid digest producing a lead sulphate that undergoes titration for determination of

the lead content. Two samples from Phase V had over limit results (23 % to 27 % lead). At their

lead values, the tolerance level for reporting the grade with the titration method is ±2.5 %.

At SGS, gold is analyzed by fire assay with an atomic absorption finish using a 30g charge. The

detection range for this method was 5 ppb to 2,000 ppb.

Silver was analyzed from a 2 g charge using a multi-acid digest with atomic absorption finish (0.3 g/t

to 300 g/t Ag detection range). Over limit results were re-analyzed by fire assay with an atomic

absorption finish using a 50 g charge.

Copper, lead and zinc are analyzed by inductively coupled plasma – optical emission spectroscopy

(ICP-OES) using a four acid digest. Detection limits are: 10 ppm to 10 % for copper, 20 ppm to

10 % for lead and 10 ppm to 10 % for zinc. Over limit results are reanalyzed using the same method

but with a sodium peroxide fusion. The over limit detection limit is 0.01 % for each metal.

Quality Assurance and Quality Control

Blanks, standards and pulp duplicates were inserted into the series of underground drill core samples

submitted for assay. Typically, standard and blank samples were placed at the start and finish of the

sampled interval within a hole. Approximately two sample intervals per hole were selected to have

pulp duplicates prepared, and another two intervals per hole were selected for preparation of core

duplicates. Additional quality control samples were inserted into the sequence as deemed necessary,

e.g. a blank inserted in the sample sequence after a sample expected to have very high grade to

monitor the quality of the assays. Prior to the Phase IV and V drilling programs, 6,561 underground

core samples, 229 core duplicates, 225 pulp duplicates, 231 blanks and 226 standards were submitted

for assay. Overall, approximately one in eight samples submitted for assay were used for quality

control. In addition, 574 pulp samples were selected by the lab as analytical checks.



The same quality assurance and control procedure was used in the Phase IV and V drilling programs.

Associated with the drill holes completed and reported herein for the Phase IV and V drilling

programs: 2,277 underground core and 150 surface core samples, 106 core duplicates, 106 pulp

duplicates, 112 blanks and 103 standards were submitted for assay. Overall, approximately one in

six samples submitted for assay were used for quality control.

In addition, 54 pulp samples were selected by the lab as sample checks for gold and silver and 59

samples for the base metals. Nine hundred and thirty ALS Chemex laboratory standard analyses and

502 blanks were completed, in addition to 11 repeat gold and 542 repeat silver assays. The quality

assurance/quality control (QA/QC) of these samples is presented in Section 16.

Although the author has not visited the assay labs used to analyze Cozamin samples, they are

reputable facilities which have been monitored using an appropriate QAQC program. In the opinion

of the author, the sample preparation, analysis and security practises follow accepted industry

standards and the results demonstrate an acceptable level of analytical accuracy and precision.

13.3 Underground Channel Sample Preparation and Analytical Procedures

The underground channel samples were analyzed at both SGS Toronto (using the same methods as

the drill core samples) and at the on-site lab at Cozamin prior to mid-2006 (Table 13.4). SGS

Toronto was used as the primary laboratory and the site laboratory as a check. Pulp samples were

analyzed on-site by fire assay with an atomic absorption finish for copper, silver, lead, zinc and iron.

From mid-2006 the Cozamin site laboratory has been used as the primary laboratory and check

samples sent to SGS in 2006 and ALS Chemex in 2007. The same methods described in section

15.2 for the drill hole samples have been used for the underground check samples submitted to SGS

and ALS Chemex.

Table 13.4: Primary and check laboratories used for Cozamin channel samples

Principal Laboratory Check Laboratory From To Level

SGS Cozamin site lab April 2005 April 2006 7, 8, 9

Cozamin site lab SGS Sept 2006 December 2006 8, 9, 9.3, 9.6, 10, 10.3

Cozamin site lab ALS Chemex January 2007 Present 8, 9, 9.3, 9.6, 10, 10.3

Quality Assurance and Quality Control

Blanks, standards and pulp duplicates were inserted into the series of underground samples

submitted for assay. Standard and blank samples are inserted into the sample sequence

approximately 1 in 15 samples, and pulp duplicates every 20 samples. Additional quality control

samples were inserted into the sequence as deemed necessary. In the opinion of the author, an

acceptable number of standards, blanks and duplicates were submitted and the results demonstrate an

acceptable level of analytical accuracy and precision.



14 Data Verification

14.1 Database Validation

A review of the database was conducted in order to verify the integrity of the contained data. Of the

365 drill holes used in the resource estimation, 20 holes were randomly selected for manual

validation. The contained information, including collar locations, down-hole survey data, geology

codes and assay values were verified back to the original source. The collar location and directional

data was traced back to the original survey sheets. The geology data was traced back to the original

drill logs and the assay data was compared to the original assay certificates.

In addition to the data described above, 50 sample intervals were selected which represent the

highest grade copper samples present in the database. The assay results of these samples were also

compared to the original assay certificates.

There were no errors found in the drill hole collar locations and survey data in the 20 holes randomly

selected for validation. However, while developing the resource model for the Cozamin deposit,

there were instances found where discrepancies exist between the drill hole results and the location

of the mineralized zone in the underground development. As a result, an extensive series of check

surveys were done throughout the mine which identified a series of areas where errors had

previously been made. “Closed loop” survey checks now show that the underground development in

the mine is in the correct locations in the mine’s digital database.

As stated previously in this report, a series of drill holes show mineralized intervals which do not

correlate with the underground (mapped) information and, as a result, the information from these

holes has been excluded from the model. It appears that there are errors in the location of these holes

and further investigation is required before the information from these holes can be utilized. In some

cases, holes may have to be re-drilled in order to properly confirm the results.

Thirty of the early Capstone surface drill holes were initially tested for silver content using

gravimetric method. The lab method was then changed to what is considered a more appropriate

method consisting of a 4-acid digestion with AA finish. Samples from the initial thirty holes were

reanalyzed using 4-acid AA and the certificates were reissued to reflect this. However, some of the

silver grades in the database have not been changed from gravimetric to AA. This is not considered

to be a material error as the gravimetric results tend to be lower than the AA method.

Other than the discrepancy in silver grades described above, there was only one assay value error

noted in the database.



The data in the spreadsheets provided by Capstone contains copper assay results at both two and

three decimal place precision. Some of the sample values have been truncated to two decimal

accuracy (verses rounding of the data). All available copper assay data has been rounded to two

decimal places accuracy prior to loading into the MineSight® system.

The results the data verification indicate that the database is sound and reliable for the purposes of resource estimation.

14.2 Site Visit Validation

Robert Sim, QP, visited the Cozamin mine property on September 3-4, 2008 and again from March

24-26, 2009. The 2008 site visit included an underground tour which included visual inspection of

exposures of the mineralized zone in numerous levels in the mine (Level 8 through Level 11).

Drilling activities, in hole U-114, were observed at a drill station located in the hangingwall of the

deposit on L8. The equipment and practices followed by drilling personnel meet accepted industry

standards.

Underground channel sampling practices were also observed during the underground tour. The

sampling crew exhibited tenacity and skill in the collection of these samples.

During both site visits, a series of randomly selected drill hole intervals were reviewed and, in all

cases, the type and content of copper, zinc and lead minerals observed support the assay results

present in the database.

The core logging, sampling and core storage facilities at the Cozamin mine were inspected and found

to utilize appropriate equipment and all facilities to be clean and organized. The drill core handling

and sampling procedures following on the property were also observed and discussed during site

visits. These practices follow accepted industry standards.

14.3 Conclusions

Observations during the site visit confirm the physical presence of the drilling activities completed

on the deposit and the sampling procedures have been followed according to accepted industry

standards. Observations of the contained mineralogy in the rocks support the assay results and these,

as described in Section 13, have been monitored through an appropriate QAQC program.

The results of the data verification indicate that the database is sound and reliable for the purposes of

resource estimation.



15 Adjacent Properties

Information concerning adjacent properties has not been included in this report as they are not

pertinent.



16 Mineral Processing and Metallurgical Testing

A site visit to Capstone’s Cozamin project was conducted by Jeff Woods, QP, during mid-February

of 2009. The majority of the data used to compile this section was collected from site personnel

during the site visit. A complete review of the processing facilities was completed including the

processing circuit, support equipment, record keeping, assay facilities and historical data. SRK

found the staff to be capable and receptive of suggestions made during the visit.

The current processing facilities and equipment, for the most part, are in good repair, maintained and

functioning at design capacities. Exceptions relate to equipment purchased used and placed into

operation.

The laboratory, maintenance and storage facilities are functional and suitable for use.

Bacis, the most recent operating company at Cozamin, operated a flotation concentrator on site from

1998 to 1999 with an approximate 750 tonne per day production rate. The unit operations included

in the Bacis process are very similar to the current operations, using crushing, grinding flotation,

thickening and filtration to produce individual copper, lead and zinc concentrates. The majority of

the silver reports to the lead and zinc concentrates.

Production at the Cozamin concentrator began in August of 2006 at a nominal production design rate

of 1,000 tonnes per day. In July of 2007, nominal production was expanded to 2,200 tonnes per day

of ore. Additional plant modifications and refinements have been completed or are under way to

improve production throughput, metal recoveries and concentrate grades. These upgrades include

the installation of flash column flotation to improve lead and zinc production, the addition of

concentrate re-grind and magnetic separation to remove iron and improve concentrate grades.

SRK considers the flow sheet to be robust and suitable for purpose.

16.1 Metallurgical Testing

16.1.1 Process Mineralogy

Primary minerals of interest are chalcopyrite, sphalerite, galena and argentite. Deleterious species

include pyrite, pyrrhotite and arsenopyrite. Gangue minerals are primarily, silica, calcite, chlorite,

epidote and minor sericite It has been determined that the phase of pyrrhotite changes as the deposit

gets deeper. This chemical change also results in a change in “metallurgical” response of the

pyrrhotite in the concentrator circuit. Flotation operating parameters as well as adjustments to

reagent type and dosage will be required as operations continue and different pyrrhotite species are

encountered.



16.1.2 Metallurgical Testing

Pre-production metallurgical was completed by SGS and RDI under the direction of Capstone.

Testing confirmed the suitability of the previous flow sheet used by Bacis and was used to develop

the new concentrator’s design criteria.

On-going metallurgical testing is currently being conducted on site at Cozamin’s metallurgical

laboratory. This test work is used to fine-tune the current process and to add in process flow sheet

optimization, reagent selection and dosage. The test work methods and analytical procedures used

are common to the industry and suitable for its use.

Several process improvements have been developed in the Cozamin metallurgical laboratory and

implemented in production. These include;

Determination of concentrate regrind requirements;

Alternative flotation reagents;

Optimization of process pH;

Potential for the removal of undesirable iron content to upgrade the zinc concentrate, among

others.

Owing to the continual changes in process mineralogy as underground operations are developed,

SRK strongly recommends that the metallurgical laboratory be expanded to increase the quantity and

size of bench test work, i.e. flotation.

16.2 Process Description and Flow Sheets

16.2.1 Crushing and Screening

The crushing process flow sheet is illustrated in Figure 16.1. Ore is presently trucked from the

headframe bin and underground ramps to a surface stockpile to allow blending to produce a

consistent copper feed grade. The surface stockpile of approximately 10,000 tonnes is reclaimed by a

front end loader which feeds material to a 100 tonne bin. Ore reports to the 20” x 36” (0.5 m x 0.9 m)

primary jaw crusher via belt feeder. Crusher product is conveyed to the secondary 1.52 m x 3.66 m

vibrating screen ahead of the 4’ (1.22 m) secondary standard head cone crusher. Screen oversize is

fed to the secondary crusher with screen undersize combined with secondary screen undersize

product. This material is conveyed to a 1.83 m x 4.88 m vibrating screen with oversize material

conveyed to the tertiary crusher (1560 Omni cone crusher) and undersize material being conveyed to

the two mill feed fine ore bins. Tertiary crusher product is returned to the 1.83 m x 4.88 m screen.

Two 1,100 tonne capacity fine ore bins are available each feeding one of the two grinding lines in the

milling circuit. Each bin provides approximately 20 hours storage for each grinding line.

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16.2.2 Milling

The current milling process flow sheet is presented in Figure 16.2. The milling section is composed

of two primary ball mills operating in parallel. Each mill is 3.65 m in diameter by 4.27 m long. The

original design feed rate to each mill was a nominally 50 tonnes per hour, but has been increased to

approximately 75 tonnes per hour. Grinding product size is an 80% passing (P80) 100 mesh. Each

ball mill is operated in closed circuit with a cyclone pack composed of 0.66 m diameter cyclones.

Cyclone under flow reports back to the respective grinding mill with the cyclone overflow from both

circuits reporting to a common flotation conditioning tank. Lime is added to the grinding circuit for

pH control throughout the circuit. Flotation reagents currently zinc sulfate. Lime and the collector

S-7583 are added to the grinding circuit as well.

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16.2.3 Flotation

The original process flow sheet has been expanded to include a flash flotation cell for the recovery of

copper and lead. Figure 16.3 illustrates the current flotation flow sheet at Cozamin. Ground slurry

from the grinding circuit reports to the flash flotation cell for initial copper and lead flash flotation.

Concentrate from flash flotation report directly to the copper and lead separation flotation. Tailings

from flash flotation report by gravity to banks of rougher and scavenger flotation cells (6-OK 16

cells) for additional recovery of copper and lead. The copper-lead rougher concentrates report to a

two stage cleaning system. The original second stage cleaner cells have been replaced with a column

cleaner which has improved the overall concentrate grade.

Copper-Lead rougher flotation tailings report to the zinc conditioner tank prior to zinc rougher

flotation where reagents are added to depress deleterious minerals and activate the zinc

mineralization. The zinc rougher concentrate reports to a closed circuit regrind and additional

liberation of zinc mineralization. Product from the regrind circuit reports to two stages of zinc

concentrate cleaning. A column cell has been added to the circuit to improve zinc concentrate grade.

Tailings from the first cleaner stage report to final tails.

Individual copper and lead concentrates are produced from the copper-lead cleaner concentrate via

selective flotation. Reagents are added to promote lead mineral flotation and suppress the flotation

of copper mineralization. The copper-lead flotation rougher tails (copper concentrate) reports

directly to the copper concentrate thickener. The lead concentrate undergoes two stages of cleaning

before being transferred to the lead concentrate thickener.

F-0

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16.2.4 Concentrate Dewatering and Filtration

Copper concentrate is pumped from the 16m diameter concentrate thickener and dewatered using a 2

m x 4 m disc filter (see Figure 16.4). Product moisture is approximately 10%. Copper concentrate

can be stored in the inside bins (capacity 1,500 tonnes) or outside on a concrete pad (capacity

4,000 tonnes). Concentrate is trucked to port (approximately 600 kilometers) and sampled as the

material is transferred to the port warehouse and becomes the property of the buyer.

Zinc concentrate is pumped from the 8 m diameter thickener to the 1.3 m diameter x 4 m disc filter.

Product moisture is approximately 10% and is stored in the inside bins with a capacity of

1,000 tonnes. The material is then transported to the port and sampled the same as the copper

concentrate.

Lead concentrate is pumped from a 4 m diameter thickener to a 1.3 m diameter x 2 m long drum

filter. The final moisture is approximately 8% and this material is stored inside (capacity 400 tonnes)

prior to shipment by truck to the port. All concentrate trucking is done by third party. All trucks are

weighed both empty and full at the mine site and the port.

