TECHNICAL REPORT ON THE COZAMIN PROJECT, ZACATECAS … · 23.11.2007 · Technical Report on the...

Technical Report on the Cozamin project, Zacatecas State, Mexico October 2007

Capstone Mining Corp. 1

1 TITLE PAGE

TECHNICAL REPORT ON THE COZAMIN PROJECT, ZACATECAS STATE, MEXICO

COZAMIN PROJECT, ZACATECAS STATE, MEXICO

Zacatecas Mining District (Centred near: 22

o47’00”N, 102

o34’00”W)

Prepared by:

Michelle S. Stone, Ph.D., P.Geo. CAPSTONE MINING CORP.

Suite 1980 - 1055 West Hastings Street Vancouver, B.C., Canada V6E 2E9

and

Robert B. Barnes, B.Sc., M.B.A., P.Eng.

CAPSTONE MINING CORP. Suite 1980 - 1055 West Hastings Street

Vancouver, B.C., Canada V6E 2E9

and

Jenna Hardy, M.Sc., M.B.A., P.Geo. NIMBUS MANAGEMENT LTD.

535 East Tenth Street North Vancouver, B.C., Canada V7L 2E7

For

CAPSTONE MINING CORP. Suite 1980 – 1055 West Hastings Street

Vancouver, B.C., Canada V6E 2E9

October 31, 2007



2 TABLE OF CONTENTS 1 TITLE PAGE ........................................................................................................................... 1 2 TABLE OF CONTENTS ........................................................................................................ 2 3 SUMMARY .............................................................................................................................. 5

3.1 Property Description, Location and Access ................................................................ 5 3.2 Property Ownership and Terms of Agreement ........................................................... 5 3.3 Geological Setting ........................................................................................................... 5 3.4 Property Geology and Mineralization ........................................................................... 5 3.5 Exploration Concept........................................................................................................ 6 3.6 Status of Exploration, Development and Operations ................................................ 6 3.7 Resource Estimation for the Cozamin Project ............................................................ 6 3.8 Reserves........................................................................................................................... 8 3.9 Conclusions and Recommendations............................................................................ 9

4 INTRODUCTION .................................................................................................................... 9 5 RELIANCE ON OTHER EXPERTS................................................................................... 10 6 PROPERTY DESCRIPTION AND LOCATION............................................................... 10

6.1 Location and Property Status ...................................................................................... 10 6.2 Location of Mineralization and Workings................................................................... 11 6.3 Terms of Agreements ................................................................................................... 11 6.4 Environmental Liabilities and Required Permits....................................................... 15 6.5 Environmental and Permitting Considerations and Activities Highlights .............. 15

7 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY .................................................................................................................... 16 8 HISTORY ............................................................................................................................... 18

8.1 Prior Ownership and Ownership Changes ................................................................ 18 8.2 Historical Mineral Resource and Mineral Reserve Estimates ................................ 20

9 GEOLOGICAL SETTING.................................................................................................... 20 9.1 Regional Geological Setting ........................................................................................ 20 9.2 Property Geology........................................................................................................... 23

10 DEPOSIT TYPES ............................................................................................................... 23 10.1 Mineral Deposit Type.................................................................................................. 23 10.2 Geological Concept Used For Exploration of the Property................................... 24

11 MINERALIZATION............................................................................................................. 24 12 EXPLORATION .................................................................................................................. 26 13 DRILLING ............................................................................................................................ 32 14 SAMPLE METHOD AND APPROACH .......................................................................... 42

14.1 Diamond Drill Core Sampling .................................................................................... 42 14.2 Underground Chip Sampling ..................................................................................... 44

15 SAMPLE PREPARATION, ANALYSES AND SECURITY......................................... 45 15.1 Sampling Personnel.................................................................................................... 45 15.2 Drill Core Sample Preparation and Analytical Procedures................................... 45 15.3 Underground Channel Sample Preparation and Analytical Procedures ............ 48

16 DATA VERIFICATION ...................................................................................................... 49 16.1 Introduction................................................................................................................... 49



16.2 QA/QC Summary for Phase IV and V Drill Samples ............................................. 50 16.3 QA/QC Summary for the 2006 – 2007 Underground Chip Channel Samples .. 56

17 ADJACENT PROPERTIES .............................................................................................. 66 18 MINERAL PROCESSING AND METALLURGICAL TESTING ................................. 66

18.1 Introduction................................................................................................................... 66 18.2 Ore Processing ............................................................................................................ 66 18.2.1 Ore Types ................................................................................................................. 66 18.2.2 Previous Plant Performance .................................................................................. 66

19 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ........................... 74 19.1 2005 Initial Resource Estimate ................................................................................. 74 19.2 2007 Updated Resource Estimate ........................................................................... 76 19.2.1 Database, software and three dimensional models ........................................... 76 19.2.2 Domain Modelling .................................................................................................... 77 19.3.3 Compositing .............................................................................................................. 78 19.2.4 Block model .............................................................................................................. 81 19.2.5 Density....................................................................................................................... 81 19.2.6 Block Interpolation ................................................................................................... 81 19.2.7 Classification............................................................................................................. 82 19.3 Mineral Reserves Estimation..................................................................................... 84 19.3.1 Introduction ............................................................................................................... 84 19.3.2 Cut-off Grade ............................................................................................................ 85 19.3.3 Recovery and Dilution ............................................................................................. 85 19.3.4 2007 Cozamin Reserves ........................................................................................ 85 19.3.5 Reconciliation ........................................................................................................... 89 19.4 Conclusions.................................................................................................................. 90

20 OTHER RELEVANT DATA AND INFORMATION....................................................... 90 21 INTERPRETATION AND CONCLUSIONS ................................................................... 90 22 RECOMMENDATIONS ..................................................................................................... 92 23 REFERENCES AND SOURCES OF INFORMATION................................................. 94 24 SIGNATURE, STAMP AND DATE ................................................................................. 96 25 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES .......................... 97

25.1 Mining Operations ....................................................................................................... 97 25.1.1 Introduction ............................................................................................................... 97 25.1.2 Mining Method .......................................................................................................... 98 25.1.2.1 Mechanized Cut and Fill Stoping ....................................................................... 98 25.1.2.2 Long Hole Stope Mining .................................................................................... 100 25.1.4 Ore and Waste Handling ...................................................................................... 102 25.1.5 Mine Ventilation...................................................................................................... 102 25.1.6 Pumping .................................................................................................................. 103 25.1.7 Operations Expansion to 2,200 tpd Capital Project ......................................... 103



25.1.6 Milling....................................................................................................................... 104 25.2 Production and Development Schedules .............................................................. 104 25.3 Mining Costs .............................................................................................................. 105 25.3.1 Mine Operating Costs (in $US)............................................................................ 105 25.3.2 Capital Cost (in $US)............................................................................................. 105 25.3.3 Mill Operating Costs .............................................................................................. 106 25.3.4 General and Administrative Operating Costs .................................................... 106 25.4 General and Administrative - Infrastructure .......................................................... 106 25.4.1 Site Access ............................................................................................................. 106 25.4.2 Personnel ................................................................................................................ 106 25.4.3 Safety and Environmental Department .............................................................. 107 25.4.4 Power Supply and Distribution ............................................................................ 107 25.4.5 Water Supply .......................................................................................................... 109 25.4.5.1 Process Water..................................................................................................... 109 25.4.5.2 Domestic Water .................................................................................................. 109 25.5 Environmental Considerations ................................................................................ 109 25.6 Financial Analysis ..................................................................................................... 112 25.6.1 Mine Operating Costs ........................................................................................... 112 25.6.2 Plant Operating Costs ........................................................................................... 112 25.6.3 General and Administrative Costs ...................................................................... 112 25.6.4 Total Operating Costs ........................................................................................... 113 25.7 Capstone 5 Year Mine Plan..................................................................................... 113 25.7.1 Processing Issues.................................................................................................. 114 25.7.2 Capital Expenditures ............................................................................................. 114 25.7.3 Marketing................................................................................................................. 115 25.7.4 Mine Plan ................................................................................................................ 115 25.7.5 Sensitivity Analysis ................................................................................................ 116

26 CERTIFICATES................................................................................................................ 117 APPENDIX 1........................................................................................................................... 125

QA/QC ANALYSIS PHASES IV AND V ......................................................................... 125 APPENDIX 2........................................................................................................................... 164

QA/QC ANALYSIS CHIP CHANNEL SAMPLES .......................................................... 164 APPENDIX 3........................................................................................................................... 184

SPECIFIC GRAVITY DETERMINATIONS PHASE IV AND V DRILL PROGRAMS.............................................................................................................................................. 184



3 SUMMARY 3.1 Property Description, Location and Access The Cozamin project (Cozamin) is located in the Morelos Municipality of the Zacatecas Mining District near the southeastern boundary of the Sierra Madre Occidental Physiographic Province in north-central Mexico. The project area is located approximately 3.8 km from the city of Zacatecas. Cozamin is accessible by two-wheel drive vehicles via paved roads to the project area boundary where gravel roads in good condition provide access to the mine and most of the surrounding area. The Cozamin project consists of 21 mining (exploitation) and 11 exploration concessions covering approximately 2,016 ha. One additional claim is pending approval by the Mexican Mines Department (an additional 311 ha). 3.2 Property Ownership and Terms of Agreement On January 21, 2004, Capstone Mining Corp. (Capstone) through its 100% owned subsidiary Capstone Gold S.A. de C.V. (Capstone Mexico) entered into an option agreement with Grupo Minero Bacis S.A. de C.V. (Bacis) to acquire the Cozamin project. In June 2006, Capstone acquired 100% interest in the project subject to a 3% Net Smelter Return/tonne of ore (NSR). In February 2007 Capstone paid Bacis the equivalent of 1 million (M) shares as final settlement of the deal. 3.3 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) in the southern Sierra Madre Occidental Physiographic Province in north-central Mexico. The dominant structural features that localize mineralization are of Tertiary age, and are interpreted 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 metasedimentary rocks. 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. 3.4 Property Geology and Mineralization The veins at Cozamin are hosted mainly in volcanic and sedimentary rocks of the Chilitos formation and partly in Triassic metasedimentary rocks of the Zacatecas Formation. The principal vein on the property, the Mala Noche, occupies a series of anastamozing faults and has been traced on surface for more than 5 km. The Mala Noche vein strikes east-west and dips on average at 60º to the north. Mineralization in the Mala Noche vein at Cozamin is interpreted to have been episodic. Early epithermal alteration and mineralization have been superimposed and replaced by higher temperature pyrrhotite and pyrite dominant mineralization. This later



episode of mineralization contains a copper-silver phase that provides the economic values of most interest at Cozamin. This phase of mineralization is interpreted to have a mesothermal origin that is associated with a telescoping, intrusive related, hydrothermal system. 3.5 Exploration Concept The area with the largest historical workings at Cozamin, the San Roberto mine, was selected as the principal exploration target. Widely spaced exploration drill holes by prior operators suggested that mineralization in the Mala Noche vein extended at least 100m below the historical workings and contained significantly higher copper grades (with undiminished silver grades) than encountered in the historical production. Surface exploration drill holes by Capstone subsequently confirmed these higher copper grades over significant mineable widths. This exploration drilling demonstrated that the higher grade copper-silver dominant mineralization was continuous along strike for 1.4 km and extended more than 350m below the historic workings. 3.6 Status of Exploration, Development and Operations Exploration at Cozamin commenced in 2004. By October 2005, the first resource estimate had been completed and a pre-feasibility study was underway. Meanwhile, Capstone commenced underground development and rehabilitation of the processing plant at Cozamin. Additional underground drilling through to mid-2006 increased the resources at Cozamin and indicated that the copper-silver-lead-zinc ore was continuous and mineable over a strike length of at least 1.4 km and 300m depth. In June 2006, refurbishment of the plant was completed and commercial production commenced in September 2006 at 1,000 tpd. The first concentrates were shipped in November 2006. An operational expansion to 2,200 tpd was announced in 2006. Phase IV and V drill programs subsequently commenced to increase and upgrade the current resources. Drilling was completed in July. In October 2007, the expansion was complete and Cozamin was producing 2,200 tpd. Capstone also announced a significant increase in resources which are presented in this report. A revised mine plan has been produced and reserves estimated (section 19). The reserves provide a 5 year mine life at the current expanded production capacity. 3.7 Resource Estimation for the Cozamin Project 2005 Initial Resource Estimate An initial NI 43-101 compliant resource for the Mala Noche vein was estimated by independent consultant Gary Giroux., P.Eng. (Giroux Consultants Ltd.) using data available as of October 5, 2005 (Christopher and Giroux 2005). The database used for the initial resource estimate consisted of 37 exploration surface diamond drill holes numbered CG-04-01 to 33 and CG-05-34 to 37 and 66 underground diamond drill holes numbered CG-05-U01 to U64, U69 and U70. Forty two historic Bacis channel samples were used in the estimation in addition to 6 lines of Capstone channel samples from Level 8 of the San Roberto mine. Capstone geologists determined the entire interval for the Mala Noche vein in each drill hole. The vein interval used for this



estimation was the entire width for the vein regardless of the grade tenor. Grades for copper, silver, lead, zinc and gold were estimated using 2m composites and ordinary kriging. Results are summarized below (Table 3.1). Detailed tables are included in the 2005 Christopher and Giroux technical report available on the SEDAR website. Table 3.1: Resources estimated for the Cozamin mine in 2005 reported with a 1% copper cut-off.

Million lbs* Million ozs* Tonnes

Cu (%)

Ag (g/t)

Pb (%)

Zn (%)

Au (g/t) Cu Zn Pb Ag

Measured 560,000 2.00 85.46 0.70 1.03 0.07 24.7 8.6 12.7 1.5 Indicated 2,430,000 1.86 70.54 0.49 1.05 0.05 99.7 26.3 56.3 5.5 Measured + Indicated 2,990,000 1.89 73.33 0.53 1.05 0.05 124.4 34.9 69.0 7.0 Inferred 3,550,000 1.67 56.99 0.17 1.11 0.04 130.7 13.3 86.9 6.5 Total 6,540,000 1.77 64.46 0.33 1.08 0.05 255.1 48.2 155.9 13.5 * 1 kilogram = 2.2 lbs; 31.103 grams = 1 oz. Troy 2006 Updated Resource Estimate In July 2006, an updated resource estimate was completed by Gary Giroux, P.Eng., of Giroux Consultants Ltd (Table 3.2). Grades for copper, silver, lead, zinc and gold were estimated using 2m composites and ordinary kriging. Details of the resource update are provided in the 2006 technical report by Christopher, Stone and Giroux available on the SEDAR website. Major changes or differences from the October 2005 study included the following: • Data from an additional 159 lines of channel sampling and 46 underground drill holes. These drill holes were both infill to and deeper than previous drill holes. • An additional 2,071 specific gravity determinations bringing the total to 2,990. • A reinterpretation by Capstone of the Mala Noche vein breaking it into a sulfide-rich core surrounded by a mineralized envelope. Three dimensional solids were created for each domain, and each domain was modeled and estimated separately. Table 3.2: Resources estimated for the Cozamin mine in 2006 reported with a 1% copper cut-off.


Cu (%)

Ag (g/t)

Pb (%)

Zn (%)


Measured 550,000 2.58 86.56 0.48 1.07 0.05 31.3 5.8 13.0 1.5

Indicated 2,210,000 2.48 85.13 0.47 1.14 0.05 120.9 22.9 55.6 6.0 Measured + Indicated 2,760,000 2.50 85.41 0.47 1.13 0.05 152.2 28.7 68.6 7.5 Inferred 3,090,000 2.22 79.07 0.27 1.29 0.06 151.3 18.4 87.9 7.9 Total 5,850,000 2.35 82.06 0.37 1.21 0.06 303.5 47.1 157 15.4 * 1 kilogram = 2.2 lbs; 31.103 grams = 1 oz. Troy 2007 Updated Resource Estimate The Phase IV and V drill programs proposed in the October 2006 technical report (Christopher, Stone and Giroux) and partially reported in the March 2007 technical report (Stone and Giroux) were completed in July 2007. This provided information



from 5 additional surface and 69 underground drill holes to update the 2006 resource estimate. An additional 1,125 lines of underground channel sampling were also completed and used in preparation of the revised estimate (Table 3.3). The estimate was prepared by Michelle Stone, P.Geo., of Capstone. Grades for copper, silver, lead, zinc and gold were estimated using 2m composites and inverse distance squared. Table 3.3: 2007 Cozamin resource model estimate.


Cu (%)

Ag (g/t)

Pb (%)

Zn (%)


Measured 2,591,705 2.48 87.11 0.43 1.18 0.03 141.7 24.6 67.4 7.3 Indicated 2,896,158 2.59 86.37 0.32 1.14 0.04 165.4 20.4 72.8 8.0 Measured + Indicated 5,487,863 2.54 86.56 0.37 1.15 0.04 307.1 45.0 140.2 15.3 Inferred 3,162,838 2.36 80.50 0.18 1.03 0.04 164.6 12.6 71.8 8.2 Total 8,650,702 2.47 84.35 0.30 1.11 0.04 471.7 57.6 212.1 23.5

* 1 kilogram = 2.2 lbs; 31.103 grams = 1 oz. Troy 3.8 Reserves The mineral reserves estimate reported at $US40 NSR are shown in Table 3.4. These proven and probable reserves are derived from measured and indicated resources with consideration of mining method, mine plans, mine configuration, mine schedules, mine dilution, recoveries and appropriate costs (processing, general, administration and capital) to demonstrate economic viability. Metal prices used in the NSR calculation were: $8.50/oz silver, $2.25/lb copper, $1.00/lb zinc and $0.60/lb lead. The estimate was prepared by Robert Barnes, P.Eng., of Capstone. Table 3.4: Summary of the mineral reserves at the Cozamin mine on August 31, 2007 reported using $US40 NSR/tonne. Classification Tonnes Cu (%) Ag (g/t) Pb (%) Zn (%) Proven 1,809,719 2.32 84 0.45 1.17 Probable 1,915,248 2.42 81 0.34 1.19 Total 3,724,967 2.37 82 0.40 1.18

Reconciliation of Mineral Reserves. Mineral reserves are adjusted annually by the amount mined, by additions and deletions resulting from new geological information and interpretation, in conjunction with changes in operating parameters and metal prices. However, proven and probable mineral reserves are not usually revised in response to short-term fluctuations in the metal markets. The following is a reconciliation of the proven and probable mineral reserves at Cozamin to September 30, 2007.



Table 3.5: Reconciliation of mineral reserves at Cozamin. Tonnes Opening balance, February 2006* 2,260,000Additions 1,896,373Less Tonnes Milled 431.406Closing balance as of September 30, 2007 3,724,967

*2006 reserves were calculated using metal prices of: $1.25/lb copper, $6.25/oz silver, $0.38/lb lead and $0.50/lb zinc. 3.9 Conclusions and Recommendations The 2007 surface and underground drill programs have been successfully completed. The results of the drilling has increased and upgraded the 2006 resources and reserves. Capstone has doubled its commercial production at Cozamin and currently has reserves sufficient for more than 5 years of production at a rate of 2,200 tpd. The economic analysis based on actual operating costs, reserves and metal prices significantly below current prices indicate robust economics. The underground drill program should continue infill drilling to upgrade the inferred resources to measured and indicated within the main zone in the Mala Noche vein 600 metres east and west of the San Roberto shaft. Additional drilling should be planned to test down dip and to the east and west on the Mala Noche vein. 4 INTRODUCTION

This Technical Report was prepared for Capstone Mining Corp. (Capstone) as a requirement associated with the company filing its Annual Information Form. This report presents an update on exploration and development activities completed by Capstone after the submission of the Technical Report dated February 23, 2007 (Cristopher, Stone and Giroux). This report also includes an update of the resources and reserves at Cozamin. This report has been prepared using data obtained from diamond drilling, underground channel sampling, geological mapping and the mining/processing operations at Cozamin from the period 1 March 2004 to 15 October 2007. Information and conclusions from numerous Capstone reports and press releases during this active period of exploration and development are used throughout this report and are referenced in section 23. Qualified Persons responsible for preparation of this report are Michelle Stone (P.Geo., Senior Geologist, Capstone; sections 1 - 19.2, 22 and 23), Robert Barnes (Vice President of Operations, Capstone; sections 19.3 - 25.4 and 25.6 - 25.7), and Jenna Hardy (P.Geo., Nimbus Management Ltd.; section 25.5).



Michelle Stone has completed 12 site visits to Cozamin between January 2006 and October 2007. Robert Barnes has been at site supervising the startup and expansion of the Cozamin mine since January 2005. Jenna Hardy routinely visits the mine to monitor the environmental program. 5 RELIANCE ON OTHER EXPERTS The writers are required by NI 43-101 to include descriptions of the property titles and terms of legal agreements that are presented in the following sections. The writers have not reviewed these property agreements and title documents. However, these documents have been reviewed by Capstone’s attorney of law in Mexico, Carlos Galvan Pastoriza. All 33 current titles have been reviewed to date and the writers have accepted his opinion that these titles and legal agreements are valid. 6 PROPERTY DESCRIPTION AND LOCATION 6.1 Location and Property Status The Cozamin polymetallic base metal project is located in the Morelos Municipality of the Zacatecas Mining District near the southeastern boundary of the Sierra Madre Occidental Physiographic Province in north-central Mexico (Figure 6.1). The project area is centered near coordinates 22º47’00”N latitude and 102º34’00”W longitude on the 1:250,000 Zacatecas topographic map sheet (F13-6).

Figure 6.1: The location of the Cozamin project near the southeastern boundary of the Sierra Madre (shaded dark yellow) in central Mexico.



The Cozamin project consists of 11 exploration and 21 mining (exploitation) concessions covering approximately 2,016 ha (Table 6.1 and Figure 6.2). One additional concession is pending approval for registration with the Mexican Mines Department (an additional 311 ha). Mining concessions registered in the name of Luis Jorge Smith Barajas (Capstone’s Exploration Manager based in Mexico), are in the process of being registered in the name of Capstone at the Public Registry of Mining of the Federal Mining Bureau of the United Mexican States. All other concessions are currently registered in the name of Capstone.