F-0

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Zin

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16.2.5 Tailings Handling

Tailings from the Cozamin concentrator is first neutralized to remove cyanide and then pumped to

the tailings pond area (See Figure 16.5). A hydrocyclones is used to effect a partial sand slimes

separation, with the sands being used to build up the tailings dam. The slimes portion reports to the

pond where the solids are allowed to settle and clear supernatant collected and recycled back to the

pond.

Review of the tailings pond design and operations were not included in the scope of SRK’s

involvement in the project.

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16.3 Production

Production at the Cozamin concentrator began in August of 2006. Initial production rates averaged

31 K tonnes per month for first three months of operation. For the last quarter of 2008 production

rate averaged 74 K tonnes per month with metal recoveries averaging, 90% copper, 71% silver,

66.5% lead and 55% zinc. The last quarter 2008 metal recoveries compare favourably to those

reported by the Bacis’ operation c.1999 of 80% copper, 70% lead and 65% zinc.

16.3.1 Production Data Collection and Compilation

Process production data is compiled daily by shifts. Data is entered into spread sheets for analysis

and reporting. Tonnages are determined by weightometers; both in the crushing circuit and mill feed

conveyors. The weightometers are inspected and calibrated weekly. The calibration data is recorded

and tracked to determine problems with the equipment.

The Cozamin concentrator utilizes an Outo Kumpu Courier 5 XRF analyzer for determine of mineral

concentrations is several of the process streams. The unit operates in conjunction with automatic

samples and a multiplexer allowing on-line, real time analyses. Samples are collected daily by shift

from the automatic samples are submitted to the laboratory for analyses. Analytical data is used for

mass balance calculations and analyzer calibration.

The current use of spreadsheets for data logging and tracking is a bit cumbersome owing to the

segregation of data by monthly periods. SRK recommends that production, analytical and

mechanical data be imported into a database format. This will allow for quick trans-period data

analyses and trending as well as normal production reporting.

16.3.2 Production Data

Pertinent concentrator data for the 2007 and 2008 are included in Table 16.1 and graphically

presented in Figures 16.6, 16.7 and 16.8. As can be seen from examination of the operating data,

production rates have increased owing to the expansion completed in 2007 and modifications

completed to date. Metal recoveries have generally increased, while maintaining or improving

concentrate grades.

Primary improvements to the flowsheet and operation conditions post expansion are:

May 2008: Addition of the regrind circuit to the zinc cleaner circuit;

September 2008; Inclusion of column flotation in the zinc cleaner circuit;

September 2008; Addition of column flotation to the Cu-Pb circuit; and

September 2008; Change of flotation collector.



Table 16.1: Actual Monthly Concentrate Grade and Flotation Recovery

Month Tonnage Concentrate Grade (%) Recovery (%)

Cu Pb Zn Ag Cu Pb Zn

J-07 33,794 23.6 67.4 41.9 65.7 83.8 51.8 46.6

F-07 29,684 24.8 54.2 39.1 65.8 84.4 53.4 51.2

M-07 36,680 22.5 55.8 42.7 68.3 87.0 46.7 39.2

A-07 35,703 22.0 64.3 47.8 68.5 86.0 55.5 47.2

M-07 39,894 20.8 69.3 46.5 71.5 85.9 53.3 45.2

J-07 44,246 22.7 68.6 39.2 66.0 84.8 55.6 44.7

J-07 59,262 22.3 67.7 39.5 66.9 87.2 42.3 44.4

A-07 57,654 22.0 63.4 33.0 69.1 86.8 47.8 43.8

S-07 59,300 20.8 50.0 40.4 70.0 86.7 46.8 52.7

O-07 63,573 21.6 65.0 42.8 68.5 86.9 50.7 54.3

N-07 71,335 22.0 65.5 33.6 66.2 83.7 54.1 51.4

D-07 66,506 22.9 62.2 35.6 75.3 89.0 50.5 12.8

J-08 70,164 22.1 61.2 34.3 66.0 82.6 53.8 38.6

F-08 66,280 21.2 59.4 36.0 70.8 87.0 60.0 41.3

M-08 66,230 22.5 64.0 35.6 67.6 83.2 65.3 49.6

J-08 71,841 20.0 69.9 44.7 72.6 88.6 66.2 61.1

J-08 66,073 21.1 61.1 43.8 73.1 89.2 64.7 51.9

A-08 69,001 23.2 64.5 39.3 73.0 89.8 65.0 45.0

S-08 67,068 23.0 62.6 48.5 74.1 88.9 78.5 43.9

0-08 73,673 25.3 60.0 45.6 69.5 90.3 70.9 55.7

N-08 66,787 24.3 62.8 47.9 72.0 91.1 59.0 55.1

D-08 81,866 24.3 63.9 46.1 70.6 90.0 69.5 53.1

Tot/Ave 1,296,614 22.5 63.0 41.0 69.9 87.2 58.5 47.0

Q1 2009 248,325 24.1 67.1 44.9 71.1 91.3 63.2 54.0



Figure 16.6: Concentrator Monthly Tonnage

Figure 16.7: Concentrator Monthly Metal Recoveries



Figure 16.8: Concentrator Monthly Concentrate Grades

For reference monthly reagent consumptions for the operating year 2008 are presented in the Table

16.2. The reagent suite selected is typical of the mineralogy requirements of similar operations.

Consumptions are normal for an operation of this size.

Table 16.2: Reagent Consumption

Reagent Consumptions (kg/t)

Month SodiumCyanide

Zinc Sulfate

A-3418 Promoter

7583 Copper Sulfate

Lime Floc-

culent

SodiumMeta-

bisulfite Frother

Jan-08 0.288 0.482 0.061 0.155 1.730 0.005 0.018

Feb-08 0.364 0.495 0.054 0.111 0.703 0.008 0.028

Mar-08 0.446 0.461 0.054 0.094 2.065 0.008 0.026

Apr-08 0.423 0.603 0.039 0.165 1.582 0.007 0.041

May-08 0.272 0.492 0.004 0.137 2.067 0.004 0.027

Jun-08 0.309 0.644 0.028 0.177 1.692 0.006 0.043

Jul-08 0.316 0.410 0.030 0.159 1.923 0.004 0.227 0.028

Aug-08 0.300 0.547 0.029 0.020 0.163 1.419 0.003 0.283 0.045

Sep-08 0.337 0.671 0.037 0.112 1.355 0.002 0.306 0.049

Oct-08 0.448 0.624 0.016 0.204 1.384 0.002 0.353 0.025

Nov-08 0.329 0.576 0.021 0.157 1.809 0.003 0.260 0.030

Dec-08 0.293 0.366 0.024 0.140 0.074 0.002 0.354 0.020



16.4 Conclusions

Operations and metallurgical performance at Cozamin are typical of other operations of similar

mineralogy, metallurgy and size. The technical and operations staff are competent and highly skilled

allowing for continual process optimization and modification recommendations to meet production

requirements.

The processing facilities are in good repair showing routine wear and tear. Housekeeping at

Cozamin is of a high quality allowing for safe and efficient maintenance and operations activities.



17 Mineral Resource and Mineral Reserve Estimates

17.1 Mineral Resource Estimate

17.1.1 Introduction

The mineral resource estimate was prepared under the direction of Robert Sim, P.Geo with the

assistance of Bruce Davis, FAusIMM. Mr. Sim is the independent Qualified Person within the

meaning of NI 43-101 for the purposes of mineral resource estimates contained in this report.

Estimations are made from 3-dimensional block models based on geostatistical applications using

commercial mine planning software (MineSight® v4.50.01). The project limits area based in the

UTM coordinate system using a nominal block size of 10 m x 3 m x 3 m with the long axis oriented

along the general W-E strike of the deposit. Sample data is derived from a combination of surface

and underground drill holes plus a series of closely-spaced channel samples taken across the strike of

all development drifts in the orezone. Due to a lack of hangingwall development, drill holes are

often “fanned” out from several designated drill stations, primarily on the level 8of the mine. The

pierce points of drill holes are designed to intersect the mineralized zone at approximately 50 m

spacing. Below the 1,900 m elevation, drill spacing increases to an average of 100 m to 200 m.

The resource estimate has been generated from drill hole sample assay results and the interpretation

of a geologic model which relates to the spatial distribution of copper, zinc, lead, silver and gold in

the deposit. Interpolation characteristics have been defined based on the geology, drill hole spacing

and geostatistical analysis of the data. The resources have been classified by their proximity to the

sample locations and are reported, as required by NI43-101, according to the CIM standards on

Mineral Resources and Reserves.

17.2 Geologic Model, Domains and Coding

The geologic model for the Cozamin deposit is interpreted to be a structural zone hosting an

anastomizing vein system containing varying amounts of pyrrhotite, chalcopyrite, sphalerite, and

pyrite.

The mineralized zone (“Minzone”) domain has been interpreted using a combination of geology

codes, identifying the Mala Noche vein zone plus the presence of mineralization generally above a

grade of 0.4 %CuEq. The copper equivalent grade (CuEq) is calculated in drill hole and channel

samples using the formula:

CuEq=Cu %+(Zn %*0.533)+(Pb %*0.667)+(Augpt*0.583)+(Aggpt*0.01)

Metal prices: $1.50/lb Cu, $0.80/lb Zn, $1.00/lb Pb, $600/oz Au, $10/oz Ag. No adjustments were

made for metallurgical recovery.



Attempts were made to retain a relatively smooth and consistent trend to the Minzone domain.

There are instances where there are intervals of mineralization intersected in drill holes which do not

follow the general trend of the Minzone. Rather than generate irregular deviations in the Minzone

interpretation to “capture” these intervals, these apparent anomalous results have been ignored from

the current model. Further investigation is recommended to better understand how this type of

sulphide occurrence is related to the overall zone of mineralization.

The Minzone is interpreted predominantly as a single zone but there is an internal “dead” zone and a

separate mineralized HW zone that have been interpreted near the area of current underground

development. Most likely, other similar zones exist but these are difficult to interpret based on

current drill hole data. There is evidence that, locally, a separate FW zone may be present in some

areas but this has been left out of the interpreted domain at this stage. Further evaluation is

warranted.

The Minzone domain has been identified as one continuous structural domain over a strike length of

approximately 3.3 km and to a depth of over 1 km below surface. Potentially economic

mineralization is varies throughout the zone; To the west is the San Roberto zone comprised

primarily of copper mineralization with minor zinc, lead, silver and very little gold. To the east is

the relatively smaller San Rafael zone which is more zinc-rich and contains minor lead, silver and

copper. The gold grades at San Rafael are higher than at San Roberto.

In developing the 3-dimensional Minzone domain, a total of 31 drill holes in the San Roberto area do

not fit the interpretation. The collar location, down-hole survey data or some other aspects of these

holes are possibly in error and, as a result, they have been omitted from the geology model

interpretation and subsequent grade interpolation. A review of the source of these errors is

recommended.

There are no indications of significant near-surface leaching or zone of supergene enrichment.

The geology database does not contain alteration assemblage information. There are no indications

that the alteration facies plays any role in the distribution of mineralization in the deposit.

The limits of the property boundaries and previous mined out areas have been taken into account in

the final resource tabulation.

17.3 Available Data

The drill hole database was provided by Capstone’s Cozamin mine geology personnel on

December 8, 2008 in the form of an excel spreadsheet file. There are a total of 365 diamond core

drill holes (DH) in the database with a cumulative length of 105,225 metres. As stated previously,

31 holes have been discarded from the database due to irregularities in their locations.



This leaves a total of 334 holes for a total of 100,408 metres of which 148 (52,854 m) are collared

from surface and 186 (47,554 m) are collared from underground drill stations. The distribution of

drill holes is shown in Figure 17.1.

Figure 17.1: Drill Hole Distribution

The majority of the underground drilling has been conducted from a series of stations on level 8

(2,270 m elevation) which are pushed into the hangingwall of the deposit. Drill holes, fanned from

drill stations at a variety of orientations, are designed to intersect the mineralized zone at regular

intervals. Above an elevation of 1,900 m, drill hole (“DH”) intersections are spaced at an average of

50 m intervals. Below 1,900 m, there are only several holes down-dip of the San Roberto zone,

spaced at approximately 200 m intervals.

In addition to the DH data, there are a series of chip-channel (CH) sample data taken at

approximately 4 m intervals across the strike of the deposit in all development drifts in the San

Roberto mine area. These samples tend to influence a relatively small portion of the resource but the

high density of this data type provides detailed information in the area of active mining.

Drill hole samples range from 0.11 m to 3.5 m in length with an average of 0.52 m. Channel

samples range from 0.06 m to 7 m in length and average 0.75 m. DH samples have been analyzed

for Cu, Pb, Zn, Ag and Au. Channel samples are analyzed at the local mine laboratory for Cu, Pb,

Zn and Ag (there is no gold data in the CH sample data).



Prior to importing the sample data into MineSight, values identified with “<” symbol (denoting

“below the detection limit”) were assigned values of one half of the detection limit (1/2DL). While

validating the database, it was found that a portion of the samples with assayed grades <DL were set

to zero by Capstone personnel. Therefore, there is a combination of both zero and 1/2DL treatment

of these low grade values in the MineSight® data used to generate the resource model. Although

this is an inconsistent approach, it is not considered significant with respect to the resource

estimation.

There is no recorded recovery data in the database. Discussions with Capstone geology personnel

indicate that core recoveries have always been very good. This fact is supported by observations of a

series of randomly selected core intervals during the authors site visit.

During the site visit, the author inspected the location of numerous channel samples and also was

present when the sampling crew was collecting a channel sample. This is done with a crew of three

people using scaffolding to allow access to the drift back. The surface is initially “cleaned” with a

hammer and chisel to remove all large lose material. The sample intervals are then measured and

marked across the drift back with white paint (typically 0.75 m in length). Sample intervals are chip-

channelled with hammer and chisel and tarpaulins are used, laid on the scaffolding and another held

by two helpers, to catch the sample material. The sample is placed in a bag and a sample tag is filled

out with channel number and from-to information.

It appears that the crew does a consistent job in collecting channel samples. When asked why they

do not use a saw to cut true channel samples they said the very hard (siliceous) rocks used up the

blades too quickly and there was too much dust generated. It is the author’s opinion that the

precision of the samples may be poor on an individual basis but the sheer volume of data generated

allows for accurate local estimates of mining scale volumes. Note that channel sample data is

clustered and only influences a relatively small portion of the overall (global) mineral resource.

There are a total of 2,681 individual channels in the database with an average length of 5.36 m (note

that each “channel” is comprised of a series of individual samples that average 0.75 m in length).

Wide zones of mineralization are often mined in multiple passes, each of which have been CH

sampled. Therefore, this average CH sample length is related to average mining drift width rather

than average deposit thickness.

The geologic information is restricted to the designation of “vein” but this appears to be influenced

by economic parameters. It appears the geologist logging the core has often defined the “vein” as

the interval which is expected to be mined, rather than a designation based on some geologic criteria.

As a result, the Minzone domain has been generated using a combination of this geology information

together with the presence of CuEq grades.



Table 17.1: Statistical Summary of Sample Assay Data

Element # samples Total length

(m)Min Max Mean

Standarddeviation

Copper 35,280 27,317 0 16.60 1.13 1.89

Zinc 35,280 27,317 0 36.03 1.06 2.18

Lead 35,280 27,317 0 32.54 0.35 1.57

Silver 35,280 27,317 0 1714.0 51.4 81.07

Gold 16,342 13,060 0 97.40 0.10 0.94

SG 4,092 1,724 1.51 6.37 2.83 0.32

17.4 Compositing

Compositing of drill hole samples is carried out in order to standardize the database for further

statistical evaluation. This step eliminates any effect related to the sample length which may exist in

the data.

In order to retain the original characteristics of the underlying data, a composite length is selected

which is a reasonable reflection of the average original sample length. The generation of longer

composites results in some degree of smoothing which could mask certain features of the data.