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 are determined to be correct. Capstone purchased 122 ha of surface land rights in 3 parcels at Cozamin from the Ejido Hacienda Nueva and the Ejido La Pimienta (Figure 6.2). A payment of 50% has been made up front and the remaining 50% will be paid when the properties complete registration. One parcel of 11 hectares has already completed registration and the purchase for that parcel finalized. The Cozamin property requires land rental and government fee payments on the mining concessions. In 2006 the taxes totalled $US28,372. A similar amount is expected for 2007. The project is 100% owned by Capstone subject to a 3% net smelter royalty (NSR) payable to Bacis. 6.2 Location of Mineralization and Workings The Cozamin project covers approximately 5.5 km of the trend of the Mala Noche vein system (Figure 6.3). The Mala Noche vein has been developed and mined by several operators with the mineralized trend consolidated by Penoles before acquisition by a group of miners headed by Mr. Jacek Zanewicki, and then by Bacis 1996. A long section view of the entire Mala Noche vein system (Figure 6.3) shows several of the named workings and mines. 6.3 Terms of Agreements By option agreement dated October 23, 2003, superseded by an agreement dated January 21, 2004, Capstone entered into an option agreement with Bacis to acquire the Cozamin property and four other Bacis silver exploration projects. Capstone earned a 90% interest in the projects by incurring $US10,000,000 (of which $5,000,000 on Cozamin) in exploration and development expenditures over five years, with a minimum year one expenditure of $US1,000,000 (spent). Capstone paid $US 250,000 and issued 1 million common shares to Bacis on the date of approval (January 23, 2004) of the Joint Venture by the TSX-V (issued), and 1 million shares on the anniversary of the date of approval for the next two years (issued). Upon Capstone earning its 90% interest, Bacis had the following options:



Table 6.1: Concession details as of September 30, 2007 (100% Capstone).

CONCESSION NAME TITLE

NUMBER TYPE HECTARESDATE GRANTED (DD/MM/YYYY)

EXPIRY DATE (DD/MM/YYYY)

ACUEDUCTO 224469 EXPLORATION 13.559 13/05/2005 12/5/2011 ACUEDUCTO FRACCION 1 224470 EXPLORATION 9.598 13/05/2005 12/5/2011 LA GLORIA 224474 EXPLORATION 4.1372 13/05/2005 12/5/2011 LA PARROQUIA 224471 EXPLORATION 1.2601 13/05/2005 12/5/2011 LA SIERPE 224503 EXPLORATION 4.2638 13/05/2005 12/5/2011 LA SIERPE FRACCION 1 224504 EXPLORATION 0.0108 13/05/2005 12/5/2011 SAN LUIS I* 223325 EXPLORATION 290.6121 2/12/2004 1/12/2010 SAN LUIS II* 224466 EXPLORATION 133.8409 13/05/2005 12/5/2011 SAN LUIS II FRACCION I* 224467 EXPLORATION 2.1713 13/05/2005 12/5/2011 SAN LUIS II FRACCION II* 224468 EXPLORATION 2.4654 13/05/2005 12/5/2011 UNIFICACION CARLOS 224657 EXPLORATION 542.5265 27/05/2005 26/05/2011 EL RANCHITO 228343 MINING 11.2997 8/11/2006 7/11/2056 EL RANCHITO FRACC 1 228344 MINING 0.6189 8/11/2006 7/11/2056 GABRIELA II 203364 MINING 18.9438 19/07/1996 18/07/2046 LA LIGA 217237 MINING 20.1817 2/7/2002 1/7/2052 LA PROVIDENCIA 223954 MINING 60 15/03/2005 14/03/2055 LA SECADORA 219630 MINING 9 26/032003 25/03/2053 LA VETA 228345 MINING 1.4533 8/11/2006 7/11/2056 LORENA 227712 MINING 318.5825 28/07/2006 27/07/2056 ORLANDO 225620 MINING 11.7899 23/09/2005 22/09/2055 PLATEROS 188806 MINING 9 29/11/1990 28/11/2040 PLATEROS DOS 208838 MINING 50 15/12/1998 14/12/2048 SAN BONIFACIO 217858 MINING 40.8518 27/08/2002 26/08/ 2052 SAN JACINTO FRACCION 1 202437 MINING 78.7955 24/11/1995 23/11/2045 SAN JACINTO FRACCION 2 202438 MINING 17.7846 24 /11/1995 23/11/2045 SAN JUDAS 226699 MINING 14.5989 17/02/2006 16/02/2056 SAN NICOLAS 200150 MINING 5.3697 15/07/1994 14/07/2044 SANTA BARBARA FRACCION 1 218259 MINING 82.9691 17/10/2002 16/10/2052 SANTA BARBARA FRACCION 2 202645 MINING 9.5938 8/12/1995 7/12/2045 SANTA BARBARA FRACCION 4 202628 MINING 0.4585 8/12/1995 7/12/2045 SANTA LUCIA 195187 MINING 18.7267 25/08/1992 24/08/ 2042 SARA 228086 MINING 231.9436 29/09/2006 28/09/2056 ANABEL 229238 PENDING 310.77 27/03/2007 PENDING TOTAL (EXCLUDING PENDING) 2016.4071

* Mining concessions registered in the name of Luis Jorge Smith Barajas (Capstone’s Exploration Manager based in Mexico), are in the process of being registered in the name of Capstone at the Public Registry of Mining of the Federal Mining Bureau of the United Mexican States.



Figure 6.2: Concession map of the Cozamin project area showing the boundary of the land for which surface rights are rented from the local Ejido.



Figure 6.3: Long section view of the Mala Noche vein system showing the location of historic workings.



• Bacis may maintain its 10% carried interest by participating in production costs. • Bacis may sell its carried 10% interest to us in exchange for the value of the

project to be determined by an appraiser jointly appointed, which price may be paid in cash or our common shares at our choice.

• Bacis may exchange its 10% carried interest in each of the projects for the right to instead receive an additional 1.5% net smelter return (“NSR”).

Capstone also acquired the option to acquire an interest in any of the projects individually. To acquire the Cozamin property, Capstone had to, in addition to having paid $US250,000 and delivered 1,000,000 shares to Bacis on the Acceptance Date, spend a minimum of $US1,000,000 on or before the first anniversary of the Acceptance Date on any of the projects (spent) and at least $US5,000,000 in total on the Cozamin property (spent). Capstone must pay $US1,000,000 in cash or shares to Bacis 180 days after the San Roberto mine has been put into production (done). Capstone also had to assume an existing debt which derives from credit granted by the Fideicomiso de Fomento Minero (“FIFOMI”) to Bacis to develop the Cozamin property (paid). Capstone completed the obligations in order to acquire a 90% interest in the Cozamin property and have exercised the option in regard to the Cozamin property. On June 30, 2006, Bacis converted its 10% interest in the Cozamin project to a 1.5% NSR, thus leaving Bacis with a 3% NSR royalty regarding the Cozamin project. Capstone has transferred its rights to option to acquire the silver exploration properties to Silverstone Resources Corp, (Silverstone). 6.4 Environmental Liabilities and Required Permits The writers are not aware of any significant environmental liabilities related to the current Cozamin project. Dispersed tailings from historic operations are present downstream from the current operation in drainages below the mine and below the tailings impoundment. A number of the historic workings have old waste dumps of which some contain sulphides. However, there are no permanent water sources in these arroyos, except for flows derived from activities related to mine dewatering during operations. The veins are characteristically low or moderate sulphidation. Country rocks hosting the veins contain significant neutralizing capacity, and limestone is one of the main units found east of the Zacatecas area, so if necessary neutralizing material would not be expensive to obtain. Baseline environmental studies and evaluation of the tailings area have been conducted, and surface and subsurface water quality monitoring are ongoing. These will be used to develop operational strategies for management of site waters and historic wastes. 6.5 Environmental and Permitting Considerations and Activities Highlights Highlights of Capstone’s environmental and permitting activities (see section 25 for details) include:

• Capstone retained J.L. Hardy, P.Geo., and principal of Nimbus Management Ltd. of Vancouver, B.C. to oversee environmental impact and permitting requirements.



• In February 2005, Capstone commissioned Clifton Associates Ltd. of Guadalajara, Jalisco, Mexico to complete the necessary environmental assessment to support submission of the MIA (Environmental Impact Assessment), ETJ (Change of Land Use) and ER (Risk Assessment) for an expanded Cozamin mine project.

• Capstone submitted its MIA for the expanded Cozamin mine project on July 20, 2005. The MIA “dictamen”, valid for ten years, outlying required terms and conditions was received August 29, 2005.

• Capstone submitted its ETJ for Phases 1 and 2 of the tailings dam expansion on the 27th of June 2005, and received its dictamen February 14, 2006.

• Capstone submitted its ER in December 2005 and received its dictamen on the 29th of August 2006.

• Capstone received its operating permit (LAU) from SEMARNAT in October 2006.

• The MIA for the expanded operation was approved by SEMARNAT in April 2007 and has an operational term of 10 years.

• In September 2007 Capstone initiated participation in the Mexican government sponsored environmental audit program.

7 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY Access to the Cozamin project is by paved roads from the city of Zacatecas to the property boundary. Gravel roads in good condition provide access to the office, mine and plant area, and most other parts of the property (Figure 7.1). The Cozamin mill, in the west-central part of the property, is situated approximately 3 km north-northwest of the Zacatecas city centre. Zacatecas is the largest city in Zacatecas State, Mexico, with a population of approximately 135,000. An international airport services the city with multiple flights daily to and from Mexico City. Local residents of Zacatecas have a strong mining tradition and provide the Cozamin project with a knowledgeable source of labour. Drilling and mining contractors are available in Durango, Zacatecas, Fresnillo and other areas of Mexico. The Cozamin project is located in the Western Sierra Madre Physiographic Province (Sierra Madre Occidental Province) near the boundary with the Mesa Central Province (Central Plateau Province). The Zacatecas area is characterized by abrupt north-northwesterly trending mountains with the Sierra Veta Grande to the north and the Sierra de Zacatecas to the south. The city of Zacatecas is located in a fault line valley that separates the two sierras (Ponce and Clark, 1988). The elevation in the Zacatecas area is approximately 2,400m, and the local relief is approximately 200m.



Figure 7.1: Location of the Cozamin mine office, plant and other facilities near the city of Zacatecas. 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 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, crasicaule bushes and a variety of nopales. Standing bodies of water are dammed as most streams are intermittent. Capstone owns and operates a 2,200 tpd process plant, an operational underground mine and headframe, laboratory and several service buildings at the Cozamin mine site. The buildings are connected to the local power grid. The area has sufficient water for exploration and mining. Capstone purchased 122 ha of surface land rights in 3 parcels at Cozamin from the Ejido Hacienda Nueva and the Ejido La Pimienta (Figure 6.2). A payment of 50% has been made up front and the remaining 50% will be paid when the properties complete registration. One parcel of 11 hectares has already completed registration and the purchase for that parcel has been finalized.



Tailings are deposited into the existing tailings dam located immediately south and east of the mill facilities and process buildings (Figure 7.1). The proposed plan, as engineered by Vector Engineering (see Feasibility Report on the Cozamin project: Rodger 2006, available on the SEDAR website), is to initially store an additional 4 million tonnes of tailings in this facility. The first lift of expansion was a 12m lift constructed on a modified centre line method which will store approximately another 760,000 m3. This first lift will provide 18 months containment at the expanded processing rate before another lift will be required. 8 HISTORY 8.1 Prior Ownership and Ownership Changes In pre-Hispanic times, the area was inhabited by Zacatecan Indians who mined native silver from the oxidized zone of argentiferous vein deposits in the Zacatecas Mining District (Figure 8.1). 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 (March), and the San Bernabe and Los Tajos del Panuco mines (June and November, respectively) on the Mala Noche vein system. The initial operations worked only the oxides for silver and some gold, and later the sulfide 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 CRM (1992) to be 750,000,000 ounces of silver from 20M tonnes grading over 900 g/t silver and approximately 2.5 g/t gold. Lead, zinc and copper have also been recovered but the production and grades were not estimated. Minera Cozamin was established in 1980 by Penoles to consolidate concession holdings over the Mala Noche vein. Penoles established a 250 tpd mill before vending the property to a group of miners headed by Mr. Jacek Zanewicki. In 1996 the property was acquired by Bacis for $US6.8M. Bacis expanded the mill to a 750 tpd flotation plant, and processed 250,000 tonnes of ore grading 1.2% copper, 90 g/t silver, 0.5 g/t gold, 1.8% zinc and 0.6% lead. A significant reduction in metal prices occurred between 1997 and 1999, and Bacis was forced to close the underground mine. The crushing unit was moved off the property but the remainder of the mill was kept intact.