Sample intervals are relatively small in the database with approximately 88 % less than 1 m in length

and an overall average length of 0.65 m. A standard composite length of 1 m is considered

appropriate for use in the development of the grade model.

Drill hole composites are length-weighted and have been generated “down-the-hole”, meaning that

composites begin at the top of each hole and are generated at 1 m intervals down the length of the

hole. The contacts of the Minzone domain are honoured during compositing of drill holes. Several

holes were randomly selected and the composited values were checked for accuracy. No errors were

found.

17.5 Exploratory Data Analysis

Exploratory date analysis (“EDA”) involves the statistical summarization of the database in order to

quantify the characteristics of the data. One of the main purposes of this exercise is to determine if

there is evidence of spatial distinctions in grade which may require the separation and isolation of

domains during interpolation. The application of separate domains prevents unwanted mixing of

data during interpolation and the resulting grade model will better reflect the unique properties of the

deposit. However, applying domain boundaries in areas where the data is not statistically unique

may impose a bias in the distribution of grades in the model.

A domain boundary, which segregates the data during interpolation, is typically applied if the

average grade in one domain is significantly different from that of another domain. A boundary may

also be applied where there is evidence that there is a significant change in the grade distribution

across the contact.



17.5.1 Basic Statistics by Domain

The basic statistics for the distribution of copper, zinc, lead, silver and gold both inside and outside

of the Minzone domain have been evaluated. As expected, the results show the mineralization inside

the domain to be significantly different from outside.

Table 17.2: Statistical Summary of Composited Sample Data inside Minzone Domain

Element # samples Total length (m) min max mean Std dev

Copper 18,741 18,292 0 15.22 1.61 1.89

Zinc 18,741 18,292 0 30.77 1.46 2.12

Lead 18,741 18,292 0 31.26 0.47 1.48

Silver 18,741 18,292 0 1,266.0 72.3 77.80

Gold 4,803 4,731 0 94.10 0.18 1.31

SG 4,044 1,729 1.51 6.37 2.83 0.32

17.5.2 Contact Profiles

The nature of grade trends between the two domains is evaluated using the contact profile which

graphically displays the average grades at increasing distances from the contact boundary. Contact

profiles which show a marked difference in grade across a domain boundary, are an indication that

the two data sets should be isolated during interpolation. Conversely, if there is a more gradual

change in grade across a contact, the introduction of a “hard” boundary (i.e. segregation during

interpolation) may result in much different trends in the grade model – in this case the change in

grade between domains in the model is often more abrupt than the trends seen in the raw data.

Finally, a flat contact profile indicates no grade changes across the boundary. In the case of a flat

profile, “hard” or “soft” domain boundaries will produce similar results in the model.

Contact profiles were generated to evaluate the change in grade in all five metals across the Minzone

domain boundary. In all cases there is a significant “step” in grade across the contact. This indicates

that the Minzone domain is significantly different from the surrounding rocks and that the domain

has been developed in such a way that it captures the majority of the mineralized material associated

with the deposit.

17.5.3 Conclusions and Modeling Implications

The results of the EDA indicate that the Minzone domain envelopes samples representing

mineralization which are significantly different from the surrounding area. This indicates that the

domain should be utilized as a hard boundary domain during block grade estimate in the resource

model.

Although there are instances where significant zones of mineralization have been intersected in drill

holes outside of the Minzone domain, these tend to occur as isolated drill “hits” where the nature of

continuity is not understood. As a result, estimations of resources outside of the Minzone domain

have not been conducted because continuity of this mineralization cannot be demonstrated.



The Minzone domain represents a continuous zone of potential mineralization measuring some

3.3 km in strike length and extending for about 1 km down dip. Within this domain there are two

main concentrations of potentially economic mineralization identified by Capstone as the San

Roberto and San Rafael zones. San Roberto is the larger of the two and is located on the western

half of the Minzone domain.

It is primarily a copper deposit but contains appreciable zinc, lead and silver but relatively low gold

grade. San Rafael located on the eastern part of the domain and is higher in zinc and low in

contained copper. It also contains appreciable gold content (~1 g/t Au on average) in comparison to

San Roberto.

During model interpolations, these two main areas have been separated in order to allow for specific

changes to the estimation parameters during interpolation. These two areas have also been presented

separately in the mineral resources table.

17.6 Bulk Density Data

Measurements for bulk density have been conducted by Capstone on a series of samples of drill core

using the water displacement method. Solid pieces of drill core, measuring 10-15 cm in length are

weighed in air. The volume of the piece of core is determined by submerging in water in a graduated

beaker. The bulk density is calculated using the following formula:

Bulk density = weight in air / volume of water displacement

Initially pieces of core were sealed in wax to prevent any effects from porosity in the rocks. Porosity

is not considered a significant issue at Cozamin and therefore, the wax sealing part of the procedure

was eliminated.

The accuracy of the electronic scale is periodically checked with standard weights.

A total of 4,902 measurements were made of which 857 were duplicates to demonstrate the ability to

repeat the results. Three anomalous values were removed from the database with errors apparently

probably attributed to keying errors of the weight of the sample.

Samples selected for bulk density measurements are concentrated in the vicinity of the Minzone

domain. Density values range from 1.51 t/m3 to 6.37 t/m3 with a mean of 2.83 t/m3 and a standard

deviation of 0.32.

The sample distribution is reasonably extensive throughout the area of potential economic interest

with density data present in 135 of 365 holes in the database. The method of determining bulk

density measurements is appropriate and the duplication results demonstrate that the results are

reproducible. The results obtained are appropriate for a deposit which exhibits a variable proportion

of contained sulphides. It is the author’s opinion that the bulk density database is appropriate to use

in estimating mineral resources for the Cozamin deposit.



17.7 Evaluation of Outlier Grades

Histograms and probability plots of the distribution of all elements were reviewed in order to identify

the existence of anomalous outlier grades in the composite database. In addition, a decile analysis of

the data was also conducted in order to quantify the distribution of the contained metals with respect

to the sample density.

Potential outliers, which are identified through the various statistical tests, are ultimately reviewed

visually for their location in relation to the surrounding data. Decisions are made how best to deal

with these samples during interpolation in the grade model. Outliers are dealt with in two possible

ways; top-cutting of the high values or through the application of an “outlier limitation” where

samples above a defined threshold are limited to a maximum distance of influence during grade

interpolation.

San Roberto and San Rafael were evaluated separately due to some potential differing characteristics

of the two zones. Top cutting, where applicable, was applied to the complete database. Outlier

limitations were zone-specific. The results of the evaluations for potential anomalous grades are

listed in Table 17.3

Table 17.3: Outlier Grade Controls

Element Top-Cut Limit Outlier Limits (1)

% Metal Loss in Model San Roberto San Rafael

Copper n/a 10 % 2.5 % -1.7 %

Zinc n/a 15 % 15 % -2.1 %

Lead 15 % (20 comps) 15 % 5 % -2.7 %

Silver n/a 600 g 200 g -1.8 %

Gold 5 g/t (18 comps) 3 g 4 g -28 %

(1) All outlier thresholds limited to maximum 15 m influence during interpolation

The percentage of metal lost in the model due to top-cutting and outlier limitations is considered appropriate for copper, zinc, lead and silver. The much more extensive reduction in gold is due to a combination of overall low grade, a skewed data set and less data due to the absence of gold data in the channel samples.

17.8 Variography

The degree of spatial variability in a mineral deposit depends on both the distance and direction

between points of comparison. Typically, the variability between samples increases as the distance

between samples also increases. If the degree of variability is related to the direction of comparison,

then the deposit is said to exhibit anisotropic tendencies which can be summarized with the search

ellipse. The semi-variogram is a common function used to measure the spatial variability within a

deposit.



The components of the variogram include the nugget, the sill and the range. Often samples

compared over very short distances (even samples compared from the same location) show some

degree of variability. As a result, the curve of the variogram often begins at some point on the y-axis

above the origin – this point is called the “nugget”. The nugget is a measure of not only the natural

variability of the data over very short distances but also a measure of the variability which can be

introduced due to errors during sample collection, preparation and assaying.

The amount of variability between samples typically increases as the distance between the samples

becomes greater. Eventually, the degree of variability between samples reaches a constant,

maximum value. This is called the “sill” and the distance between samples at which this occurs is

referred to as the “range”.

The spatial evaluation of the data in this report has been conducted using a correlogram rather than

the traditional variogram. The correlogram is normalized to the variance of the data and is less

sensitive to outlier values, generally giving better results.

Variograms were generated using the commercial software package Sage 2001© developed by

Isaacs & Co. Multidirectional variograms were generated for composited drill hole and channel

samples located within the Minzone domain. The results are summarized in Table 17.4.

Table 17.4: Variogram Parameters

1st Structure 2nd Structure

Element Nugget S1 S2Range

(m)AZ Dip

Range (m)

AZ Dip

Copper 0.200 0.583 0.217

25 62 65 1470 98 1

21 99 -21 138 8 12

10 3 -14 42 12 -78

Zinc 0.350 0.434 0.216

11 274 6 1403 11 60

7 20 68 354 107 4

3 1 -21 178 19 -29

Lead 0.550 0.389 0.061

12 100 -6 430 91 3

8 28 70 243 0 19

7 8 -19 73 189 71

Silver 0.300 0.547 0.153

69 29 84 135 64 -8

8 272 3 68 333 -2

6 182 6 21 50 82

Gold 0.606 0.269 0.126

186 163 41 1948 87 2

149 296 38 862 358 -35

28 228 -26 178 354 55

(Correlograms conducted on 1 m DH composite data. All exponential models with practical range.)



17.9 Model Setup and Limits

A block model was initialized in MineSight® and the dimensions are defined in Table 17.5. The

selection of a nominal block size measuring 10 m x 3 m x 3 m is considered appropriate with respect

to the current drill hole spacing as well as the selective mining unit (“SMU”) size typical of an

operation of this type and scale. The larger dimension is oriented along the W-E strike of the

deposit.

Table 17.5: Block Model Limits

Direction Minimum Maximum Block size (m) # Blocks

East 746,450 749,400 10 295

North 2,523,450 2,524,449 3 333

Elevation 1,600 2,620 3 340

Blocks in the model have been coded on a majority basis with the Minzone domains. During this stage, blocks along a domain boundary are coded if >50 % of the block occurs within the boundaries of that domain.

The proportion of blocks which occur below the bedrock and topographic surfaces are also calculated and stored within the model as individual percentage items. These values are utilized as a weighting factor in determining the in-situ resources for the deposit.

17.10 Interpolation Parameters

The block model grades for copper have been estimated using Ordinary Kriging (“OK”). The results

of the OK estimation were compared with the Hermitian (“Herco”) polynomial change of support

model (also referred to as the Discrete Gaussian correction). This method is described in more detail

in Section 17.11.

The Cozamin OK model has been generated with a relatively limited number samples in order to

match the change of support or Herco grade distribution. This approach reduces the amount of

smoothing (averaging) in the model and, while there may be some uncertainty on a localized scale,

this approach produces reliable estimations of the recoverable grade and tonnage for the overall

deposit.

All grade estimations use length weighted composite drill hole sample data. Grade estimates are

limited to the area within the Minzone domain boundary. The interpolation parameters are

summarized by domain in Table 17.6.

During grade estimations, the search orientations have been designed to follow a mineralization

“trend” surface created from the average between the hangingwall and footwall surfaces of the

Minzone domain. A temporary elevation item is assigned to all composited drill hole samples and

model blocks which is “relative” to the elevation from the trend surface.



Using the relative elevations during grade estimations, results in a distribution of grades in the

model, which more closely follows the overall subtle variations in the trend of the domain.

The resulting model is a better reflection of the grades in relation to overall interpretation of the

nature of the mineralization in this deposit.

Table 17.6: Interpolation Parameters for San Roberto Area

Interpolation Domain

Search Ellipse Range (m) # Composites Other

X Y Z (1) Min/block Max/block Max/hole

Copper 200 200 10 7 36 6 1 DH per octant

Zinc 200 200 10 7 36 6 1 DH per octant

Lead 200 200 10 8 45 7 1 DH per octant

Silver 200 200 10 7 36 6

Gold 200 200 10 6 15 5

(1) Z search distance relative to trend plane of Minzone domain (trend = avg HW-FW surfaces)

Table 17.7: Interpolation Parameters for San Rafael Area

Interpolation Domain

Search Ellipse Range (m) # Composites Other

X Y Z (1) Min/block Max/block Max/hole

Copper 200 200 10 7 36 6 1 DH per octant

Zinc 200 200 10 7 36 6 1 DH per octant

Lead 200 200 10 8 45 7 1 DH per octant

Silver 200 200 10 7 36 6

Gold 200 200 10 6 15 5 1 DH per octant

(1) Z search distance relative to trend plane of Minzone domain (trend=avg HW-FW surfaces)

Estimates for specific gravity (“SG”) have been made within the Minzone domain using the inverse distance to the power of two (“ID2”) interpolation method. SG estimates use relative elevations to help control the search orientation in relation to the trend of the Minzone domain. Block density values are estimated using a maximum of 30 composites with no more than 5 composites from a single drill hole.

17.11 Validation

The results of the modeling process were validated through several methods. These include a

thorough visual review of the model grades in relation to the underlying drill hole sample grades,

comparisons with the change of support model, comparisons with other estimation methods and

grade distribution comparisons using swath plots.

17.11.1 Visual Inspection

Detailed visual inspection of the block model has been conducted in both section and plan to ensure

the desired results following interpolation. This includes confirmation of the proper coding of blocks

within the Minzone domain. The distribution of block grades were compared relative to the drill

hole samples in order to ensure the proper representation in the model.



17.11.2 Model Checks for Change of Support

The relative degree of smoothing in the block model estimates were evaluated using the Discrete

Gaussian or Hermitian Polynomial Change of Support method (described by Journel and Huijbregts,

1978). With this method, the distribution of the hypothetical block grades can be directly compared

to the estimated OK model through the use of pseudo-grade/tonnage curves. Adjustments are made

to the block model interpolation parameters until an acceptable match is made with the Herco

distribution. In general, the estimated model should be slightly higher in tonnage and slightly lower

in grade when compared to the Herco distribution at the projected cut-off grade. These differences

account for selectivity and other potential ore-handling issues which commonly occur during mining.

The Herco distribution is derived from the declustered composite grades which have been adjusted to

account for the change in support as one goes from smaller drill hole composite samples to the large

blocks in the model. The transformation results in a less skewed distribution but with the same mean

as the original declustered samples.

Selectivity is expected to be made on the primary elements in each area of the deposit; copper in San

Roberto and zinc in San Rafael. Figures 17.2 and 17.3 show that the desired degree of correlation

has been achieved in these two areas.

Figure 17.2: Grade and Tonnage Curves of Herco Copper at San Roberto



Figure 17.3 Grade and Tonnage Curves of Herco Zinc at San Rafael

17.11.3 Comparison of Interpolation Methods

For comparison purposes, additional models were generated using both the inverse distance weighted

(“ID”) and nearest neighbour (“NN”) interpolation methods (the NN model was made using data

composited to 3 m intervals). The results of these models are compared to the OK models at a series

of cut-off grades in a series of grade/tonnage graphs. Overall, there is very good correlation between

all grade models in both the San Roberto and San Rafael areas. Examples are shown in Figures 17.4

and 17.5. Reproduction of the model using different methods tends to increase the confidence in the

overall resource.



Figure 17.4: GT Comparison OK, ID and NN Copper at San Roberto

Figure 17.5: GT Comparison OK, ID and NN Zinc at San Raphael



17.11.4 Swath Plots (Drift Analysis)

A swath plot is a graphical display of the grade distribution derived from a series of bands, or

swaths, generated in several directions through the deposit. Grade variations from the OK model are

compared using the swath plot to the distribution derived from the declustered (NN) grade model.