Figure 8.1: Plan showing the distribution of mineralized veins near Zacatecas.



Capstone obtained an option to acquire the Cozamin property in January 2004, and retained Peter Christopher & Associates Inc. to prepare a technical report on the property. In order to evaluate the project and prepare a technical report in the form required by NI 43-101, the Cozamin project was examined by Christopher on September 12th and October 11th 2003. Subject to Christopher’s examination, a technical report was prepared for submittal to the TSX-V. On January 23, 2004, Capstone received regulatory approval for the option agreement dated January 23, 2004, between Capstone, Bacis and Capstone Mexico. Capstone received the right to acquire an undivided 90% beneficial interest in five properties (Cozamin, Promontorio, Montoros, Copala and Claudia). By June 2006, Capstone had acquired 100% interest in the Cozamin property subject to a 3% NSR. 8.2 Historical Mineral Resource and Mineral Reserve Estimates In 1998, sampling of surface and underground workings and drill holes was used by Bacis to calculate resources and reserves (Table 8.1). The resources and reserves were calculated by Mariano Carrizales (now Superintendent – Geology at Cozamin) using a cross sectional method including underground channel sampling and diamond drill hole information. Rigourous quality assurance/quality control programs were not in place during this time and therefore the quality of the assays can not be properly determined. The resources and reserves presented in Table 8.1 are a historical estimate. They were not prepared to NI 43-101 standards, are not current and should not be relied upon. For the purpose of this report, the historic resources and reserves reported below have been converted to Inferred resources (Table 8.1). Table 8.1: Bacis historic resources for the Cozamin mine.

Resource Category Tonnes Au (g/t) Ag (g/t) Cu (%) Pb (%) Zn (%) Inferred 4,040,084 0.37 93.19 1.65 0.50 1.87

National Instrument 43-101 compliant resources were estimated for Capstone in 2005. Resources were updated in 2006, and in 2007, both the resources and reserves have been updated (see section 19 of this report). 9 GEOLOGICAL SETTING 9.1 Regional 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; Figure 8.1). The district is in the southern Sierra Madre Occidental Physiographic Province near the boundary with the Mesa Central Physiographic Province in north-central Mexico (Figure 6.1). 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 9.1). 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). 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.



Figure 9.1: Geology of the San Roberto mine area.



The host rocks for the Mala Noche vein are intercalated carbonaceous metasedimentary rocks and andesitic volcanic rocks ranging in age from Triassic to Cretaceous, and Tertiary rhyolite intrusive rocks and flows (Figure 9.1). 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 weak 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 mineralization in the San Roberto mine at Cozamin is relatively close to the intrusive centre for the entire district. 9.2 Property Geology The dominant mineralized vein on the Cozamin project is the Mala Noche (Figures 8.1 and 9.1). 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. Production from the Mala Noche commenced in 1548. There are at least 18 shafts that provide access to the historical workings at Cozamin (Figure 6.3). 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. Structural Geology The Mala Noche vein system occupies a system of anastamozing 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. Alteration Moderate propyllitic wall rock alteration is generally limited to 3m 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 druzy quartz, coarse crystalline masses, and a stockwork of quartz veinlets. 10 DEPOSIT TYPES 10.1 Mineral Deposit Type Mineralization in the Mala Noche vein at Cozamin appears to have been episodic. Early epithermal mineralization and alteration (represented by sulfide pseudomorphs



of carbonates and possibly barite) have been overprinted by higher temperature pyrite and pyrrhotite dominant mineralization in a telescoped, intrusive related hydrothermal system. Zinc sulfides dominate in the upper historic workings at the San Roberto mine where historic resources averaged 3.16% zinc, 0.95% copper and 85 g/t silver. At intermediate and deeper levels in the San Roberto mine, copper grades typically exceed 2% and associated silver grades are frequently above 85 g/t. Copper-silver mineralization, that is often associated with high pyrrhotite concentrations, extends at least 350m below the historic workings. This mesothermal mineralization has potential to extend to depth. 10.2 Geological Concept Used For Exploration of the Property 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 100m 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 3 drill sections, comprised of 2 drill holes each, all intersected significant economic mineralization over true widths varying from 3.2m to 14.9m. These 3 drill sections were distributed over 550m of strike extent beneath the historic workings. At that point, Capstone management decided to drill single hole sections every 100m beneath the San Roberto workings. These holes targeted the Mala Noche at approximately 2,150 metres above sea level (masl) which is 65m 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 6.64m in true width and had weighted grade averages of 2.61% copper, 91.25 g/t silver and 1.38% zinc. These significantly higher copper grades and undiminished silver grades are associated with significant amounts of pyrrhotite. This reinforced management’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 350m below the historic workings. 11 MINERALIZATION The Mala Noche vein in the San Roberto mine workings shows contained sulfides 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 sulfide 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 sulfide 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 sulfide 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 most common economic sulfide mineral and occurs 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. Mala Noche vein mineralization at the San Roberto mine is interpreted to be episodic with early epithermal mineralization replaced by higher temperature mineralization in a telescoped, hydrothermal system which is probably related to a buried intrusive body. Lead and zinc mineralization dominate the upper historic workings whereas copper dominant mineralization extends over 350m below historic workings.



12 EXPLORATION Exploration completed to date at the Cozamin project has focused on defining the copper-silver-lead-zinc mineralization associated with the Mala Noche vein in and around the historic workings of the San Roberto mine. Five phases of drilling (approximately 61,970m) have been completed and define resources down to approximately 500m depth and locally down to approximately 800m below the surface. Only 7 of the 223 drill holes have tested the mineralization beyond a distance of 200m along strike from the current mine development. Exploration in the mine area has been successful in adding to and upgrading the resources reported in the 2005 pre-mining, NI 43-101 compliant resource estimate (see section 19). Results to date indicate that the copper-silver-lead-zinc mineralization associated with the Mala Noche vein in the San Roberto mine area is continuous along strike and extends to at least 700m depth. Below is a summary of the completed exploration phases. 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 7,484.44m surface NQ-diameter diamond drill program (holes CG-04-01 to CG-04-19, Figure 12.1).

• 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 10,483.27m surface NQ-diameter diamond drill program (holes

CG-04-20 to CG-04-37, Figure 12.2) that mainly tested the Mala Noche vein at elevations between the 1,900m and 2,050m 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 17,687.70m underground definition NQ-diameter diamond drill program (holes CG-U01 to CG-U114, Figure 12.3).

• 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 4,824.56m surface PQ/NQ-diameter diamond drill program (holes CG-06-38 to CG-06-39 and CG-07-40 to CG-07-42, Figure 12.2) that tested the Mala Noche vein at elevations between approximately 600-700m below surface of the San Roberto mine.

• Completed a 21,441.10m underground NQ-diameter diamond drill program (holes CG-06-U115 to CG-06-U124, and CG-07-U125 to CG-07-U183, Figure 12.2). These holes were designed to infill and extend the 2006 estimated resources.

• Significant results of these drill programs are presented in section 13.



Figure 12.1: Map showing the location of the Mala Noche vein and Phase I diamond drill holes on the Cozamin project.



Figure 12.2: Map showing the location of the Mala Noche vein and Phase II diamond drill holes on the Cozamin project.



Figure 12.3: West-east long section showing drill hole intercepts from the Phase III underground diamond drilling program on the Cozamin project.



Figure 12.4: West-east long section showing drill hole intercepts (blue) from the Phase IV and V diamond drill hole programs on the Cozamin project.



13 DRILLING Fourty two surface and 183 underground exploration holes have been drilled at Cozamin in five phases of exploration from April 2004 through July 2007. Drill hole orientiations are generally perpendicular to the strike of the overall structural trend of the Mala Noche vein, the main exception being the deeper surface holes. Table 13.1 summarizes the drilling contractors used at Cozamin. Table 13.1: Drilling companies used at Cozamin. SURFACE Britton Brothers 2004-2005 PHASE I 37 holes 17,967.71m Eastman Single Shot

Major Drilling 2006-2007 PHASE V 5 holes 4,824.56m FLEXIT SensIT UNDERGROUND Canrock Drilling 2005-2006 PHASE II 78 holes 10,031.24m Reflex EZ-Shot

Explor 2005 PHASE II 1 hole 305.70m Reflex EZ-Shot Tecmin 2005-2006 PHASE III 35 holes 7,417.37m Reflex EZ-Shot Tecmin 2006-2007 PHASE IV 69 holes 21,441.10m Reflex EZ-Shot

Drill holes are located using a total stations TOPCON instrument, model GTS-236W. Down hole survey readings were recorded on average every 50m using either an Eastman Single Shot, FLEXIT SensIT or Reflex EZ-Shot instrument as indicated in Table 13.1. Survey readings are generally taken every 50-150m for surface holes and every 50-100m 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, however the average variation is 1.3º. The higher variations generally occur in the interval between the planned drill hole collar and the first reading. Descriptions of each drilling phase completed on the Cozamin project and a summary of significant intersections is presented below. Phases I-III The Phase I and II surface exploration drill programs totalled 17,967.71m 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 17,736.31m of NQ-diameter diamond core that were drilled in 2005 and the first half of 2006. Significant intersections are presented in Tables 13.2 and 13.3. 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 13.2: Results from the Phase I and II surface drilling.