On a local scale, the NN model does not provide reliable estimations of grade but, on a much large

scale, it represents an unbiased estimation of the grade distribution based on the underlying data.

Therefore, if the OK model is unbiased, the grade trends may show local fluctuations on a swath plot

but the overall trend should be similar to the NN distribution of grade.

Swath plots have been generated in three orthogonal directions for distribution of all five elements in

the Cozamin deposit. Examples in the EW direction for the distribution of copper and zinc are

shown in Figures 17.6 and 17.7.

Figure 17.6: Copper Swath Plot



Figure 17.7: Zinc Swath Plot

There is good correspondence between the models in all elements. The degree of smoothing in the

OK model is evident in the peaks and valleys shown in the swath plots. Deviations tend to occur for

two reasons. First, reduced tonnages near the edges of the deposit tend to accentuate the differences

in grade between models. Second, differences in grade become more apparent in the lower-grade

areas – these typically are the flanks of the deposit where the density of drilling often decreases.

17.12 Resource Classification

The mineral resources for the Cozamin deposit have been classified in accordance with the CIM

definition standards for mineral resources and mineral reserves (CIM, 2005).

The classification parameters are defined in relation to the distance to sample data and are intended

to encompass zones of reasonably continuous mineralization.



The parameters are based on the results of a study of geostatistical methods which define categories

based on confidence limits. Measured resources are defined as material in which the predicted grade

is within ±15 % on a quarterly basis, at a 90 % confidence limit.

In other words, there is a 90 % chance that the recovered grade for a quarter-year of production will

be within ±15 % of the actually achieved production grades. Similarly, indicated resources include

material in which the yearly production grades are estimated with ±15 % at the 90 % confidence

level.

The method of estimating confidence intervals is an approximate method that has been shown to

perform well when the volume being predicted from samples is sufficiently large (Davis, 1997). In

this case, the smallest volume where the method would most likely be appropriate is the production

from a three month period. Using these guidelines, an idealized block configured to approximate the

volume produced in one month is estimated by ordinary kriging using a series of idealized grids of

samples. Relative variograms for copper equivalent grades are used in the estimation of the block

(Relative variograms are used rather than ordinary variograms because the standard deviations from

the kriging variances are expressed directly in terms of a relative percentage.). Note that an

equivalent grade for copper is used due to the presence of several payable metals with copper having

the greatest contribution to value.

The kriging variances from the ideal blocks and grids are divided by twelve (assuming approximate

independence in the production from month to month) to get a variance for yearly ore output. The

square root of this kriging variance is then used to construct confidence limits under the assumption

of normally distributed errors of estimation.

The results of the evaluation indicate that quarterly production can be estimated within ±15 % at the

90 % confidence limit with holes spaced at 15 m intervals. Annual production forecasts, at similar

confidence levels, can be made based on drilling spaced at 60 m intervals. These results are used to

define the classification criteria listed below.

Measured Resources – Model blocks with copper grades estimated by a minimum of three drill

holes or channel samples located within an average distance of 12 metres.

Indicated Resources – Model blocks with copper grades estimated by a minimum of three drill

holes or channel samples located within a maximum average distance of 45 metres.

Inferred Resources – Model blocks which do not meet the criteria for Measured or Indicated

resources but and are within a maximum distance of 80 metres from a single drill hole.

As previously stated, there are no gold analyses available in the channel sample data. Therefore, the

degree of confidence in the gold estimate in the San Roberto area is considered lower than that of the

other four elements in the model. However, the fact that the relative gold grades in the San Roberto

area are very low, it is considered impractical to state the gold portion of the resource separately.



There have been no adjustments made in the presented resources to account for the lack of gold data

in the channel sample data.

Measured resources are confined to underground development areas where dense channel sample

data is present. Indicated resources are located (generally) above the 1,900 m elevation in the San

Roberto area and over the consistently drilled portion of San Rafael.

The distribution of resources by class is shown in Figure 17.8.

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17.13 Mineral Resource Estimate

The Cozamin mineral resources are segregated into the San Roberto area, summarized at a series of

copper cut-off grades and the San Rafael area, summarized at a series of zinc cut-off grades.

Highlighted in Tables 17.8 and 17.9 is the “base case” cut-off grade of 1.15 % copper for the San

Roberto area and 3 % zinc for the San Rafael area. These base case cut-offs are appropriate in

relation to the current operating costs at the Cozamin mine.

Mineral resources exclude all historical (pre-Capstone) and all underground production conducted by

Capstone as of December 31, 2008. Resources are constrained by the Capstone property boundary.

Both the San Roberto and San Rafael zones exhibit relatively continuous zones of potentially

economic mineralization. This continuity, coupled with the fact that all resources occur within

900 m of the surface, suggests that any or all of the resource exhibits reasonable prospects for

economic extraction by underground mining methods. It is important to realize that the tables below

list mineral resources – these are not mineral reserves as the economic viability has not been

demonstrated.

There are no known factors related to environmental, permitting, legal, title, taxation,

socio-economic, marketing or political issues which could materially affect the mineral resource or

mineral reserve estimates.



Table 17.8: San Roberto Mineral Resource Estimate Summary

Cut-off Grade (Cu %)

Ktonnes Cu(%)

Zn(%)

Pb(%)

Ag(g/t)

Au (g/t)

SG(t/m3)

Measured

0.50 2,287 2.00 1.10 0.43 79.2 0.068 2.93

1.00 1,908 2.24 1.03 0.42 84.1 0.064 2.93

1.15 1,749 2.35 1.01 0.41 86.0 0.063 2.93

1.50 1,373 2.63 0.96 0.37 90.0 0.061 2.94

2.00 947 3.04 0.93 0.31 95.1 0.057 2.94

2.50 616 3.48 0.94 0.24 99.0 0.056 2.94

3.00 386 3.92 0.93 0.18 101.3 0.056 2.95

Indicated

0.50 12,303 1.35 1.26 0.30 54.8 0.074 2.91

1.00 7,296 1.76 1.21 0.27 61.2 0.065 2.92

1.15 6,077 1.90 1.20 0.25 63.0 0.063 2.92

1.50 3,963 2.22 1.16 0.23 66.9 0.060 2.93

2.00 2,091 2.67 1.12 0.21 72.8 0.056 2.96

2.50 1,030 3.13 1.04 0.21 79.5 0.052 3.00

3.00 490 3.59 0.94 0.21 87.1 0.050 3.04

Measured + Indicated

0.50 14,590 1.45 1.23 0.32 58.6 0.073 2.91

1.00 9,204 1.86 1.17 0.30 65.9 0.065 2.92

1.15 7,826 2.00 1.15 0.29 68.2 0.063 2.93

1.50 5,336 2.33 1.11 0.26 72.8 0.060 2.93

2.00 3,038 2.79 1.06 0.24 79.7 0.056 2.95

2.50 1,646 3.26 1.00 0.22 86.8 0.054 2.98

3.00 876 3.74 0.94 0.20 93.4 0.053 3.00

Inferred

0.50 4,782 0.95 1.06 0.21 42.4 0.073 2.81

1.00 1,623 1.42 0.98 0.19 49.0 0.063 2.84

1.15 1,100 1.58 0.95 0.17 52.5 0.065 2.85

1.50 504 1.93 1.03 0.18 62.4 0.071 2.92

2.00 181 2.32 1.04 0.16 72.3 0.075 2.98

2.50 41 2.75 0.88 0.15 73.4 0.076 3.13

3.00 5 3.13 0.62 0.10 77.4 0.089 3.26


(2) The “base case” cut-off grade of 1.15 %Cu is highlighted in table.



Table 17.9: San Rafael Mineral Resource Estimate Summary

Cut-off Grade (Zn %)

Ktonnes Cu(%)

Zn(%)

Pb(%)

Ag(g/t)

Au (g/t)

SG(t/m

3)

Indicated

2.0 3,431 2.97 0.21 0.40 33.8 0.441 2.75

2.5 2,407 3.29 0.22 0.43 36.0 0.469 2.75

3.0 1,467 3.64 0.23 0.47 38.3 0.482 2.76

3.5 720 4.07 0.25 0.50 41.4 0.489 2.78

4.0 328 4.48 0.24 0.52 44.3 0.462 2.80

4.5 135 4.87 0.24 0.56 47.5 0.468 2.81

5.0 41 5.22 0.25 0.61 51.3 0.518 2.82

Inferred

2.0 2,642 2.61 0.09 0.37 24.0 0.436 2.65

2.5 1,161 3.11 0.12 0.47 30.2 0.514 2.68

3.0 556 3.55 0.14 0.57 35.8 0.609 2.68

3.5 256 3.92 0.14 0.65 39.5 0.675 2.67

4.0 83 4.32 0.15 0.72 41.9 0.714 2.66

4.5 19 4.76 0.17 0.81 45.4 0.709 2.65

5.0 3 5.16 0.20 0.73 51.4 0.855 2.66


(2) The “base case” cut-off grade of 3 %Zn is highlighted in table.

17.14 Mineral Reserve Estimate

The Cozamin mineral reserves were estimated using the mineral resource model provided by Rob

Sim, P.Geo. of SIM Geological Inc. (“SGI”). The mineral reserve estimate was prepared by SRK

under the supervision of Gordon Doerksen, P.Eng., a Qualified Person under NI 43-101.

The mineral reserve estimate was prepared in accordance with CIM Best Practices guidelines which

define mineral reserves as:

“The economically mineable part of a Measured or Indicated Mineral Resource demonstrated by at

least a Preliminary Feasibility Study. This study must include adequate information on mining,

processing, metallurgical, economic, and other relevant factors that demonstrate, at the time of

reporting, that economic extraction can be justified. A Mineral Reserve includes diluting materials

and allowances for losses that may occur when the material is mined”

The mineral resource model was imported into Gemcom GEMS™ software. Mining shapes were

created to define the limit of economic mining blocks from which mineral reserves were estimated.

GEMS™ was used to interrogate the resource model and report material within the confines of the

mining shapes. Mining recovery and dilution were calculated individually for each stope with due

consideration given to the mining method.



17.14.1 Assumptions

The primary phases in the reserve estimation process are:

1. Generate the resource model;

2. Create and report the tonnes and grades within the mining shapes, and;

3. Calculate recovery and dilution.

4. Prepare a categorised summary of the mineral resource estimate.

Net Smelter Return Block Model

The resource block model provided by Robert Sim included: silver, copper, zinc, lead and gold

grades; in situ density; resource classification; and copper equivalent values for each block. A Net

Smelter Return (“NSR”) value was calculated for each block to determine the economic value of

each tonne of material in the block. The parameters used for the NSR model are shown in Table

17.10. The NSR formula used to determine the value of blocks and subsequently estimate the

mineral reserves was as follows:

$22.97 x %Cu + $2.61 x %Zn + $3.48 x %Pb + $0.07 x g/t Ag = NSR $/tonne

No value was used for gold as it does not provide economic benefit.



Table 17.10: Mineral Reserve Estimate NSR Parameters

Parameter Units Cu Zn Pb Ag

Metal Prices US$/lb for Cu, Zn, Pb

US$/oz for Ag 1.50 0.50 0.45 4.00

Cu Concentrate

Flotation Recovery % 91 56

Concentrate Grade g/dmt or % 25 Variable

Unit deduction units 1.00 50

Percent payable metal % 96.5 95

Smelter costs $/dmt 70.79 -

Refining charges $/payable oz or lb 0.07 0.35

Zn Concentrate



Unit deduction units 0.08 3

Percent payable metal % 85 70


Refining charges $/payable oz or lb - -

Pb Concentrate



Unit deduction units - -

Percent payable metal % 95 95


Refining charges $/payable oz or lb - 1.00

Transport costs were estimated from existing contracts

Cut-Off Grade

An economic NSR cut-off value (“COV”) of $35 per tonne was estimated based on historical and

projected costs developed from first principles. The basis for the COV came from budget operating

costs of $18/t for mining, $12.50/t for processing and $4.50/t for general and administration costs.

For cut and fill (“C&F”) mining shapes, only material from blocks with values greater than the cut-

off were included in the reserve estimates. The selectivity of the C&F method allows for only

economic blocks to report to the mill. Non-economic C&F blocks, if required to be mined, were

assumed to be left in the stope as fill or in an adjacent mined out area.

The design of long-hole mining shapes was based on an overall block cut-off value and width.

Almost all material within a long-hole shape must be mined and sent to the mill, so all long-hole

mining shapes were designed to have an average pre-dilution NSR value greater than the cut-off

value of $35 /t. When the mining blocks were defined, dilution was calculated according to the

mining method.



Blocks with values slightly below the COV were examined and included in the reserve estimate if

their location and geometry were deemed appropriate to reduce their mining cost to a point of being

economic but slightly less than the general $35/t COV. This situation occurred for five of the 26

mining blocks and none of the blocks had a COV of less than $31/t.

Mining Shapes

The Cozamin deposit was divided into individual mining blocks, each representing a group of

planned stopes. Based upon assumed mining and economic constraints, three dimensional mining

shapes within each block were created to encompass regions of material to be mined. These solids

were created using GEMS and are analogous to wireframes or 3D Faces used in other software

packages. Using a block model with density, grade and NSR block values, the volumetric module of

GEMS was used to calculate and report the tonnes, grades and value of material contained within

each block.

Sill (horizontal) and vertical pillars were explicitly designed by Khosrow Aref, P.Eng., Cozamin’s

geotechnical consultant, for the planned mining blocks. Vertical pillars were designed to provide

additional support around fault zones. Sill pillars were estimated on a preliminary basis with a

height to width ratio of 1:1. The individual mining shapes were designed to exclude both pillar

material and material that has been mined in the past at Cozamin (Figure 17.9).

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Cut and Fill Mining Blocks

The outlines of cut and fill stopes were digitized in long section by outlining the blocks that had a

value greater or equal to the cut off value. SGI provided a three dimensional geological solid of the

interpreted ore body and using the solid, the thickness of each cut and fill stope was defined.

Digitized outlines of each cut and fill stope were imported into GEMS and was projected to create a

solid. The geology solid was then used to create an intersection solid between the geology solid and

the projected cut and fill shape solid. This intersection solid is the projected solid trimmed to the

hanging wall and footwall of the geology solid (Figure 17.10).

Figure 17.10: Illustrative Steps Taken to Define Cut and Fill Mining Shapes

LH Stope Mining Blocks

Each LH solid was constructed from rings digitized on vertical sections perpendicular to the strike of

the ore body. The rings were digitized in GEMS (Figure 17.11-1) to closely match to the outlines of

the geological solid, reduce the number of inferred blocks included in the shape, reduce the number

of measured and indicated blocks below the NSR cut-off included in the shape and create final shape

geometries that would be practical to mine with a long hole mining method. GEMS was then used to

create three dimensional solids by connecting the digitized rings, as shown in (Figure 17.11).

From past mining experience at Cozamin, it was observed that internal and external dilution within a

stope typically had grades proportional to the overall average grades of that stope. These dilution

grades were typically greater than: 33 % of the copper grade; 33 % of the lead grade; 33 % of the

zinc grade; and 20 % of the silver grade of the stope the dilution was mined from. These grades

(dilution grades) have been assumed and calculated for all internal and external dilution included in

this resource estimate.



Figure 17.11: Steps to Create Long Hole Mining Shape

Dilution Calculations

External and internal dilution tonnages and grades were factored individually into the reserve

estimate for each mining block.

External dilution is material that is introduced from outside of the planned mining shape. External

dilution due to overbreak was assumed to have dilution grade and was applied to each mining block

as follows:

Cut and fill stopes – 10 % of mining block tonnage.

Long hole stopes with average undiluted mining width less than 7 metres – 14 % of mining

block tonnage.

Long hole stopes with average undiluted mining width greater than 7 metres – 10 % of mining

block tonnage.

For the same hanging wall area, equal quantities of dilution will have a greater effect on the smaller

tonnage of a narrow stope than on the larger tonnage of a wide stope.

External dilution due to unintentional mucking of waste material from the floor of a stope was only

applied to cut and fill stopes since the floors of long hole stopes will not be built on fill material.

Each cut and fill mining block had 3 % of mined tonnage added with zero grade to account for the

mucking of waste.

Internal dilution includes material within the mining shape that is mined and send to the mill with a

value below the NSR cut-off, as well as any material classified as inferred that reports to the mill.

Based on prior mining experience, material classified as inferred was assumed to contain dilution

grades of metals, as defined earlier in this section.



The cut and fill mining method is selective and it was assumed that no material below NSR cut-off

will report to the mill. For this reason, no internal dilution included in the reserve estimates of cut

and fill blocks. The long hole mining method must send almost all mined material to the mill and,

for this reason, all material within each long hole mining shape was included in the reserve estimate.

The material reported by GEMS Volumetric did not include recovery factors or dilution. Recovery

factors and dilution were applied to each mining block specific to the mining method to be used and

geometric parameters of each mining block.

17.14.2 Reserve Estimate

A summary, by mining block, of the mineral reserve estimate is presented in Table 17.11.

Only measured and indicated mineral resource blocks were used in the estimation of mineral

reserves. Inferred resources contained within a mining shape were assigned a grade equivalent to

dilution. As stated previously, dilution grade, including any inferred resource blocks were calculated

as the block grade factored down to 33 % of the block copper grade; 33 % of the block lead grade;

33 % of the block zinc grade; and 20 % of the block silver grade.



Table 17.11: Mineral Reserve Estimates by Stope Block (Dec. 31, 2008)

BLOCK Tonnes Ag

(g/t) Cu(%)

Pb(%)

Zn (%)

Dilution(%)

Cut and Fill Stoping Blocks

8W 110,000 103 1.11 0.38 2.46 13

9C 22,000 79 1.22 0.57 1.43 13

8E 36,000 79 0.64 1.95 2.98 13

9E 66,000 82 1.23 0.67 1.05 13

Narrow Long Hole Stoping Blocks

10T 47,000 54 1.18 0.06 0.74 15

11T 322,000 49 1.28 0.12 0.95 22

13T 69,000 42 1.07 0.31 1.45 10

9W 156,000 74 1.22 0.40 1.80 14

10W 335,000 65 1.12 0.33 1.33 13

11W 362,000 56 1.38 0.17 1.11 14

12W 617,000 55 1.49 0.13 0.91 15

13W 207,000 53 1.33 0.11 1.06 20

11C 433,000 71 2.39 0.28 1.37 14

12C 573,000 66 1.97 0.10 1.35 14

13C 638,000 45 1.71 0.15 1.51 14

14C 451,000 39 1.44 0.15 1.53 15

11E 343,000 63 1.84 0.77 0.92 14

13E 714,000 44 1.79 0.18 0.78 15

14E 144,000 39 1.49 0.18 0.91 14

15E 247,000 37 1.12 0.18 1.15 18

Wide Long Hole Stoping Blocks

12T 318,000 66 1.28 0.23 0.88 10

10C 96,000 60 1.90 0.11 0.65 10

10E 218,000 60 1.65 0.35 0.66 10

12E 472,000 77 2.64 0.23 0.66 10

11EE 751,000 83 1.65 0.81 0.85 10

12EE 349,000 58 1.85 0.38 0.90 10

Total 8,097,000 60 1.66 0.29 1.10 13

*Dilution is taken into account in block tonnes and grade



17.14.3 Risks to the Mineral Reserve Estimate

Geotechnical Conditions

If the competency of the rockmass is less than expected, then additional ground support may be

required. In addition, stope dimensions may need to be adjusted to allow for larger or additional

pillars or a smaller stope size. Increased ground support requirements would increase the mining

cost and therefore the COV, thus reducing the reserve estimate. If the resulting stope size is less than

the minimum practical mining size, then this material could not be included in the reserve estimate.

Economic Conditions

Increased labour and supplies costs can increase the overall mining cost and COV, leading to the

potential exclusion of some material from being included in the reserve estimate. A decrease in

metal prices, or an increase in transportation, smelting, or processing costs will decrease the NSR

value of all material, and this may exclude additional material from being included in the reserve

estimate.

Process Recovery

Process recoveries are one of the key NSR components and as such, any reduction in process

recovery, particularly of copper, would reduce the value of the ore and may make lower grade areas

uneconomic.

Dilution

Dilution of ore grades with waste or low grade material can be an issue in any mine but in particular

with LH stoping operations. To avoid excessive dilution, LH stopes must be carefully designed with

particular reference to geology, assay information, geotechnical conditions. Drilling and blasting

must be optimised to reduce overbreak. Dilution creates a lower grade feed material to the plant and

raises operating costs per unit of metal produced

Metal Prices

Metal prices have the single biggest impact on the NSR and subsequently, mineral reserve

estimation. A reduction in the price of metals, especially copper below their assumed levels would

have a negative impact on reserves. C&F stoping provides the best opportunity for reacting to

fluctuating metal prices.



17.14.4 Mineral Reserve Classification

The mineral reserve classifications were estimated based on the NI 43-101 definitions:

“A ‘Probable Mineral Reserve’ is the economically mineable part of an Indicated and, in some

circumstances, a Measured Mineral Resource demonstrated by at least a Preliminary Feasibility

Study. This Study must include adequate information on mining, processing, metallurgical,

economic, and other relevant factors that demonstrate, at the time of reporting, that economic

extraction can be justified.”

“A ‘Proven Mineral Reserve’ is the economically mineable part of a Measured Mineral Resource

demonstrated by at least a Preliminary Feasibility Study. This Study must include adequate

information on mining, processing, metallurgical, economic, and other relevant factors that

demonstrate, at the time of reporting, that economic extraction is justified.

Mineral reserves were classified as proven reserves if they were derived from measured resources

and probable reserves if they were derived from indicated resources. Table 17.12 shows the mineral

reserve estimate by classification.

Table 17.12: Mineral Reserve Estimate (Dec. 31, 2008)

Classification Tonnes Ag

(g/t) Cu(%)

Pb(%)

Zn(%)

Proven 1,606 76.91 2.02 0.44 0.97

Probable 6,491 55.38 1.57 0.26 1.13

Total 8,097 59.65 1.66 0.29 1.10



18 Other Relevant Data and Information

18.1 Historical Production Data

Production data since the start of the Capstone operation can be found in Table 18.1 below.

Numerous production improvement initiatives allowed has enabled the operation to steadily improve

throughput levels since start-up in 2006.

Table 18.1: Historical Plant Production Data

Head Grade

Month Ktonnes Ag

(g/t) Cu(%)

Pb(%)

Zn(%)

Jan to May 2006

Jun 14.8 67 0.86 0.91 1.71

Jul 17.2 66 1.27 0.52 1.61

Aug 28.4 66 1.35 0.58 1.81

Sep 30.5 65 1.30 0.62 1.89

Oct 31.1 79 1.73 0.83 2.08

Nov 31.4 73 1.58 0.62 1.57

Dec 32.0 70 1.49 0.55 1.61

2006 TOTAL 185.5 70 1.42 0.65 1.77

Jan 2007 33.8 63 1.43 0.53 1.44

Feb 29.7 71 1.84 0.51 1.27

Mar 36.7 73 1.80 0.47 1.39

Apr 35.7 71 1.60 0.59 1.67

May 39.9 70 1.57 0.72 1.86

Jun 44.2 63 1.41 0.65 1.68

Jul 59.3 78 1.57 0.51 1.29

Aug 57.7 72 1.70 0.53 1.02

Sep 59.3 68 1.77 0.51 1.18

Oct 63.6 69 1.55 0.64 1.32

Nov 71.3 67 1.72 0.62 1.27

Dec 66.5 73 2.14 0.52 1.42

2007 TOTAL 597.6 70 1.69 0.57 1.37

Jan 2008 70.2 70 1.82 0.63 1.48

Feb 66.3 72 1.52 0.59 1.24

Mar 66.2 69 1.42 0.73 1.33

Apr 65.0 64 1.79 0.57 1.07

May 69.2 61 1.61 0.59 1.37

Jun 71.8 63 1.60 0.51 1.50

Jul 66.1 67 1.47 0.60 1.66

Aug 69.0 71 1.76 0.40 1.42

Sep 67.1 57 1.69 0.56 1.62

Oct 73.7 54 1.53 0.43 1.26

Nov 66.8 52 1.63 0.50 1.05

Dec 81.9 57 1.60 0.53 0.82

2008 TOTAL 833.2 63 1.62 0.55 1.31

Q1 2009 248.3 56 1.96 0.33 0.81



18.2 Reconciliation of Production Data

Two reconciliations were conducted to determine the accuracy of a) the mine production versus mill

production results and b) the resource model tonnage and grade estimate versus the mill production.

Annual reconciliation results are shown in Tables 18.2 and 18.3. Both reconciliations show excellent

correlation of mine, resource model and mill tonnages and the differences are well within industry

accepted standards. The reconciliation results show future tonnage and grade estimates are

reasonably supported by past practice.

Table 18.2 : Mine vs. Mill Reconciliation

Year Parameter Tonnes Ag (g/t) Cu (%) Pb (%) Zn (%)

2006

Mine Estimate 188,492 87 1.67 0.59 1.88

Plant Sampled 185,483 70 1.42 0.65 1.77

Difference -3,009 -17 -0.25 0.06 -0.11

Percent Difference -2 % -19 % -15 % 10 % -6 %

2007

Mine Estimate 615,685 81 1.76 0.46 1.45

Plant Sampled 597,631 70 1.69 0.57 1.37

Difference -18,054 -11 -0.07 0.10 -0.08


2008

Mine Estimate 821,693 78 1.69 0.54 1.34

Plant Sampled 833,176 63 1.62 0.55 1.31

Difference 11,483 -15 -0.07 0.01 -0.03

Percent Difference 1 % -19 % -4 % 2 % -2 %

Total

Mine Estimate 1,625,870 80 1.71 0.52 1.45

Plant Sampled 1,616,290 66 1.62 0.57 1.39

Difference -9,580 -14 -0.09 0.05 -0.06


Table 18.3: Resource Model vs. Mill Reconciliation

Year Parameter Tonnes Ag (g/t) Cu (%) Pb (%) Zn (%)

2006

Model Estimate* 186,262 75 1.57 0.44 1.50

Plant Sampled 185,483 70 1.42 0.65 1.77

Difference -779 -5 -0.15 0.21 0.27

Percent Difference 0 % -7 % -10 % 47 % 18 %

2007

Model Estimate* 597,544 74 1.66 0.52 1.28

Plant Sampled 597,631 70 1.69 0.57 1.37

Difference 87 -4 0.03 0.05 0.09

Percent Difference 0 % -6 % 2 % 9 % 7 %

2008

Model Estimate* 821,356 70 1.59 0.56 1.32

Plant Sampled 833,176 63 1.62 0.55 1.31

Difference 11,820 -7 0.03 -0.01 -0.01

Percent Difference 1 % -10 % 2 % -2 % -1 %

Total

Model Estimate* 1,608,169 71 1.58 0.52 1.32

Plant Sampled 1,616,290 66 1.62 0.57 1.39

Difference 8,121 -5 0.04 0.05 0.07

Percent Difference 1 % -7 % 2 % 9 % 5 %

*Models generated and reported by Rob Sim



19 Additional Requirements for Technical Reports on Development Properties and Production Properties

19.1 Mining Operations

19.1.1 Deposit context

Mineral reserves at the Cozamin Mine are located within the Mala Noche vein and splays from the

Mala Noche vein. The Mala Noche vein is typically much wider than the high-grade mined portion

of the vein and as a result, dilution from the footwall and hangingwall is usually mineralized.

The mineable vein widths vary from 2 m to greater than 20 m with the average at approximately 6 m.

The dip of the Mala Noche vein varies between 40o and 90o and averages approximately 60o. The

strike length of the deposit is approximately 1.7 km. The mine is relatively shallow with reserves

extending to less than 700 m depth.

Ground conditions in the mine are generally favourable with wide spans observed to be stable with

minimal ground support. In areas where significant faults transect the orebody, the ground

conditions can be poor and vertical pillars are established.

No significant constraints relating to rock temperature or groundwater are anticipated.

19.1.2 Mining Methods

The Cozamin mine currently utilizes two mining methods, mechanized cut and fill (“C&F”) and

longhole (“LH”) open stoping. Initial mine production was based on using C&F stoping. As greater

knowledge of the deposit and geotechnical characteristics were gained, LH stoping was introduced,

leading to higher production rates and reduced mining costs. Approximately 0.23 Mt (3 %) of the

mineral reserves are planned for extraction by C&F method and 7.9 Mt (83 %) by LH methods.

Avoca mining will be used for extraction of approximately 14 % (1.1 Mt) of the resource and will be

used primarily in the eastern part of the resource in the very wide veins. Avoca mining combines the

efficiency of LH stoping with continual backfill placement to maintain ground stability. Although it

has not yet been tested at Cozamin, SRK believes the method will be successful, as it is used at other

mines in similar conditions

In SRK's opinion, the mining methods in use and planned for Cozamin are appropriate for the

deposit context.



19.1.3 Mine Development

Detailed mine development layouts have been prepared for the first year of the LOM plan. The

general sizes of development headings are as follows.

Ramps: 4 m wide x 5 m high

Sublevels: 4 m wide x 4.5 m high

(usually mined to the extent of the ore zone)

Access x-cuts, drawpoints: 4 m wide x 4 m high

Raises: 2 m x 2 m

Primary mine development is carried out by a Mexican mining contractor with Capstone personnel

carrying out all other mining development and production.

19.1.4 Mine Access

There are three main access routes to the mine; the San Ernesto ramp on the west end of the mine,

the shaft in the central part of the mine and the Guadalupana ramp at the east end of the mine. The

shaft currently hoists of ore from the Level 10 and is planned to be deepened as the mine advances.

The lowest level of the mine is planned to be Level 15. The lower portions of the mine will be

accessed via a ramp that will be used for ore haulage directly to surface as well as ore haulage to the

shaft at Level 11.5. Ramp access is planned to be driven to this level with haulage of ore taken up to

Level 11.5.

19.1.5 Material Handling

Ore

Ore is mucked from stopes and in-ore development ends using LHDs. The LHDs transfer the ore

into trucks or into ore passes. Ore is either hauled to surface via the La Guadalupana or San Ernesto

ramps or taken to the shaft and dumped on the crusher grizzly.

Ore trucks are weighed on scales located near the mill after which the ore is dumped onto the Run Of

Mine (“ROM”) stockpile. Ore is re-handled from the ROM stockpile to the primary jaw crusher by a

loader. Oversized ore (> 0.5 m) is broken by a hydraulic rock breaker mounted on an excavator.

Waste

Development waste is used exclusively within the mine as backfill in the cut and fill stopes or

deposited in the mined out LH stopes. The waste is transported directly using LHDs or loaded into

trucks depending on haul distance.



19.1.6 Mine Ventilation

A primary ventilation survey of the Cozamin mine was conducted by an independent contractor in

January 2009. The survey estimated a total intake flow of approximately 130 m3/s of air moving

through the main ramps, shaft and raises.

Mine ventilation is achieved with the use of three exhaust fans, one 260 kW 350 Joy axial vane fan

located on surface above the main central ventilation raise and two 112 kW Jet Air fans located

underground off of the San Ernesto and Guadalupana ramps at either end of the mine.