Hole # From (m) To (m)

Length (m)

True Width (m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

CG-04-01 287.50 296.00 8.50 3.20 0.03 126.59 1.93 1.06 2.85CG-04-02 306.80 311.95 5.15 4.20 0.22 129.12 2.89 0.12 0.73CG-04-03 326.20 341.40 15.20 14.90 0.08 99.63 4.35 0.01 0.27CG-04-04 340.00 349.75 9.75 7.24 40.90 2.12 2.60including 346.75 349.00 2.25 2.10 67.30 3.42 4.13CG-04-05 351.00 356.75 5.75 5.47 62.90 1.06 0.85including 353.87 356.25 2.38 2.38 135.30 2.46 1.61CG-04-06 417.40 426.30 8.90 6.85 62.20 1.49 0.23including 417.40 419.30 1.90 1.77 185.40 2.91 1.17CG-04-07 377.00 380.00 3.00 2.65 82.10 0.85 13.93including 377.00 379.00 2.00 1.76 119.10 1.19 10.45CG-04-08 336.80 340.00 3.20 2.80 97.80 1.60 1.60CG-04-10 483.30 496.20 12.90 9.00 107.60 2.50 1.90including 488.10 495.30 7.20 5.00 148.10 3.40 2.10CG-04-11 367.60 373.90 6.30 5.60 126.10 2.50 2.60including 369.70 372.00 2.30 2.00 258.80 4.30 2.40CG-04-12 291.50 296.80 5.30 4.00 115.90 2.90 0.90including 291.50 295.00 3.50 2.70 153.60 3.70 1.20CG-04-13 321.80 333.50 11.70 8.90 134.00 1.40 1.60including 330.40 333.50 3.10 2.40 279.10 1.80 1.40CG-04-15 294.90 302.80 7.90 6.00 75.50 4.90 0.20including 296.60 299.10 2.50 1.90 133.60 8.70 0.80CG-04-16 317.40 321.90 4.50 3.90 62.20 1.50 0.50CG-04-18 344.40 361.40 17.00 14.10 82.20 3.90 0.40including 347.20 354.90 7.70 6.40 145.60 7.00 0.40CG-04-19 190.40 198.30 7.90 6.00 58.20 1.60 0.40including 191.80 194.10 2.30 1.70 106.00 4.40 0.50CG-04-20 312.70 321.00 8.30 6.80 100.30 1.40 4.80including 317.60 320.70 3.10 2.60 138.70 1.90 2.70CG-04-21 464.60 476.80 12.20 8.50 11.10 0.20 5.90CG-04-22 450.70 452.30 1.60 1.40 30.20 1.70 0.10CG-04-23 135.00 135.70 0.70 0.60 30.50 0.00 3.80CG-04-24 417.00 417.90 0.90 0.80 18.20 0.10 9.40CG-04-25 267.80 269.40 1.60 1.40 0.20 0.20 4.50CG-04-26 647.00 653.20 6.20 4.60 32.00 2.00 0.10including 647.00 649.60 2.60 1.90 41.40 2.80 0.10CG-04-27 589.80 593.70 3.90 5.90 62.10 1.80 0.80including 585.80 593.70 7.90 2.90 91.60 2.50 1.40CG-04-28 639.30 644.40 5.10 3.80 40.80 2.00 0.10including 642.00 644.40 2.40 1.80 62.30 3.00 0.20CG-04-29 510.20 512.90 2.70 3.20 158.90 5.50 1.10including 510.50 513.90 3.40 2.10 219.50 7.60 0.90CG-05-33 666.20 668.20 2.00 1.70 14.00 0.20 0.00 8.00CG-05-34 104.10 111.00 6.90 6.10 37.00 0.00 3.00 6.90and 565.90 572.50 6.60 6.00 40.00 1.70 0.00 0.10




Length (m)

True Width (m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

CG-04-01 287.50 296.00 8.50 3.20 0.03 126.59 1.93 1.06 2.85CG-04-02 306.80 311.95 5.15 4.20 0.22 129.12 2.89 0.12 0.73CG-04-03 326.20 341.40 15.20 14.90 0.08 99.63 4.35 0.01 0.27CG-04-04 340.00 349.75 9.75 7.24 40.90 2.12 2.60including 346.75 349.00 2.25 2.10 67.30 3.42 4.13CG-04-05 351.00 356.75 5.75 5.47 62.90 1.06 0.85including 353.87 356.25 2.38 2.38 135.30 2.46 1.61CG-04-06 417.40 426.30 8.90 6.85 62.20 1.49 0.23including 417.40 419.30 1.90 1.77 185.40 2.91 1.17CG-04-07 377.00 380.00 3.00 2.65 82.10 0.85 13.93including 377.00 379.00 2.00 1.76 119.10 1.19 10.45CG-04-08 336.80 340.00 3.20 2.80 97.80 1.60 1.60CG-04-10 483.30 496.20 12.90 9.00 107.60 2.50 1.90including 488.10 495.30 7.20 5.00 148.10 3.40 2.10CG-04-11 367.60 373.90 6.30 5.60 126.10 2.50 2.60including 369.70 372.00 2.30 2.00 258.80 4.30 2.40CG-04-12 291.50 296.80 5.30 4.00 115.90 2.90 0.90including 291.50 295.00 3.50 2.70 153.60 3.70 1.20CG-04-13 321.80 333.50 11.70 8.90 134.00 1.40 1.60including 330.40 333.50 3.10 2.40 279.10 1.80 1.40CG-04-15 294.90 302.80 7.90 6.00 75.50 4.90 0.20including 296.60 299.10 2.50 1.90 133.60 8.70 0.80CG-04-16 317.40 321.90 4.50 3.90 62.20 1.50 0.50CG-04-18 344.40 361.40 17.00 14.10 82.20 3.90 0.40including 347.20 354.90 7.70 6.40 145.60 7.00 0.40CG-04-19 190.40 198.30 7.90 6.00 58.20 1.60 0.40including 191.80 194.10 2.30 1.70 106.00 4.40 0.50CG-05-35 252.60 254.30 1.70 1.50 79.80 0.90 0.00 3.70



Table 13.3: Results from the Phase III underground drilling.


Length (m)

True Width (m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

CG-05-U02 43.80 50.00 6.20 6.00 95.80 2.20 0.30 0.70CG-05-U03 54.80 61.90 7.10 5.50 95.60 1.60 0.20 0.70CG-05-U04 62.00 68.40 6.40 3.90 82.40 1.10 1.60 0.80CG-05-U05 42.10 55.00 12.90 10.90 110.80 1.30 1.90 1.00CG-05-U06 49.30 58.30 9.00 6.70 122.20 1.70 1.60 1.10CG-05-U07 32.00 35.70 3.70 3.40 105.90 0.60 8.40 7.50and 46.30 51.00 4.70 4.30 63.20 1.50 0.20 0.60CG-05-U08 40.50 43.10 2.60 1.90 144.60 1.20 12.80 9.20and 50.00 53.70 3.70 2.70 56.40 1.20 0.10 0.90CG-05-U09 50.80 52.00 1.20 1.20 90.30 2.50 0.00 0.00CG-05-U10 49.80 57.30 7.50 6.80 67.20 2.20 0.10 0.50CG-05-U11 52.90 66.70 13.80 10.60 59.20 2.10 0.00 0.10CG-05-U12 57.50 63.30 5.80 3.80 23.00 1.70 0.00 0.10CG-05-U13 30.80 32.70 1.90 1.50 6.10 0.00 0.00 3.20CG-05-U17 40.60 52.40 11.80 8.70 132.70 1.30 1.60 0.50CG-05-U18 39.70 53.00 13.30 9.70 94.10 1.30 0.50 0.60CG-05-U19 39.80 62.10 22.30 18.50 104.70 1.30 3.50 1.50CG-05-U20 45.90 50.60 4.70 3.20 233.00 1.50 3.90 0.90and 66.00 75.60 9.60 6.50 139.50 1.90 0.60 0.40CG-05-U21 50.50 55.20 4.70 2.90 122.90 1.50 0.50 2.10CG-05-U22 68.70 73.30 4.60 1.80 31.40 1.50 0.00 0.10CG-05-U25 55.80 63.30 7.50 3.50 90.90 1.30 0.20 2.50CG-05-U26 51.40 54.40 3.00 2.30 73.50 1.40 0.00 4.30CG-05-U27 42.30 49.30 7.00 6.90 82.80 1.40 0.10 1.70CG-05-U29 57.90 61.30 3.40 2.60 105.20 1.60 0.60 1.70CG-05-U30 63.80 69.50 5.70 3.70 106.20 2.10 0.20 2.40CG-05-U31 52.50 61.50 9.00 6.80 51.50 0.60 0.60 4.70CG-05-U32 86.00 102.00 16.00 2.30 134.70 2.80 0.60 3.40CG-05-U33 65.00 70.00 5.00 2.80 82.70 0.90 0.60 3.00CG-05-U34 115.00 132.50 17.50 14.80 138.00 2.50 0.20 1.00CG-05-U35 144.00 165.50 21.50 14.60 88.90 2.20 0.60 0.40CG-05-U36 116.50 127.50 11.00 7.10 142.60 4.20 0.90 3.10CG-05-U37 161.00 167.50 6.50 2.40 36.20 2.10 0.00 1.10CG-05-U38 131.80 143.00 11.20 9.00 90.70 2.30 2.30 1.30CG-05-U39 155.00 175.90 20.90 7.80 96.10 2.80 0.20 1.30CG-05-U40 143.30 149.30 6.00 3.80 75.90 2.40 0.10 2.90CG-05-U42 123.50 132.30 8.80 6.70 52.50 1.50 0.60 1.50CG-05-U43 129.30 134.30 5.00 3.10 81.20 2.70 0.10 0.50CG-05-U44 110.00 115.30 5.30 3.60 58.80 2.10 0.10CG-05-U45 114.00 127.00 13.00 10.50 106.60 2.90 0.30 1.30including 114.50 120.50 6.00 4.90 151.70 3.40 0.70 2.40CG-05-U46 113.50 128.80 15.30 12.00 193.20 3.40 1.80 1.00including 120.80 126.80 6.00 4.70 360.90 5.90 3.20 0.50CG-05-U47 104.00 112.80 8.80 7.20 46.40 1.80 0.10including 108.30 112.30 4.00 3.30 78.00 3.10 0.10



CG-05-U48 141.30 155.00 13.70 11.30 93.40 2.00 0.90 0.70CG-05-U50 108.50 123.50 15.00 10.40 84.10 1.90 0.80including 116.00 123.50 7.50 5.20 93.90 3.20 0.40CG-05-U51 93.80 96.80 3.00 2.80 54.30 1.30 0.10 0.10CG-05-U52 115.50 122.00 6.50 5.50 34.10 1.00 0.20 5.10CG-05-U53 170.00 179.80 9.80 2.30 98.70 2.00 0.90 1.60CG-05-U54 169.00 184.50 15.50 7.70 70.60 3.90 0.10including 177.50 184.00 6.50 3.20 114.40 6.70 0.20CG-05-U55 77.50 80.50 3.00 2.00 47.50 0.30 0.90 3.00CG-05-U56 54.00 59.00 5.00 4.80 100.20 1.30 0.40 3.50CG-05-U58 67.30 79.00 11.70 7.90 96.30 2.40 0.20 2.90including 74.50 79.00 4.50 3.00 175.40 4.60 0.10CG-05-U59 60.50 64.00 3.50 2.70 54.50 1.20 0.40 4.50CG-05-U60 75.50 79.30 3.80 2.00 145.50 1.60 4.90 5.20CG-05-U61 134.80 138.00 3.20 1.50 212.70 1.80 19.00 3.60CG-05-U61 181.50 193.00 11.50 5.30 40.50 2.40 0.20including 181.50 193.00 11.50 1.80 83.10 4.80 0.20CG-05-U62 122.00 147.00 25.00 16.20 90.20 3.30 0.20 0.80including 127.50 137.00 9.50 6.10 123.00 5.00 0.40CG-05-U64 126.00 135.50 9.50 7.50 21.70 0.60 5.90including 129.50 132.00 2.50 2.00 48.40 1.60 0.60CG-05-U67 142.30 148.50 6.20 4.20 54.20 1.80 0.30including 144.50 148.50 4.00 2.70 80.70 2.70 0.40CG-05-U68 124.00 130.00 6.00 3.65 108.10 2.90 0.07 1.30CG-05-U69 175.50 185.00 9.50 2.90 37.00 2.10 0.40CG-05-U70 116.00 123.50 7.50 6.80 113.60 4.50 0.20CG-05-U71 170.50 180.50 10.00 5.00 78.90 4.20 0.60including 170.50 175.50 5.00 2.50 145.20 7.70 0.70CG-05-U72 167.50 174.50 7.00 3.80 190.80 5.40 1.40including 171.00 174.50 3.50 1.90 219.60 8.50 0.50CG-05-U73 99.00 126.00 27.00 19.70 60.70 2.00 0.30including 102.00 106.70 4.70 3.40 141.90 2.80 1.00CG-05-U74 50.50 54.50 4.00 4.00 104.90 2.10 2.70CG-05-U75 155.50 163.50 8.00 5.60 147.70 8.30 0.20 0.30CG-05-U76 72.50 91.50 19.00 13.70 88.40 2.70 0.80including 84.00 91.00 7.00 5.10 115.10 4.00 0.10CG-05-U77 252.80 257.00 4.20 1.60 46.10 1.10 0.10 1.70CG-05-U78 95.00 107.00 12.00 11.20 91.30 3.30 0.40 1.70CG-05-U79 162.30 165.50 3.20 2.30 86.90 2.10 0.10 2.80CG-05-U80 198.00 214.50 16.50 7.80 60.60 3.10 0.10 0.90including 205.00 214.50 9.50 4.50 83.00 4.60 0.10 1.00CG-05-U81 107.00 110.80 3.80 2.60 81.10 2.10 0.10 0.30and 129.30 135.00 5.70 3.90 81.30 2.00 0.90 0.10CG-05-U82 188.30 197.00 8.70 4.48 80.80 2.90 0.10 0.60CG-05-U83 175.00 189.00 14.00 6.54 93.00 3.30 0.10 0.70including 176.50 182.00 5.50 2.57 159.70 6.70 0.03 0.30CG-05-U84 147.50 155.80 8.30 5.60 86.50 4.00 0.40 2.20including 148.00 150.50 2.50 1.70 185.60 9.50 0.10 0.30CG-05-U85 176.80 185.00 8.20 3.60 78.20 2.10 0.10 1.50