Secondary ventilation is achieved with auxiliary ventilation fans, ducting and return air raises as

required.

19.1.7 Mobile Equipment

The mine has a fleet of modern mobile equipment that is sufficient for current production

(Table 19.1). As the mine extends to depth additional haul trucks will likely be needed as haul

cycles are lengthened.

Table 19.1: Major Underground Mobile Equipment (Cozamin and contractors)

Equipment Type Model No. of units

Load-haul-dump (“LHD”)

Toro 005 3

Toro 007 4

Atlas Copco 6yd 2

E/H Jumbo drill, single boom 2xAxera 5 14’, 1xAxera 5 16’, 2xOldenburg 16’ 5

Longhole drill Atlas Copco Simba 4.5” DTH 1

Boart Stopemate 2

Haul Trucks 10t – single axle 6

Rock Bolter Sandvik 1

19.1.8 Life of Mine Development and Production Schedule

The mine production rate is established at 1.015 Mt/year to match the budgeted mill capacity of

2,900 tpd based on a 350 day operating year. The LOM plan does not include any significant

stockpiling of lower grade ore.

Table 19.2: LOM Operating Schedule

Parameter UnitAnnual Estimate

2009 2010 2011 2012 2013 2014 2015 2016 2017

Ore Production Kt 1,015 1,015 1,015 1,015 1,015 1,015 1,015 734 258

Total Development m 7,835 6,077 5,058 3,835 5,863 7,450 6,210 4,302 660

*Development metres in 2009 include 170 m of capital development

** Development metres in 2010 include 1,850 m of capital development



19.2 Recoverability

The Cozamin concentrator recoverability used in the LOM plan and the economic analysis are based

on historical results and adjusted for expected plant performance (See Table 19.3)

Table 19.3: LOM Flotation Recovery Estimate

Parameter Unit Ag Cu Pb Zn

LOM Recovery Estimate % 74 91 60 65

19.3 Markets

Copper, lead and zinc concentrates produced by Cozamin are sold under contract to three separate

reputable commodity purchasing companies who market the concentrates to various smelters. The

marketability of the Cozamin concentrate is dependent upon metal grade and the quantity of

deleterious elements in the concentrate.

Cozamin also has a precious metal sales agreement for 100 % of its silver production through a silver

stream purchase agreement with Silverstone Resources Corp. Silverstone is in the midst of being

purchased by Silver Wheaton Corp. with the deal expected to be finalized in May 2009.

19.4 Contracts

Cozamin Mine typically has numerous contracts for supplies, services. In SRK’s opinion, the most

significant contracts were for mine development and concentrate haulage, the terms of which were

within accepted industry practice.

Cozamin’s concentrate and silver sales agreements are summarized in Table 19.4.

Table 19.4: Metal and Concentrate Purchase Contracts

Metal/Concentrate Purchaser Contract Period

% of Production

Metal Price

Copper Concentrate

Trafigura Beheer 2008 to 2011 100 % Cu: LME Cash Settlement

Ag: London Silver Spot

Lead Concentrate Louis Dreyfus 2009 to 2011 100 % Pb: LME Cash Settlement

Ag: London Silver Depot

Zinc concentrate MRI Trading 2009 to 2011 100 % Zn: LME Cash Settlement

Ag: London Silver Spot

Silver Silverstone Resources*

2007 to 2017 100 % The lesser of US$4,00/oz and

the prevailing Ag price

On March 12, 2009, Silverstone Resources Corp. announced that it entered into a definitive agreement with Silver Wheaton

Corp. in which Silver Wheaton will acquire all outstanding shares of Silverstone in exchange for 0.185 common shares of

Silver Wheaton for each common share of Silverstone. The transaction, if approved, is expected to conclude in May 2009.

The details of the concentrate and metals sales agreements were reviewed by SRK and found to be

within the range of industry norms.



19.5 Environmental Considerations

19.5.1 Environmental Assessment and Permitting

This summary of the environmental assessment and permitting requirements is based on work

undertaken for Capstone under the supervision of Nimbus Management Ltd., J. L. Hardy, P.Geo.,

Principal.

The Cozamin Mine lies within a regionally mineralized area that has seen extensive historic mining

over more than 475 years. Host rocks surrounding the mineralized bodies are anomalous in base and

precious metals, providing a halo of elevated metals values that extend a considerable distance

beyond known workings.

Numerous old mine workings, excavations and dumps, as well as some historic tailings are present,

both on, and adjacent to, the Cozamin project site; some lie on mining lands held by Capstone and

others are held by third parties.

Environmental impacts within the project site resulting from historic activities are evident. As well

there are obvious impacts from present day (though sometimes intermittent) operations of

surrounding mines by third parties. The impacts have been discussed, though not necessarily

completely documented, in historic reports.

Environmental impacts within the project site resulting from historic activities are evident. As well,

there are obvious impacts (though sometimes intermittent) from present day operations of

surrounding mines by third parties. The impacts have been discussed, though not necessarily

completely documented, in historic reports.

Prior to Capstone’s involvement in the Cozamin project, several environmental studies had been

completed by previous owners and the San Roberto mine had been permitted to operate at 750 tpd.

Capstone completed the following to support permitting and regulatory approvals with a view to re-

opening the mine and expanding tonnage throughput to 1,000 tpd.

An environmental impact assessment, known in Mexico as a Manifestacion de Impacto

Ambiental (“MIA”);

A detailed study of new ground needed for use as part of an expanded mining operation, known

as the Estudio Justificativo de Cambio de Uso de Suelos (“ETJ”); and

A risk assessment to include all aspects of the expanded operation, known as an Estudio de

Riesgo (“ER”).

Studies required to support the MIA included detailed analysis of: soil, water quality, vegetation,

wildlife, hydrology, hydrogeology, cultural resources and socio-economic impacts. These

investigations identified acid rock drainage and metal leaching as potential concerns manageable

with appropriate mitigation measures.



Static acid-base accounting showed that flotation tailings and some types of waste rock have the

potential to generate acid drainage. However, the country rocks surrounding the deposit have

significant neutralizing capacity and show relatively low permeability. In addition, construction

activities programmed as part of the expansions have significantly reduced identified sources of acid

drainage associated with the historic tailings impoundment, as well as downstream contamination

due to tailings spills by previous operators. Further, during operation both newly generated waste

rock, as well as waste rock deposited during historic operations, will be used as underground back

fill.

Over the longer term, mining will not generate new surface waste dumps, and will significantly

reduce the volume of the existing dumps. Additional mitigation measures involve both engineering

design and operational approaches.

The original MIA was approved by SEMARNAT (Secretaría de Medio Ambiente y Recursos

Naturales) on August 29th, 2005. It remains valid for a period of ten years, and was conditional on

acceptance of the ETJ (accepted January 20, 2006) for Phases 1 and 2 of the tailings dam expansion.

Capstone subsequently filed application for a modification to this MIA and a second ETJ, to include

additional lands leased from the neighbouring ejido for tailings dam expansion. These applications

were accepted on April 20, 2006 and July 10, 2006 respectively.

As a result of significant exploration and operational success in 2006, Capstone completed additional

documentation in support of a MIA for an expanded operation up to 2,600 tpd.

The modified MIA for the expanded operation was approved by SEMARNAT on the 18th of April

2007, and also has an operational term of ten years. With continuing exploration success, Capstone

subsequently submitted documentation for another modification to the MIA for an expansion of the

operation to 3,000 tpd.

The modified MIA for this expanded operation was accepted on the 6th of August 2008, and also has

an operational term of ten years. Capstone applied for an additional ETJ to allow inclusion of

additional lands for tailings dam and water management structures, as well as the entryway for the

new Guadalupana ramp which was approved on the 18th of March 2008.

The government’s statement of approval for the MIA, known as a “Dictamen”, includes detailed

terms and conditions of compliance, as well as an obligation to file operational reports every six

months describing progress in fulfilling the terms and conditions. The MIA Dictamenes provide

authorization for Capstone to complete activities related to the expansion within the approved project

footprint subject to the terms and conditions outlined; these represent normal environmental and

regulatory requirements as described in the MIA, and all costs are included in the operating costs

summary. Development of the required mitigation plans, closure strategy and operational

procedures is ongoing. Detailed reporting includes filing of mitigation and closure plans with

SEMARNAT.



Following a final inspection by PROFEPA (Procuraduría Federal de Protección al Ambiente en el

Estado de Zacatecas), the federal environmental attorney general (i.e. enforcement) branch of

SEMARNAT, Capstone formally received its operating permit on the 20th of October 2006. This is

known in Mexico as a Licencia Única Ambiental (LAU). An LAU for the tonnage expansion to

2,600 tpd was received on the 25th of March 2008. On the 19th January 2009, application was made

to modify the LAU for the tonnage expansion to 3,000 tpd. While unforeseen delays are always

possible, Capstone has supplied all additional information requested, project details and schedules

have been fully and completely discussed with SEMARNAT, and no delays or difficulties are

expected in obtaining approval for this modification. Under the LAU, companies are permitted to

consolidate environmental reporting data on a single form to be submitted annually know as a COA

(Cedula de Operación Anual).

An environmental management and monitoring program is currently underway and will be ongoing.

Data collected are being used to define an operational environmental management and monitoring

program, which will include appropriate environmental management and mitigation plans based on

the principle of continuous improvement. These will be reviewed and revised as necessary, on at

least an annual basis, with results reported as required to Mexican regulators.

Though some assessment and management planning remain to be completed, work to date indicates

that environmental impacts are manageable. It is expected that appropriate management solutions

can be developed within the project schedule and time frames.

In September 2007, Capstone applied to enter into PROFEPA’s National Environmental Auditing

Program (Programa Nacional de Auditoría Ambiental). This ambitious voluntary environmental

audit program is perhaps one of the most advanced programs of voluntary compliance in Latin

America. Known also as the Voluntary Audit or Clean Industry (Industria Limpia) Program, the

initiative was created by PROFEPA in 1992 to promote self-regulation and continuous

environmental improvement. Companies volunteering to join the program pay for an environmental

audit by an accredited, third party, private sector inspector. PROFEPA determines the terms of

reference of the audit, defines audit protocols, supervises the work through certification of the

independent third party auditors, and supervises compliance with the agreed-upon actions.

Companies that enter the program are exempt from the normal inspection activities carried out by

PROFEPA unless a public complaint has been issued. The audit results in an action plan that is

included in an Environmental Compliance Agreement to be signed by PROFEPA and the company

concerned. By way of this process, the Cozamin mine will be able to receive certification under the

Clean Industry Program.

In late 2007, the Cozamin operation underwent a rigorous evaluation by the certified third party

auditor to assess compliance with a broad spectrum of environmental, mine and operational safety,

health and occupational safety laws and regulations. By way of a cooperative process between

Capstone and PROFEPA, Capstone identified areas for improvement, and developed a detailed

Action Plan (with estimated costing) to achieve compliance over an approximate two year period

ending in 2010.



The first trimestral progress report on the Audit was submitted to PROFEPA on the 12th of January

2009, and the second was submitted on the 10th of March 2009. The Company anticipates no

difficulties in achieving the agreed upon Plan and schedule. Work supporting the Plan must be

verified by the independent auditor, who must issue a favourable opinion of the work completed as

compared to the work identified in the Plan before PROFEPA can issue a Clean Industry Certificate

for the operation. The Clean Industry Certificate recognizes operations that have demonstrated a high

level of environmental performance, based on their own environmental management system, as well

as total compliance with regulations.

Apart from public acknowledgement of its clean status, benefits to Capstone include agreement with

its regulators on a defined program of remediation and mitigation, and the ability to participate in no-

cost training programs established by PROFEPA. The Audit certificate is valid for two years and

can be re-authenticated and renewed by an additional audit. Overall under Capstone’s management,

the Cozamin mine has a good environmental record and a good relationship with the environmental

regulatory authorities.

Apart from the issues identified above with respect to potential for acid rock drainage and metal

leaching, other issues of environmental concern relate to potential impacts comparable to those in

underground mines of similar size with flotation tailings impoundments. These include: dust, tailings

handling/management, storm water diversion, combustibles and reagent management/handling,

waste management and disposal and noise.

With expanded production to 3,000 tpd, the present site water balance indicates that while current

underground water sources and operational water management are sufficient at the present time,

these may be unable to sustain a continuous supply of water for site use in the longer term.

Increased demand from the expanded operation as well as intervals of restricted surface and ground

water availability may challenge existing water sources.

The overwhelming majority of water is used in the process plant. Capstone has defined a

comprehensive program to address the challenge of additional supplies which includes the following:

purchase of additional water rights; evaluation of options for tailings management which would

increase water recycle to the plant; recovery of tailings pore water by installing pumping wells in the

tailings impoundment; improved management of site diversions to better segregate and capture fresh

waters entering the site; permitting and development of a new ground water well which will provide

for pump back to the plant for process use; and development and implementation of site-wide

programs of water conservation/recovery to reduce overall water use in all mine areas. Based on

preliminary results to date, these measures provide reasonable expectation that longer-term water

supply needs can be met.



Other operational aspects with environmental implications on which Capstone is presently

concentrating include: mine planning (i.e. in support of planning for progressive reclamation) with

respect to development of a more detailed schedule for prioritized placement of various sources of

waste rock for underground stope fill; waste rock characterization which may support further

definition of priorities for placement of this material underground; and improvements to erosion

control in areas adjacent for the principal arroyos which provide sources of particulate mineralized

material which may enter into watercourses with potentially deleterious effects on surface water

quality. Best management operational practices and site programs of training geared to continuous

improvement are expected to ensure further improvements in the longer term.

19.5.2 Closure Plan

Capstone has adopted a proactive approach to closure. A revised closure plan for the Cozamin Mine

was submitted to SEMARNAT as part of semestral reporting requirements in March 2009. A

number of ongoing closure activities have been completed as part of the site program of progressive

reclamation.

These include: closure of historic workings; reclamation and re-vegetation of exploration drill pads

and accessways disturbed historically and by Capstone; reclamation and re-vegetation of areas of

historic waste dumps and mining activities; clean up of historic tailings spilled downstream from the

tailings impoundment; removal of historic waste rock for use as underground fill and current

construction activities; and definition of diversion channels around the historic Chiripa

impoundment.

Planned “best practices” operational management along with sequential progressive reclamation and

closure planning will reduce new sources of contamination. Reclamation, post-closure monitoring

and follow-up will require more detailed planning, but have the general objective of leaving the land

in a useful, stable and safe condition capable of supporting native plant life, providing appropriate

wildlife habitat, maintaining watershed function and supporting limited livestock grazing. Overall

objectives include the removal of any environmental liabilities and the return of the site to a

condition that resembles pre-mining conditions or restores productivity. Final land use after closure

will be determined in consultation with Mexican authorities.

Once mining stops, equipment and underground infrastructure will be removed and the mine will be

allowed to flood. Based on historic mining following cessation of operations, ground waters are

expected to return to their original phreatic levels in a short time, with no direct point source

discharges to surface anticipated. Mine entryways will be appropriately closed. All salvageable

items will be removed from the site. Remaining quantities of chemicals, reagents, lubricants,

combustibles, etc, will be returned to suppliers, vendors or sold to third parties. Any remaining non-

hazardous waste will be removed to the municipal landfill. Hazardous waste will be removed and

disposed of to an appropriately licensed waste management facility. Buildings, other structures and

surface infrastructure will be dismantled, removed and sold (or donated) where practical.



Remaining disturbed areas will be re-sloped to re-establish natural landscape contours and (where

applicable) pre-existing drainage patterns. In selected areas as necessary erosion control measures

will be implemented. The disturbed areas will be re-vegetated with natural species approved by

SEMARNAT. Roads that will not be required after mine closure will be re-graded and re-vegetated

to approximate pre-mining conditions.

The flotation tailings are potentially acid generating and require careful management during

operations and in closure to minimize potential impacts to the environment. Capstone is currently

evaluating several possible options for future management of tailings during operations. The closure

plan identifies a number of final closure activities to maintain geochemical stability of the tailings

including: diversion channels above the impoundment to limit fresh water flowing into the tailings

from the upper watershed; re-contouring the surface of the tailings impoundment to prevent ponding,

and a final low permeability dry cover to restrict infiltration. Depending on the results of ongoing

water quality monitoring as well as characterization of tailings products, planning for closure design

may include installation of an engineered low permeability cover to limit oxygen entry into the

tailings and minimize seepage.

The impoundment will be covered and encapsulated if monitoring indicates this is required. With

careful engineering design and good quality control on construction this would appear to be a

reasonable concept.

Reclamation obligations will be funded during mining operations, and are not anticipated to involve

measures significantly different than would be expected for an underground mining operation of this

size and type.

An original preliminary closure cost estimate has been revised to include disturbances present to

December 31, 2008 and now totals $2,196,000, including progressive reclamation during operations,

clean up, rehabilitation and reclamation on closure as well as post closure inspection and monitoring.

This amount will be revised and updated on an annual basis to reflect the disturbances present to year

end, the evolving knowledge of specific site conditions and their reclamation requirements, and an

understanding of the success of ongoing progressive rehabilitation, reclamation and closure

activities, as well as prevailing costs for physical and other work related to closure.

19.6 Taxes

The engineering economic model developed for Cozamin for this report does not take into account

taxation and, therefore, the information provided in this section is only for general information.

Detailed tax calculations are typically very complex and take into account many factors of a

corporation’s entire financial performance and not just the results of an individual operation.



Mexican corporate income tax is calculated on the greater of:

28 % of net revenue (profit) after depreciation.

17 % on a cash flow basis allowing for the deduction of capital expenses in the year incurred.

A valued added tax (“IVA”) is paid to the government by Cozamin but is 100 % refundable by the

government and can be used to offset income tax.

Mexican law also has a provision for a profit sharing tax paid to employees. The tax rate is 10 % of

company profit after tax. Cozamin effectively manages the impact of this tax through its corporate

structure.

Property taxes for the mine site are approximately $20,000 per annum.

Cozamin pays a 3 % NSR royalty to Grupo Minero.

19.7 Capital Cost Estimates

Capital expenditures estimated by Capstone for the LOM plan are shown in Table 19.5. Specific

projects have been identified and budgeted in 2009. In 2010 and beyond, estimates have been made

for the required sustaining capital although specific projects have not been detailed. These later

capital expenditures will depend upon the needs of the operation and the availability of capital.

Table 19.5: Capital Cost Summary Estimate

Project UnitAnnual Capital Costs Estimate

TOTAL 2009 2010 2011 2012 2013 2014 2015 2016 2017

Mine M $ 2.5

Mill M $ 1.1

Site Services M $ 1.2

TOTAL M $ 4.8 3.5 2.0 2.5 1.5 1.0 1.5 0.5 0.0 17.3

Capital costs do not include exploration costs as these are deemed to be discretionary and not related

to the economic analysis of the LOM plan.

19.8 Operating Cost Estimates

The operating costs for the LOM plan are based on a combination of historical results and first

principals’ calculations. The increase to mill throughput over the past two years has assisted

significantly in lowering unit operating costs. As shown in Table 19.6, the actual unit costs in the

first quarter of 2009 are approximately 8 % below the 2009 budgeted unit costs and are a good

indicator that unit cost targets may be achieved in 2009.



Table 19.6: Operating Cost Summary Estimate

Cost Area 1

st Qtr 2009

Actual Unit Costs ($/t milled)

2009 Budget and LOM Estimated Unit Costs

($/t milled)

Mining 17.37 18.03

Processing 10.93 12.99

G & A 4.25 4.49

TOTAL 32.55 35.51

The total site unit operating cost of $35.51/ t milled was used for each year in the LOM plan for

engineering economic analysis. Costs projections are done on an annual basis according to cost

assumptions made at the time of budgeting. The main assumptions made in the cost are the Mexican

peso to US dollar exchange rate, labour rates and consumable prices such as diesel, explosives,

reagents, etc. The key elements of the LOM plan are shown by year in Table 19.7.

Table 19.7 LOM Plan with Annual Operating Costs

Parameter 2009 2010 2011 2012 2013 2014 2015 2016 2017 Total

Mining and Milling

Tonnes (000s) 1,015 1,015 1,015 1,015 1,015 1,015 1,015 734 258 8,097

Copper grade (%) 1.70 1.87 1.86 1.88 1.51 1.51 1.52 1.47 1.28 1.66

Zinc grade (%) 1.22 1.12 1.07 0.98 1.04 1.02 1.11 1.18 1.34 1.10

Lead grade (%) 0.51 0.45 0.38 0.34 0.19 0.15 0.17 0.16 0.16 0.29

Silver grade (g/t) 72 72 70 68 53 51 48 43 38 60

Payable Metals

Copper (Mlbs) 33.1 36.6 36.3 36.8 29.4 29.5 29.8 20.8 6.4 258.6

Zinc (Mlbs) 14.8 13.6 12.9 11.9 12.6 12.4 13.5 10.3 4.1 106.1

Lead (Mlbs) 6.5 5.7 4.9 4.3 2.4 2.0 2.1 1.4 0.5 29.9

Silver (Moz) 1.6 1.6 1.6 1.5 1.2 1.1 1.0 0.7 0.2 10.4

Cash costs (US$/lb payable Cu)

Production (on site) costs

1.09 0.99 0.99 0.98 1.23 1.22 1.21 1.25 1.44 1.11

By-product Credits for Zn, Pb & Ag

0.41 0.35 0.33 0.30 0.32 0.31 0.31 0.32 0.38 0.33

Off site cost of Cu concentrate

0.32 0.32 0.32 0.32 0.32 0.32 0.32 0.32 0.32 0.32

Total Cash Cost 1.00 0.96 0.99 1.00 1.22 1.24 1.22 1.26 1.38 1.10

19.9 Economic Analysis

A simplified, pre-tax engineering economic analysis was conducted to model the financial results of

the Cozamin LOM plan. All cost parameters were held constant through the mine life analysis

(Table 19.8). The analysis considered only 2009 and beyond and did not consider any past

performance of the operation or any investments previously undertaken. No taxes, depreciation,

interest or amortization were considered in the model. The production plan for the model was taken

from the LOM plan as described in previous sections of this report.



Table 19.8: LOM Main Economic Analysis Assumptions (all years)

Item Unit Value

METAL PRICES

Copper $/lb 2.00

Zinc $/lb 0.70

Lead $/lb 0.60

Silver $/oz 4.00

FLOTATION RECOVERY

Copper in Cu concentrate % 91

Zinc in Zn concentrate % 65

Lead in Pb concentrate % 60

Silver in all concentrates % 74

OFF-SITE COSTS

TC/RC, Transport, Payables, Penalties, Price Participation

$/t As per current contracts

OPEX

Unit mining cost $/t milled 18.03

Unit processing cost $/t milled 12.99

Unit G&A cost $/t milled 4.49

Unit cost total M$ 35.51

19.9.1 Economic Results

The simplified economic model for the operation is shown in Appendix A. The LOM (from 2009)

results of the analysis are shown in Tables 19.9 and 19.10 and Figure 19.1.

Table 19.9: LOM Economic Analysis Summary (2009 – 2017)

Item Unit Value

LOM PRODUCTION

Ore Mined Mt 8.1

Mill head grade – copper Cu % 1.66

Mill head grade – zinc Zn % 1.10

Mill head grade – lead Pb % 0.29

Mill head grade – silver Ag g/t 60

METAL PRODUCTION

Copper in Cu concentrate t Cu 122,000

Zinc in Zn concentrate t Zn 58,000

Lead in Pb concentrate t Pb 14,000

Silver in all concentrates oz Ag 11,468,000

REVENUE


Royalty (@3 %) M$ 16


COST

Total OPEX M$ 288

Capex (inc. sustaining) M$ 18

ECONOMIC RESULTS (EBITDA)

NPV0 %DR M$ 216

NPV8 %DR M$ 172



The analysis, like the LOM plan, did not use any inferred mineral resources, other than those blocks

that were integral to a stope design. When this occurred, the grade of the inferred material was

reduced to normal dilution.

Based on the assumptions in the model, the project returns a pre-tax net present value (“NPV8 %”) of

$172 M. It must be noted that this high NPV is due to the fact that the analysis was done on 2009

going forward and does not include any of the investment made prior to 2009.

Table 19.10: Pre-tax NPV by Discount Rate

Discount rate EBITDA NPV (M$)

0 % $216

8 % $172

10 % $164

12 % $156

Undiscounted EBITDA

(Annual and Cumulative)

$0

$50

$100

$150

$200

$250

2009 2010 2011 2012 2013 2014 2015 2016 2017

Project Year

$ (

Mil

lio

ns)

Undiscounted Annual EBITDA Undiscounted Cumulative EBITDA

Figure 19.1: Economic Analysis Graph

Based on the LOM plan and the economic assumptions listed previously, copper is income producer

for the mine at 84 % of total NSR metal value. Silver and zinc follow copper, each adding 7 % of

the project value. Lead contributes only 2 % of value and gold, although present in the deposit in

low grades does not contribute to the economics. Figure 19.2 shows the distribution of NSR metal

value in the LOM mineral reserve estimate.



NSR Value by Metal

Pb 2%

Zn 7%

Ag 7%

Cu 84%

Cu Pb Zn Ag

Figure 19.2: NSR Value by Metal

19.9.2 Sensitivity Analysis

Sensitivity analysis was performed using metal prices, mill head grade, capital costs and operating

costs as variables. The value of each variable was changed plus and minus 20 % independently while

all other variables were held constant. The results of the sensitivity analysis are shown in a graph in

Figure 19.3 and in Table 19.11.

As with most mining projects, the NPV of the project is most affected equally by the metals and head

grade. A 20 % increase in metal price or in mill head grade leads to an increase in the pre-tax net

present value using a 8 % discount rate (“PT-NPV8 %”) from $172 M to $254 M. Conversely, a 20 %

decrease in metal prices or plant feed grades has an $82 M negative impact on the PT- NPV8 %

reducing it from $172 M to $90 M, a 48 % decrease.

The operating costs yielded moderate sensitivity levels. A 20 % increase in operating costs over the

life of the mine decrease the PT-NPV8 % by 26 % or $44 M from $172 M from $128 M.

As most of the capital has already been spent on the project, the mine economics are not sensitive to

capital costs. A 20 % increase in the cost yields a $3 M reduction in PT-NPV8 % or about 2 %.

Table 19.11 Sensitivity Analysis

VariablePre-tax NPV8 % (M$)

-20 % 0 % +20 %

Capital Cost 175 172 169

Operating Cost 217 172 128

Metal Price 90 172 254

Grade 90 172 254



Sensitivity of Project Economics

(Pre-tax NPV8%)

0

50

100

150

200

250

300

-20% 0% 20%

Percent Change from Base Case

Pre

-ta

x N

PV

8%

(M

$)

Metal Price Capital Cost Operating Cost Grade

Figure 19.3: Sensitivity Analysis Graph

19.10 Payback

The payback period for the entire project capital spent to date was completed in within the first two

years of operation. The payback on all future capital spending will be almost immediate due to large

cash flows and minimal capital expenditures planned.

19.11 Mine Life

The LOM plan currently has the mine being depleted of reserves in 2017. There are several potential

resource areas that may, if shown to be economic, add life to the mine but are not at a stage where

they can be classed as reserves. It must be noted that these resources may never reach a reserve level

and, therefore, never be mined.

It is SRK’s opinion that, based on the possible extension of the deposit along strike and down dip,

and the past exploration success on the property, there is good potential for the extension of the mine

life.



20 Interpretation and Conclusions

The Cozamin project has been successfully developed into viable mining and milling operation that,

based on the assumptions made, shows a positive return on the mining and processing of current

reserves. The project has exceeded production expectations as estimated in the original feasibility

study.

The mine has the potential to expand its life if some of the current resources can be converted into

reserves, particularly in the San Rafael zone. There is, however, no guarantee that an increase in

reserves will be achieved and will depend upon further exploration, metallurgical, geotechnical and

hydrogeology assessments as well as market conditions such as metal prices and smelter terms.

20.1 Risks

There are many risks to the forecast life of mine plan.

Water Supply and Management

Mine water supply and the current site water balance suggest that existing water sources and water

management approaches may be unable to sustain a continuous water supply to the plant over the

longer operational term at the current operating rate of 3000 tpd. Both the lack of water and the

overabundance of water seasonally are potential issues for the site. The tailings facility has been

recently upgraded to accommodate additional tailings from the current operations, with

improvements made to the fresh water diversion structures. Cozamin personnel are working to

improve the understanding of the water balance over time. Additional rights for underground water

have been purchased. The potential need for additional water during the dry season is being

addressed by construction of a well to tap into deeper ground water downstream from the facilities to

source additional fresh water which could be pumped to the plant for process use. Though permits

for this well are in hand, the well has not yet been drilled.

Mining Control

The mine must ensure accurate drilling and blasting practices are maintained to minimize dilution,

minimize secondary breaking and optimize extraction. Adequate back-up stopes must be available

to give the mine production flexibility should dilution become a problem in a particular stope.

External Factors

Exchange rates, off-site costs and, in particular, base metal prices all have the potential to seriously

affect the economic results of the mine. Negative variance to these items from the assumptions made

in the economic model would reduce the profitability of the mine and the mineral resource and

reserve estimates.



20.2 Opportunities

The main opportunities for the project are:

Improved ore handling system for the hoisting shaft – The LH mining method has the potential

to produce oversized muck that can be a bottleneck at the shaft grizzly and loading pocket.

Improvements in drilling and blasting practices can help alleviate the problem as well as an

improvement in the underground truck dump/grizzly/rock breaker set-up.

Timely updates of the resource model will allow for better mine planning and scheduling.

Previous planning has been conducted using primarily the channel sample results – essentially a

2 dimensional approach to defining mining limits. Monthly updates of a block model using all

available sampling and mapping information will greatly improve the mine design/planning

process.

Continued improvement in metal recovery and concentrate grade as demonstrated in year to date

operating statistics.

Maximizing mill throughput on a sustained basis to reduce unit costs. Mill has operated in excess

of 3,500 tpd for periods of days which is 20% greater than LOM throughput.



additional claims which cover the down dip extension of the Male Noche vein to the east .

Review of 31 drill holes omitted from resource model due to apparent survey issues may result

in an increase in resources. For example, drill hole U62 (16.2 m 3.3% Cu) intersected


resurveying or redrilling (if required), this hole is reintroduced into the database, it may result in




21 Recommendations

Water Management

Ongoing review and assessment of the site water balance should be completed to refine the

components of the balance as well as to establish both effectiveness of the water management

measures implemented to date, and the needs for additional water in the longer term.

Improved Characterization of Acid Rock Drainage (ARD) and Metal Leaching Potential (ML) of Tailings and Waste Rock

Only limited sampling has been completed to characterize tailings entering the impoundment and

waste rock deposited in short term dumps and used in surface construction activities. The material

from the short-term dumps will ultimately be used for backfill in the underground mine. Additional

characterization is recommended on representative samples of both tailings and waste rock to

improve understanding of potential surface and ground water impacts resulting from longer term

storage in the impoundment and underground. Data could be used to define priorities for backfilling

of the more reactive material into the mine to reduce operational impacts to surface waters, as well as

in evaluating tailings management options.