including 181.00 185.00 4.00 1.70 110.40 3.40 0.10 1.90CG-05-U86 187.50 192.00 4.50 1.70 56.30 1.80 0.10 0.20CG-05-U87 159.00 169.00 10.00 4.50 84.80 2.10 0.20 1.90including 162.50 166.50 4.00 1.80 137.30 3.80 0.40 2.20CG-05-U88 181.50 193.00 11.50 5.40 93.30 2.80 0.10 1.60including 184.80 189.00 4.20 2.00 173.60 4.80 0.20 1.30CG-05-U91 166.00 170.25 4.25 1.51 60.00 2.80 0.02 0.50CG-05-U93 58.00 70.50 12.50 6.65 116.20 1.10 0.50 1.70including 64.00 69.50 5.50 2.93 185.30 1.30 0.70 2.40CG-05-U94 286.00 303.50 17.50 9.80 114.70 6.00 0.10 1.10including 291.50 303.00 11.50 6.50 141.60 7.70 0.10 0.60CG-05-U95 256.00 261.00 5.00 2.82 73.80 2.70 0.20 0.40CG-05-U96 248.50 261.00 12.50 6.30 68.60 3.00 0.06 1.10including 254.00 257.50 3.50 1.76 128.20 5.90 0.05 1.60CG-05-U97 235.00 241.25 6.25 3.34 31.40 1.10 0.03 0.30CG-05-U98 270.50 279.25 8.75 3.19 128.80 4.40 0.60 2.50including 271.00 274.50 3.50 1.27 196.90 7.40 0.10 0.50CG-05-U99 204.50 218.50 14.00 9.52 73.70 3.00 0.10 0.50including 212.00 216.50 4.50 3.06 136.20 5.50 0.04 1.20CG-05-U100 233.00 238.00 5.00 2.45 136.90 3.70 3.20 0.90CG-05-U101 226.50 251.50 25.00 7.88 114.90 4.70 0.40 1.00including 232.25 239.00 6.75 2.13 207.00 9.30 0.40 1.20CG-05-U102 269.50 278.00 8.50 3.03 82.40 2.80 0.05 1.20CG-05-U103 139.00 149.50 10.50 5.40 245.80 2.80 9.00 3.80and 175.25 182.50 7.25 3.73 113.40 4.70 0.06 0.30CG-05-U104 198.50 202.00 3.50 1.21 111.30 1.80 2.90 7.20CG-05-U105 154.00 165.00 11.00 4.38 208.80 3.30 0.20 3.50and 165.00 175.00 10.00 3.98 7.60 0.20 0.02 10.00CG-05-U106 304.00 314.50 10.50 4.46 106.80 4.50 0.20 1.20CG-05-U107 123.50 132.75 9.25 7.90 103.70 1.70 4.70 1.00CG-05-U108 140.00 156.25 16.25 7.07 95.80 3.30 1.50 0.30including 147.50 155.50 8.00 3.48 135.50 4.60 0.02 0.20CG-05-U109 358.00 364.00 6.00 2.07 60.60 2.60 0.03 1.80CG-05-U110 126.00 133.00 7.00 4.62 74.70 2.30 1.30 0.70CG-05-U111 135.50 147.00 11.50 4.66 71.20 2.70 0.01 0.20including 139.75 147.00 7.25 2.94 85.00 3.80 0.20 0.30CG-05-U112 319.50 323.00 3.50 1.90 118.90 5.50 0.10 0.60CG-05-U113 343.30 349.00 5.70 1.89 78.90 4.00 0.02 1.10CG-05-U114 282.00 294.00 12.00 7.44 34.70 1.70 0.01 0.20including 291.00 294.00 3.00 1.86 101.30 4.70 0.03 0.20

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. 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 holes were drilled with PQ-diameter pre-collars to approximately 300m depth and completed with NQ-tails. A total of 4,824.56m were drilled and 304 samples assayed for copper, silver, lead, zinc and gold at ALSChemex in Vancouver. Underground NQ-diameter drilling commenced in November 2006 and was completed in July 2007. Sixty nine holes were drilled for a total of 21,441.10m. Two thousand, two hundred and seventy seven samples assayed for copper, silver, lead, zinc and gold at ALSChemex in Vancouver.

Significant intersections are presented in Tables 13.4 and 13.5. 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 13.4: Results from the Phase IV surface drill program.

Hole # From (m)

To (m)

Length (m)

True Width (m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

CG-06-38 768.75 773.50 4.75 3.88 0.07 53.50 1.36 0.03 0.06 CG-07-41 793.00 799.50 6.50 2.89 0.00 49.54 0.13 0.04 0.04

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



Figure 13.4: West-east long section showing drill hole intercepts (blue) from the Phase IV and V diamond drill hole programs on the Cozamin project.



Table 13.5: Results from the Phase V underground drill program.


Length (m)

True Width (m)

Au (g/t)

Ag (g/t)

Cu (%)

Pb (%)

Zn (%)

CG-06-U115 151.00 152.00 1.00 0.69 0.56 98.00 1.15 0.19 1.09 CG-06-U117 56.50 58.75 2.25 1.60 0.17 78.11 1.67 0.07 3.14 CG-06-U118 399.45 416.30 16.85 10.35 0.01 17.96 0.55 0.01 0.61

including 412.50 416.30 3.80 2.33 0.04 60.37 1.58 0.03 2.23 CG-06-U119 392.00 395.75 3.75 1.71 0.04 27.33 2.12 0.01 0.46 CG-06-U120 310.50 329.50 19.00 14.30 0.00 53.87 3.62 0.04 0.59 CG-06-U121 317.50 323.00 5.50 3.79 0.00 104.95 3.91 0.34 3.14 CG-06-U122 409.00 417.00 8.00 2.80 0.00 53.93 3.84 0.01 0.13 CG-06-U123 390.15 393.00 2.85 1.23 0.14 56.14 2.83 0.01 0.30 CG-06-U124 304.20 309.50 5.30 3.18 0.02 125.75 3.83 0.22 3.42 CG-06-U125 288.75 294.00 5.25 3.63 0.03 140.10 3.17 0.37 2.93 CG-07-U126 201.00 203.80 2.80 2.35 0.16 32.20 1.18 0.07 0.47 CG-07-U127 249.50 253.85 4.35 3.60 0.09 82.07 5.60 0.03 1.26 CG-07-U129 286.00 294.50 8.50 6.40 0.12 66.29 1.67 0.09 3.39

including 286.00 291.00 5.00 3.77 0.19 98.60 2.65 0.11 3.30 CG-07-U130 333.00 334.00 1.00 0.77 0.00 55.00 1.98 0.02 2.40 CG-07-U131 187.70 193.20 5.50 4.26 0.13 116.21 2.15 0.07 0.52

including 188.20 192.50 4.30 3.33 0.11 153.17 2.87 0.09 0.52 CG-07-U132 325.50 344.50 19.00 9.95 0.00 18.13 0.49 0.03 3.00

including 327.25 333.50 6.25 3.27 0.00 32.32 1.05 0.06 3.34 CG-07-U134 229.69 260.50 30.81 18.25 0.02 19.40 0.65 0.06 0.96

including 232.50 237.00 4.50 2.67 0.03 26.78 1.03 0.02 0.16 and including 242.00 260.50 18.50 10.96 0.00 17.81 0.63 0.06 1.20 CG-07-U136 385.00 387.50 2.50 0.92 0.00 4.16 0.18 0.02 0.78 CG-07-U138 346.00 348.00 2.00 1.16 0.00 7.00 0.64 0.01 0.03 CG-07-U139 188.50 193.00 4.50 4.05 0.17 33.89 0.37 0.20 3.36

including 191.00 192.00 1.00 0.90 0.10 87.50 0.94 0.33 4.73 and 219.50 222.25 2.75 2.48 0.04 75.00 3.68 0.02 0.18

CG-07-U141 400.75 403.00 2.25 0.84 0.00 32.22 2.52 0.01 0.10 CG-07-U142 212.00 218.65 6.65 4.62 0.01 3.42 0.03 0.03 1.26 CG-07-U144 196.50 204.00 7.50 5.49 0.02 8.00 0.13 0.03 2.10

and 243.00 252.00 9.00 6.59 0.01 37.94 2.38 0.01 0.56 including 243.25 247.00 3.75 2.75 0.02 79.33 4.87 0.01 0.14

CG-07-U145 257.00 260.50 3.50 3.18 0.00 65.79 3.54 0.04 1.27 including 258.00 259.50 1.50 1.36 0.00 129.29 6.91 0.07 0.31

CG-07-U146 179.00 190.50 11.50 10.16 0.01 3.00 0.09 0.03 1.17 including 180.50 181.00 0.50 0.44 0.00 30.00 1.37 0.03 0.14

CG-07-U147 355.25 359.00 3.75 2.35 0.26 351.27 3.81 0.29 2.77 including 355.50 358.00 2.50 1.57 0.37 520.00 5.64 0.42 3.18

CG-07-U148 397.00 403.00 6.00 2.18 0.01 22.38 1.64 0.02 0.08 CG-07-U149 195.50 199.80 4.30 2.74 0.03 50.77 0.46 0.83 2.53

and 203.00 205.75 2.75 1.75 0.01 433.00 2.26 10.80 5.02 including 204.00 205.50 1.50 0.96 0.00 776.67 3.98 18.99 7.79

and 246.00 254.00 8.00 5.10 0.02 13.41 0.81 0.01 0.75 including 246.00 249.55 3.55 2.26 0.05 30.21 1.78 0.01 0.51



CG-07-U150 337.00 343.00 6.00 4.66 0.12 27.67 0.63 0.03 0.96 including 339.00 340.50 1.50 1.17 0.00 76.33 1.73 0.04 0.28

CG-07-U151 193.00 199.50 6.50 4.16 0.07 50.69 0.64 1.97 2.43 and 220.00 229.00 9.00 5.76 0.02 17.53 0.19 1.49 1.47 and 232.95 238.30 5.35 3.43 0.13 289.84 4.89 1.90 0.83 and 242.50 252.00 9.50 6.08 0.01 126.96 5.21 0.48 2.02

including 245.10 251.50 6.40 4.10 0.01 179.17 7.57 0.25 1.66 CG-07-U152 350.00 363.00 13.00 6.40 0.01 24.56 1.64 0.01 0.24

including 351.25 355.00 3.75 1.85 0.00 54.33 4.19 0.02 0.16 CG-07-U153 290.25 294.75 4.50 2.59 0.00 76.72 4.24 0.02 0.19

and 299.50 313.00 13.50 7.78 0.03 69.70 4.53 0.04 0.29 CG-07-U154 159.85 164.50 4.65 3.99 0.00 27.41 0.27 0.07 1.40

including 159.85 161.75 1.90 1.63 0.00 67.08 0.65 0.16 1.18 CG-07-U155 133.50 138.00 4.50 4.47 0.29 88.56 1.23 0.67 1.53

including 134.50 138.00 3.50 3.48 0.38 113.86 1.57 0.86 1.74 CG-07-U156 145.80 150.20 4.40 3.99 0.00 79.21 1.75 0.10 1.06



and 364.50 365.30 0.80 0.46 0.00 119.50 4.18 0.02 0.14 CG-07-U161 173.35 177.00 3.65 2.38 0.18 155.81 4.16 0.50 1.69 CG-07-U163 359.00 361.50 2.50 1.58 0.00 0.00 0.12 0.01 8.21 CG-07-U165 117.00 122.50 5.50 2.23 0.08 71.77 1.09 0.09 0.91

and 258.50 261.30 2.80 1.13 0.00 23.04 0.87 0.03 0.74 including 258.50 260.50 2.00 0.81 0.00 32.25 1.19 0.03 0.37

CG-07-U166 173.00 177.25 4.25 2.34 0.01 50.00 1.99 0.01 2.04 including 173.50 177.00 3.50 1.93 0.00 55.86 2.29 0.01 2.45

CG-07-U167 187.00 203.50 16.50 14.10 0.04 74.90 3.18 0.05 0.71 including 191.00 203.50 12.50 10.68 0.03 93.24 4.12 0.05 0.48

CG-07-U168 285.65 288.50 2.85 2.14 0.16 19.54 0.16 0.02 2.55 CG-07-U169 267.00 270.50 3.50 1.30 0.00 70.70 2.46 0.08 0.33 CG-07-U170 192.15 196.50 4.35 2.10 0.00 40.90 1.57 0.06 0.44 CG-07-U171 300.00 310.00 10.00 9.01 0.00 31.53 1.23 0.53 3.34

including 303.50 307.00 3.50 3.15 0.00 74.64 3.06 1.48 4.81 CG-07-U172 364.50 365.00 0.50 0.36 0.00 71.00 2.24 0.10 5.40 CG-07-U173 220.25 224.00 3.75 2.90 0.00 30.07 0.81 0.03 1.81 CG-07-U174 265.10 267.00 1.90 1.54 0.24 153.95 4.68 0.02 0.60 CG-07-U175 322.00 323.70 1.70 0.82 0.07 121.94 0.22 4.75 2.91 CG-07-U176 202.85 215.00 12.15 11.28 0.01 43.94 1.82 0.36 1.29

including 202.85 207.00 4.15 3.85 0.04 97.37 3.73 1.04 1.65 and 217.90 224.50 6.60 6.13 0.04 26.80 0.06 3.13 2.35

CG-07-U177 241.00 243.00 2.00 1.56 0.08 134.03 1.03 4.03 6.24 and 289.00 292.75 3.75 2.93 0.01 23.04 1.22 0.03 0.22

CG-07-U178 195.35 200.70 5.35 4.49 0.15 338.81 6.59 0.44 0.67 CG-07-U179 172.45 177.50 5.05 3.45 0.01 105.90 1.37 0.17 0.88



and 191.00 193.50 2.50 1.71 0.02 39.00 1.41 0.02 2.47 and 204.00 205.00 1.00 0.68 0.00 25.50 1.01 0.02 0.66

CG-07-U180 221.50 223.00 1.50 0.62 0.03 6.00 0.13 0.02 2.64 and 225.00 228.25 3.25 1.33 0.01 93.59 1.55 0.03 1.25


including 176.85 179.25 2.40 1.93 0.00 67.44 1.18 0.03 7.67 CG-07-U183 346.00 355.40 9.40 8.27 0.04 47.88 1.07 0.18 1.11

including 354.00 355.40 1.40 1.23 0.06 212.57 4.53 1.00 3.82 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. Drill holes CG-06-U116, CG-06-U128, CG-07-U133, CG-07-U135, CG-07-U140, CG-07-U143, CG-07-U162 and CG-07-U 164 were anomalous but did not intersect any significant mineralization. 14 SAMPLE METHOD AND APPROACH The results of two sampling methods are presented in this report: drill core cutting and underground chip sampling. 14.1 Diamond Drill Core Sampling A total of 8,294 drill core samples have been collected from the 37 surface and 114 underground diamond drill holes in the Phase I-III drilling programs. Phase IV and V samples total 2,581 from 5 surface (304 samples) and 69 underground holes (2,277). These samples were collected from the Mala Noche vein and surrounding rock over a length of approximately 1.5 km and a 925m vertical extent. Samples selected for assay from each hole consisted of the Mala Noche vein, sulfide-bearing hangingwall and footwall quartz veins, and non-mineralized rock surrounding these intervals. Sample intervals vary from 0.11m to a maximum of 3m in the surface exploration holes, and from 0.2m to 3.05m in the underground drill holes. Sample intervals in mineralized zones do not exceed 1m and are typically 0.5m in length. Significant intersections corrected for true width are presented in section 13. The following are relevant excerpts from Capstone’s work procedure for drill core handling and sampling: DRILL SITE CONTROL

• Core boxes should be delivered to the drill site in such a manner that they are clean and free of grease at the site. The drill hole number should be marked clearly on the box by the driller.

• The driller must place 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.



• The drill site geologist should check the order of the core and make a very quick log of the rock in the core box prior to transportation.

• A clean top must be placed on each box to be transported and be 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 must monitor the transportation of the core from the drill

site to the core shack. Transportation should be as gentle as reasonably possible.

CORE SHACK CONTROL

• When the core arrives at the core shack, a geologist should oversee the placement of the core boxes on the logging tables and the removal of the tops.

• The core should be cleaned and wet. The boxes should be clearly marked for

hole number and start and finish depths prior to photographing and logging for rock quality.

• The core should be logged for 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 should be marked by the geologist. The saw line for primary samples should be marked in the centre of the core with each side being roughly equivalent for apparent grade.

• The drill hole number and sample interval have to be clearly entered in the

sample book. One ticket stub should be stapled in the corresponding interval in the core box by the geologist and the other two ticket stubs should be placed in the sample bag by the sampler. The sample books must be archived at Cozamin.

• The sample interval should not exceed 0.5m in the vein and 2m in the wallrock.

Very high grade intervals should be marked out and sampled separately from lower grade zones. The boundary should be based on mineral proportions and/or texture (e.g. massive versus disseminated). However, the sample interval should not be less than 0.25m 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.



SURVEY CONTROL • The locations and orientations of the drill holes must be checked by the

surveyor prior to drilling and after the completion of each hole. The driller must identify each hole with a wood plug showing the drill hole number clearly labelled with permanent black marker. Drill hole locations should be surveyed using the total stations TOPCON surveying instrument.

• Down hole orientation surveys must be undertaken on completion of each hole.

A minimum of two survey points must be collected on shorter holes. On longer holes, survey points should be taken approximately every 75m starting at the bottom of the hole. Drill hole surveys are to be completed using the Reflex EZ-shot instrument.

14.2 Underground Chip Sampling A total of 8,057 underground chip samples were used in preparing the updated resource estimate presented in section 19 of this report. These samples were collected from Levels 7, 8, 9, 9.3, 9.6, 10 and 10.3 in the San Roberto mine. Sample lines vary in length from 1.1m to 22.45m and are generally oriented perpendicular to the vein strike. Individual sample intervals range from 0.06m to 3.8m in length (0.8m average). Line spacing is approximately every 2m in sampled areas. 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 2m line spacing and appropriate sample intervals. Intervals length can be variable and should represent the geology (eg. variations in sulphide and gangue minerals, proportions, textures and cross cutting structures).

• The mine surveyor should use the total stations TOPCON survey instrument to

determine the location of the sample lines.

• All bearings for survey purposes must be determined by turning angles from established survey points. A compass cannot be used except in a preliminary manner for starting a drift or cross-cut.

• Survey control point coordinates must be established at least every 50m in the

drifts and crosscuts. SAMPLING CONTROL

• Chip samples up to 20 cm wide are to be collected along the marked sample line. The line number and sample interval must be clearly entered in the



sample book. Two stubs should be placed in the sample bag by the sampler. The sample books must be archived at Cozamin.

• The geologist checks the sample bags and tag books for consistency before

releasing the samples to the laboratory for analysis. 15 SAMPLE PREPARATION, ANALYSES AND SECURITY 15.1 Sampling Personnel No employees, officers, directors or associates of Capstone are involved in preparation of the drill core samples used in the preparation of this report. Channel samples are prepared by Capstone employees for analysis at the on-site laboratory. Duplicate quality control samples (coarse crush and pulp) are also prepared by Capstone employees for analysis at an off- site laboratory. 15.2 Drill Core Sample Preparation and Analytical Procedures Diamond drill core samples are prepared as outlined in section 14.1. 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 should handle the core prior to sampling. This geologist should be very careful to return core 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 should be placed in order in a large

sack. The sack must be sealed with tape and by a numbered seal that prevents opening the sack without damaging the seal. Ideally, Capstone should try to use wire seals with teeth on the locking mechanism that prevent sliding the seal over the top of the sack. The sample number series of the enclosed samples should be clearly written on the exterior of the sack. The batch number, the serial numbers of the seals and the corresponding sample number series should be 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 15.1-15.3. Table 15.1: Primary and check laboratories used for Cozamin drill samples. Principal Laboratory Check Laboratory From To Phase BSi Inspectorate ALSChemex April 2004 August 2004 I ALSChemex BSI Inspectorate Sept 2004 March 2005 II SGS ALSChemex April 2005 April 2006 III ALSChemex 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 ALSChemex in Vancouver, Canada. ALSChemex has ISO registration in Canada (ISO 9001:2001 and ISO 17025). Phase II drill core samples were prepared in Hermosillo by ALSChemex 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 ALSChemex in Vancouver for analysis. Phase IV and V drill core samples were sent to ALSChemex in Hermosillo, Mexico, for preparation. The pulps are shipped directly to Canada for analysis by ALSChemex in Vancouver. Duplicate samples were sent to SGS Toronto for check analysis. Table 15.2: Laboratory methods and elements routinely analyzed.

Au Ag Cu Pb Zn

BSI Inspectorate

fire assay 2AT, atomic

absorption finish

aqua regia digest, atomic

absorption finish


absorption finish


absorption finish


absorption finish

ALSChemex

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 15.3: Analytical methods used when re-analyzing samples with over limit results.