Waste Management

The detailed design of a waste management plan (which incorporates the characterization testing

described for ARD and ML) for the entire site, including the potential for co-deposition of tailings

and waste in the underground mine will need to be studied in detail along with the closure plan to

improve planning of progressive reclamation during operations as well as develop final closure

strategies.

Mine Ventilation

A more rigorous system of mine ventilation measurement and control is recommended to ensure the

volume of air required in each area of the mine is achieved and the efficiency of the ventilation

system is maximized. The ventilation network in the mine should be shown in both detailed plan

and section view layouts with airflows measured and recorded at each junction. These layouts

should show location of fans, regulators, barricades, airflow direction and airflow volumes of fresh

and return air. The layouts should be kept up to date with hard copies available at all times in the

event of an emergency.



Dilution Monitoring and Control

It would be highly desirable to conduct periodic surveys of mined out stopes to compare mined

shapes to planned shapes. The purpose of this exercise would be to assist engineers and operations

personnel in the modification of drilling and blasting practices to minimize dilution, enhance

breakage and optimize extraction. A 3rd party blasting consultant commenced working for the

company in late March 2009.

Review of Irregular Drill Hole Results

During the development of the Minzone domain, a series of 31 holes were identified and excluded

from the database because their information contradicts the surrounding information. Initial

investigations suggest these irregularities are the result of errors made during the recording of these

holes. A thorough investigation from first principles is recommended including visual review of all

core intervals (to uncover potential drilling depth errors), collar surveys, down hole surveys, etc. In

some cases it may be necessary to re-drill some holes in order to confirm the validity of the model.

All future drilling must be closely monitored with drilling progress plotted in a timely manner and

the results compared to the surrounding information. All drill holes must be surveyed after

completion of the hole and the results checked in relation to the digital 3D surveyed development or

3D surface topography. All sample data must be reviewed and questioned in relation to the

surrounding data (assay results and density measurements). Irregularities must be investigated

immediately as it becomes more difficult to resolve problems at a later date.

Monthly Updates of the Resource Block Model for Mine Planning Purposes

Current practice at the mine is to update the resource model on a 1-2 year basis. Life of mine plans

are developed using this model. All shorter term planning is conducted on level plans using

(primarily) the channel sample data results. It is recommended that the practices in the geology

department be modified in order to provide an accurate updated resource model each month which

will provide the basis for all short, medium and long term planning. This transition from a 2-D to a

3-D approach to mine planning should result in improvements in sequencing, extraction efficiencies,

dilution, production rates and costs.

All of the recommendations above are part of the on-going operation of the mine and do not require

a special budget for their implementation.



22 Acronyms and Abbreviations Distance Other

µm micron (micrometre) oC degree Celsius

mm Millimetre oF degree Fahrenheit

cm Centimetre BTU British thermal unit

m Metre cfm cubic feet per minute

km Kilometre elev elevation above sea level

” Inch masl metres above sea level

in Inch hp horsepower

’ Foot hr Hour

ft Foot kW kilowatt

Area kWh kilowatt hour

m2

square metre Ma Million years

km2 square kilometre mph miles per hour

ac Acres ppb parts per billion

ha Hectares ppm parts per million

Volume s second

l litre SG specific gravity

m3 cubic metre usgpm US gallon per minute

ft3

cubic foot V Volt

usg US gallon W Watt

yd3 cubic yard Ohm

bcm bank cubic yard A ampere

Mbcm Million bcm tph tonnes per hour

Mass tpd tonnes per day

kg Kilogram Ø diameter

g Gram SRK SRK Consulting (Canada) Inc.

t metric tonne CIM Canadian Institute of Mining

kt kilotonne NI 43-101 National Instrument 43-101

lb Pound ABA Acid- base accounting

Mt Megatonne or million tonnes AP Acid potential

oz troy ounce NP Neutralization potential

wmt wet metric tonne NPTIC Carbonate neutralization potential

dmt dry metric tonne ARD/ML Metal leaching/ acid rock drainage

Pressure Capstone Capstone Mining Corp.

psi pounds per square inch LOM Life of Mine

Pa Pascal LH Longhole

kPa Kilopascal CF Cut and Fill

MPa Megapascal NPV Net Present Value

Elements and Compounds DR Discount Rate

Au Gold IRR Internal Rate of Return

Ag Silver EBITDA Earnings before interest, taxation, depreciation and amortization

Cu Copper QP Qualified Person as per NI 43-101

Pb Lead P.Eng. or PE Professional Engineer

Hg Mercury P. Geo. Or PG Professional Geologist

Zn Zinc

As Arsenic

CaCO3 Calcium carbonate Conversion Factors

ANFO Ammonium Nitrate/Fuel Oil 2,204.6 lb 1 tonne

1 troy ounce 31.103 48 g



23 Illustrations

All illustrations are included in the body of the report.



24 References

Capstone Gold S.A. De C.V. Mina San Roberto Programa Largo Plazo 2008-2017 Y Programa

Corto Plazo 2009. December 13, 2008

CIM Definition Standards for Mineral Resources and Mineral Reserves, CIM Standing Committee

on Reserve Definitions, December 11,2005

Davis,B.M., Some Methods of Producing Interval Estimates for Global and Local Resources, SME

Preprint 4p,April 1997

Journel and Huijbregts, “Mining Geostatistics”, 1978

Ponce, B. F., Clark, K. F., “The Zacatecas mining district; a Tertiary caldera complex associated

with precious and base metal mineralization” Economic Geology (December 1 1988)

Silverstone Resource Corp. “Silverstone Completes Cozamin Silver Transaction And Closes $32

Private Placement” Press Release, April 4, 2007.

Silverstone Resource Corp. “Silverstone Resources To Be Acquired By Silver Wheaton Corp.” Press

Release, March 12, 2009.

Silver Wheaton Corp. “Silver Wheaton To Acquire Silverstone Resources Corp., Solidifying Its

Position As The Largest Silver Streaming Company” Press Release, March 12, 2009.

Stone, M.S., Barnes, R.B, Hardy, J. Technical Report on the Cozamin Project, Zacatecas State,

Mexico, December 31, 2007



25 Date and Signature Page

This NI 43-101 Technical Report (effective date: March 31, 2009) was written by the following

Qualified Persons:

ORIGINAL SIGNED AND STAMPED

Gordon Doerksen, P.Eng.

Jenna Hardy, P.Geo.

Robert Sim, P.Geo.

Jeff Woods, CP

SRK Consulting (Canada) Inc.

Suite 2200 – 1066 West Hastings Street

Vancouver, B.C. V6E 3X2

Canada

[email protected]

www.srk.com

Tel: 604.681.4196

Fax: 604.687.5532

Capstone Cozamin Technical Report QP Letter - Doerksen.doc

QUALIFIED PERSON CERTIFICATE

I, Gordon Doerksen, P.Eng., am employed as a Principal Consultant - Mining with SRK Consulting (Canada) Inc.

This certificate applies to the revised technical report titled “Technical Report, Cozamin Mine, Zacatecas,

Mexico” dated March 31, 2009 (“Technical Report”).

I am a member of the Association of Professional Engineers and Geoscientists of British Columbia. I graduated with a B.S. (Mining Engineering) from Montana Tech in May 1990.

I have practiced my profession continuously since graduation. I have over twenty years experience in open pit and underground mining operations encompassing technical, production and management roles. I have experience with a variety of commodities at locations in North America, South America, Africa, Asia and Australia. I have three years experience as a consultant conducting and managing all levels of technical studies and reviews. I am a member of the Canadian Institute of Mining.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101 Standards of Disclosure of Mineral Projects (NI 43-101).

I visited the Cozamin Mine most recently on February 10-13, 2009.

I am responsible for Sections 1 to 11, 15, 17.14, 18 to 19.4, 19.6 to 25 and Appendix A of this Technical Report.

I am independent of Silverstone Resources Corp. and Capstone Mining Corp. as independence is described by Section 1.4 of NI 43-101.

I have been involved in previous technical reports and engineering studies at Capstone’s Minto Mine and Kutcho Project prior to the undertaking of this Technical Report.

I have read National Instrument 43-101 and this report has been prepared in compliance with that Instrument.

As of the date of this certificate, to the best of my knowledge, information and belief, the technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

__________________________

Gordon Doerksen, P.Eng. Date: May 22, 2009, Vancouver, BC

APPENDIX A

Simplified Economic Model

UN

ITS

OU

RC

ET

OT

AL

20

09

20

10

20

11

20

12

20

13

20

14

20

15

20

16

20

17

da

ys

inp

ut

3,1

50

3

50

35

0

3

50

35

0

3

50

35

0

3

50

35

0

3

50

Dis

co

un

t p

eri

od

01

2

3

4

5

6

7

8

MA

TE

RIA

L S

CH

ED

UL

E

Min

ing

& M

illi

ng

To

tal O

reM

tca

lc8

,09

7

1,0

15

1

,01

5

1,0

15

1

,01

5

1,0

15

1

,01

5

1,0

15

7

34

25

8

He

ad

Gra

de

Cu

%F

rom

LO

M

1.6

6

1

.70

1.8

7

1

.86

1.8

8

1

.51

1.5

1

1

.52

1.4

7

1

.28

Sensitiv

ity F

acto

r:P

b%

Fro

m L

OM

0

.29

0.5

1

0

.45

0.3

8

0

.34

0.1

9

0

.15

0.1

7

0

.16

0.1

6

1.0

0Z

n%

Fro

m L

OM

1

.10

1.2

2

1

.12

1.0

7

0

.98

1.0

4

1

.02

1.1

1

1

.18

1.3

4

Ag

g/t

Fro

m L

OM

6

0

7

2

72

7

0

68

5

3

51

4

8

43

3

8

Au

g/t

Fro

m L

OM

0

.06

0.0

7

0

.05

0.0

5

0

.05

0.0

6

0

.06

0.0

6

0

.06

0.0

7

NE

T S

ME

LT

ER

RE

TU

RN

Un

it N

SR

Va

lue

sC

u$

/% C

ufr

om

NS

R C

alc

32

.31

3

2.3

1$

32

.31

$

3

2.3

1$

32

.31

$

3

2.3

1$

32

.31

$

3

2.3

1$

32

.31

$

3

2.3

1$

Pb

$/%

Pb

fro

m N

SR

Ca

lc5

.31

5.3

1$

5

.31

$

5.3

1$

5

.31

$

5.3

1$

5

.31

$

5.3

1$

5

.31

$

5.3

1$

Zn

$/%

Zn

fro

m N

SR

Ca

lc4

.25

4.2

5$

4

.25

$

4.2

5$

4

.25

$

4.2

5$

4

.25

$

4.2

5$

4

.25

$

4.2

5$

Ag

$/g

/t A

gfr

om

NS

R C

alc

0.0

7

0

.07

$

0.0

7$

0

.07

$

0.0

7$

0

.07

$

0.0

7$

0

.07

$

0.0

7$

0

.07

$

Au

$/g

/t A

ufr

om

NS

R C

alc

-

-

-

-

-

-

-

-

-

-

To

tal N

SR

Va

lue

$/t

ca

lc6

4.2

7

67

.99

7

2.8

5

71

.77

7

1.9

3

58

.02

5

7.7

8

58

.45

5

6.6

2

50

.90

NS

RC

uM

$ca

lc4

34

.25

5.6

61

.46

0.9

61

.84

9.4

49

.55

0.0

34

.91

0.7

Meta

l price s

ensitiv

ity

Pb

M$

ca

lc1

2.6

2.7

2.4

2.1

1.8

1.0

0.8

0.9

0.6

0.2

facto

r:Z

nM

$ca

lc3

7.8

5.3

4.8

4.6

4.2

4.5

4.4

4.8

3.7

1.5

1.0

0A

gM

$ca

lc3

5.8

5.3

5.3

5.3

5.2

4.0

3.9

3.6

2.4

0.7

Au

M$

ca

lc0

.00

.00

.00

.00

.00

.00

.00

.00

.00

.0

Gro

ss

In

co

me

Fro

m M

inin

gM

$c

alc

52

0.3

8

6

9.0

73

.9

7

2.9

73

.0

5

8.9

58

.6

5

9.3

41

.6

1

3.1

OP

ER

AT

ING

CO

ST

Un

it m

inin

g c

ost

$/t

ore

inp

ut

18

.03

1

8.0

31

8.0

31

8.0

31

8.0

31

8.0

31

8.0

31

8.0

31

8.0

31

8.0

3

Un

it p

roce

ssin

g c

ost

$/t

ore

inp

ut

12

.99

1

2.9

91

2.9

91

2.9

91

2.9

91

2.9

91

2.9

91

2.9

91

2.9

91

2.9

9

Un

it G

&A

co

st

$/t

ore

inp

ut

4.4

9

4

.49

4.4

94

.49

4.4

94

.49

4.4

94

.49

4.4

94

.49

Sensitiv

ity facto

r:M

inin

g C

ost

M$

ca

lc1

46

.0

18

.3

1

8.3

18

.3

1

8.3

18

.3

1

8.3

18

.3

1

3.2

4.7

1.0

0P

roce

ssin

g c

pst

M$

ca

lc1

05

.2

13

.2

1

3.2

13

.2

1

3.2

13

.2

1

3.2

13

.2

9

.5

3.4

G&

AM

$ca

lc3

6.4

4.6

4

.6

4.6

4

.6

4.6

4

.6

4.6

3

.3

1.2

To

tal

OP

EX

M

$c

alc

28

7.5

3

6.0

36

.0

3

6.0

36

.0

3

6.0

36

.0

3

6.0

26

.1

9

.2

NE

T O

PE

RA

TIN

G I

NC

OM

E

NE

T O

PE

RA

TIN

G I

NC

OM

EM

$c

alc

23

2.8

6

3

3.0

37

.9

3

6.8

37

.0

2

2.9

22

.6

2

3.3

15

.5

4

.0

CA

PIT

AL

CO

ST

Min

eM

$in

pu

t2

.55

2.5

52

Mill

M$

inp

ut

1.0

8

1

.08

1

Site

Se

rvic

es &

IT

M$

inp

ut

1.1

6

1

.16

2

Exp

lora

tio

nM

$in

pu

t0

.25

0.2

50

Ge

ne

ral su

sta

inin

g c

ap

ita

lM

$in

pu

t1

2.5

0

3.5

00

2

.00

0

2.5

00

1

.50

0

1.0

00

1

.50

0

0.5

00

Sensitiv

ity facto

r:W

ork

ing

ca

pita

lM

$in

pu

t-

1.0

0T

OT

AL

CA

PIT

AL

CO

ST

M$

ca

lc$

17

.5$

5.0

$3

.5$

2.0

$2

.5$

1.5

$1

.0$

1.5

$0

.5$

0.0

Ea

rnin

gs

Be

fore

In

tere

st,

Ta

xa

tio

n,

De

pre

cia

tio

n a

nd

Am

ort

iza

tio

n (

EB

ITD

A)

Pre

-ta

xU

nd

isco

un

ted

An

nu

al E

BIT

DA

M$

ca

lc$

21

5$

28

$3

4$

35

$3

4$

21

$2

2$

22

$1

5$

4

8%

Dis

co

un

ted

EB

ITD

AM

$ca

lc$

17

2$

28

$3

2$

30

$2

7$

16

$1

5$

14

$9

$2

Un

dis

co

un

ted

Cu

mu

lative

EB

ITD

AM

$ca

lc$

28

$6

2$

97

$1

32

$1

53

$1

75

$1

96

$2

11

$2

15

Dis

co

un

ted

Cu

m.

EB

ITD

AM

$ca

lc$

28

$6

0$

90

$1

17

$1

33

$1

47

$1

61

$1

70

$1

72

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