Ag Over Limit Cu Over Limit Pb Over Limit Zn Over Limit

BSi Inspectorate

aqua regia digest, atomic absorption

finish


finish


finish

ALSChemex four acid digest,

titration

SGS

fire assay, atomic

absorption finish

four acid digest, sodium peroxide ICP-OES finish





Current Preparation Method Samples are sent to ALSChemex for preparation. Upon receipt samples are dried, weighed, crushed. Two hundred and fifty grams is split and pulverized to at least 85% passing 75 microns. Reject material is retained at ALSChemex in Hermosillo in a cold storage facility. Prepared pulps are sent to ALSChemex in Vancouver for primary analysis. Check sample pulps are sent to SGS Toronto for analysis. Current Analytical Methods At ALSChemex gold and silver are analyzed by fire assay with a gravimetric finish using a 50g charge. The detection range for this method is 0.05-1,000 ppm gold and 5-10,000 ppm silver. Silver is 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: 1-1,500 ppm silver, and 0.001-10,000 ppm for copper, lead and zinc.

Samples with over limit lead results are re-analyzed using the CON02 method. 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 (23-27% 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 - 2,000 ppb. Silver was analyzed from a 2g charge using a multi-acid digest with atomic absorption finish (0.3 – 300 g/t detection range). Over limit results were re-analyzed by fire assay with an atomic absorption finish using a 50g 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 – 10% for copper, 20 ppm - 10% for lead and 10 ppm - 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, eg. 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 1 in every 8 samples submitted for assay was 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 1 in every 6 samples submitted for assay was 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 ALSChemex 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. 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. 15.3 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 15.4). 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 ALSChemex 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 ALSChemex. Table 15.4:. 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 ALSChemex 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 every 15 samples, and pulp duplicates every 20 samples. Additional quality control samples were inserted into the sequence as deemed necessary. The QA/QC of these samples is presented in section 16. 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.



16 DATA VERIFICATION 16.1 Introduction Peter Christopher conducted a site visit to the Cozamin project on November 11th and 12th, 2004 to review core handling procedures and to collect independent samples for verification at a referee laboratory (2004 technical report on the Cozamin project, available on the sedar website). Christopher selected and supervised the splitting of 24 previously assayed intervals and secured the samples and delivered the samples to Acme Analytical Laboratories Ltd. in Vancouver. Capstone initially had every sample prepared by ALSChemex in Guadalahara, analyzed by ALSChemex in Vancouver and check analyzed by BSi Inspectorate in Reno, Nevada. No significant discrepancies were found in the data from the two laboratories. Assay results from Christopher’s check sampling confirmed previous sampling results and confirmed good core handling, sampling and analytical procedures. Prior to commencing the Phase III underground drilling program, Capstone took large samples from the dewatered underground workings and made in-house standards which were inserted into the sample streams from the underground holes and thereby significantly reduced the number of pulp duplicates sent to a second lab for checking the quality of the assays. Blanks were also prepared at site using cement. A formal written quality assurance and quality control program was developed by Hugh Willson, P.Geo., and Jenna Hardy, P. Geo., and implemented by Hugh Willson in 2005 prior to starting the Phase III underground drilling program, and has continued with minor revisions. The procedure was reviewed as part of the 2005 Technical Report by Christopher and Giroux and found to meet the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) standard of best practices. A review of the QA/QC data associated with the assays used in the 2005 resource estimate is provided in the 2005 Technical Report by Christopher and Giroux. The results demonstrated an acceptable level of analytical accuracy and precision. Analysis of the assay results of QA/QC samples associated with the drill holes subsequently completed and used in the 2006 estimate also demonstrated an acceptable level of accuracy and precision. The QA/QC analysis for assay results from samples collected as part of the Phase IV and V drill programs are summarized below and presented in detail in Appendix 1. These data demonstrate an acceptable level of accuracy and precision. In preparation for the 2007 resource update, the drill hole collar coordinates and down hole surveys were compared to the original records by the writer and updated where required. Sample intervals and assays were for approximately half of the surface and underground drill holes were also reviewed and found to be correctly documented.



16.2 QA/QC Summary for Phase IV and V Drill Samples The following sections are a summary of the QA/QC results from Phase IV and V exploration drill programs. Plots for all blanks, standards and duplicate samples are provided in Appendix 1. In this section less than detection limit values are considered equal to zero for reporting purposes so that they can be distinguished from values equal to the detection limit. Blanks One hundred and twelve blank samples of cement were submitted by Capstone as part of the Phase IV and V exploration programs. These blanks were submitted with drill holes CG-06-U115 through CG-06-U124 and CG-07-U125 through CG-07-U183. Thirteen hundred and fifty seven blanks were analyzed for gold and silver and 130 for copper, lead and zinc by ALSChemex in accordance with their internal QA/QC procedures. All samples had assay values within acceptable ranges of the expected values. Figures 16.3-16.6 show scatter plots of all elements for each the Capstone and ALSChemex blanks.

Figure 16.3: Gold and silver assays of blank samples submitted by Capstone. Three Capstone blank samples have silver values equal to the minimum detection limit for the assay method (5 g/t; Figure 16.3). One sample has a silver value of twice the minimum detection limit, however, it is well below the typical silver grades within the ore (minimum of at least 30 g/t). All other blank samples have assay values below the detection limit for each metal.



Figure 16.4: Copper, lead and zinc assays of blank samples submitted by Capstone. Copper, lead and zinc assays are generally at or ear the minimum detection limit for the analytical assay (0.01%; Figure 16.4). The largest deviation from the detection limit is one sample, which contains 0.19% copper. It is the same sample that has the high silver content (10 g/t). However, all Capstone and ALSChemex standards and the blanks associated with this assay have values that report well within their expected range of deviation.

Figure 16.5: ALSChemex blank samples analyzed for gold and silver. Seven ALSChemex blank samples have silver values equal to (2) or greater than (5) the minimum detection limit (5 g/t) for the assay method (Figure 16.5). All other blank samples have assay values below the detection limit for the respective metals.



Figure 16.6: ALSChemex blank samples analyzed for copper, lead and zinc. One ALSChemex blank had a lead value equal to the detection limit for the method and another blank had a zinc value twice the minimum detection limit. All other values were less than detection. Standards One hundred and three standards were submitted by Capstone with the drill hole samples analyzed. Figures 16.8 through 16.10 summarize the assay results. Details are shown in Appendix 1. Most values occur within the high and low range reported for the standards or within two standard deviations from the mean standard sample value (Appendix 1). None of the samples with values outside of the expected limits have suspect assays for all metals, are consistently high or low, or are associated with the same drill hole. The results of the standard analyses indicate that the assays are of good quality.



Figure 16.7: Copper assays of Capstone standards.

Figure 16.8: Silver assays of Capstone standards.



Figure 16.9: Lead assays of Capstone standards.

Figure 16.10: Zinc assays of Capstone standards. Nine hundred and thirty standards (13 different standards) were analyzed by ALSChemex as part of their internal QA/QC program. Graphs for all of these standards are shown in Appendix 1. Similar to Capstone standards, most values fall within the high and low range rexpected for the standards or within two standard deviations from the mean standard sample value (Appendix 1). None of the samples with values outside of the expected limits have suspect assays for all metals, are consistently high or low, or are associated with the same drill hole.



Pulp Preparation Duplicates Capstone submitted 106 pulp duplicates for analysis (Appendix 1). These samples were prepared by taking a second split of the pulp at the preparation stage by ALSChemex in their preparation facility in Hermosillo, Mexico. Table 16.1 summarizes the results and indicates that there is a very good correlation between the original assay and the duplicate. Table 16.1: Statistical analysis of Capstone prepared pulp duplicates. Au Ag Cu Pb Zn

R2 value 0.9942 0.9973 0.9841 0.9707 0.9167

Correlation Coefficient 0.9971 0.9986 0.9920 0.9852 0.9574

ALSChemex selected 59 pulp samples for check analysis. These samples were prepared from the pulps as received in Vancouver were the analyses are performed. The results are presented as a series of scatter plots in Appendix 1, and summarized in Table 16.2 below. The correlation between the original sample and the pulp duplicate is very good. Table 16.2: Statistical analysis of ALSChemex lab duplicates. Au Ag Cu Pb Zn

R2 value 0.9740 0.9792 0.9998 1.0000 0.9998


Core Duplicates One hundred and twenty seven core duplicates were submitted for analysis by Capstone (Appendix 1). Table 16.3 summarizes the analytical results and show a good correlation between the original assay copper, silver, lead and zinc assay and the core duplicate. Gold duplicates have a low correlation. Five duplicate pairs in particular do not compare well to the original assays. There are also several pairs with one result at the detection limit and the duplicate below detection or vice versa resulting in a bad correlation (less than detection values are plotted with a zero value to distinguish from values at the detection limit). Table 16.3: Statistical analysis of core duplicates. Au Ag Cu Pb Zn

R2 value 0.5384 0.9150 0.9338 0.9634 0.9308


Repeat Analyses and Alternative Method Check One hundred and eleven repeat analyses of gold and 111 repeat analyses of silver (not always the same samples) were also completed by ALSChemex (Appendix 1). Silver grades determined by fire assay (542 samples) were periodically checked using



the four acid digest - ICP method described in section 15. Table 16.4 summarizes the results of these checks and indicates excellent correlation of the original and duplicate assay at the analytical stage and using the alternative method. Table 16.4: Statistical analysis of ALSChemex repeat assays and alternative method check. Au Ag Ag ICP*

R2 value 1.0000 0.9989 0.9972

Correlation coefficient 1.0000 0.9995 0.9986

*Ag ICP is a check of the fire assay results with the four acid digest - ICP method. In summary, the QA/QC analysis associated with samples collected from the Phase IV and V drill holes show good accuracy and precision and indicate that the sample assay data is of high quality and suitable for resource estimation. 16.3 QA/QC Summary for the 2006 – 2007 Underground Chip Channel Samples Table 16.5 is a list of quality control samples analyzed with the channel samples at the Cozamin laboratory. Table 16.5: Quality control samples for the channel samples. QC Type Number CZ STD 4757-2 192 CZ STD 4787-2 131 CZ STD 4759 65 CDN STD 1 29 CDN STD 2 18 CZ BLANK 78 CDN BLANK 20

The Cozamin prepared standards generally assay with values in the expected range, except for lead which appears to be higher than the expected value of the standards. Most values outside of the expected range plot within two standard deviations from the mean of the sample mean. Therefore, the sample data are considered valid and acceptable to use in the resource modelling. Figures 16.11-16.14 are summaries showing the assay values of each metal for each standard. Individual graphs are presented in Appendix 2.



Figure 16.11: Silver assays of Capstone standards.

Figure 16.12: Copper assays of Capstone standards.



Figure 16.13: Lead assays of Capstone standards.

Figure 16.4: Zinc assays of Capstone standards.



Two standards from CDN Laboratories Ltd. (CDN) in Vancouver, Canada were purchased in 2007 and also used to monitor assay quality. The results are summarized in Figure 16.15-16.17. The lead values for CDN standards are slightly high, but are in general within two standard deviations of the mean standard sample value similar to the other metals. Individual graphs are presented in Appendix 2.

Figure 16.15: Silver assays of CDN standards.



Figure 16.16: Copper assays of CDN standards.

Figure 16.17: Lead assays of CDN standards.



Figure 16.18: Zinc assays of CDN standards. Two types of blanks were used for quality control: one prepared by Capstone consisting of cement and the other was purchased from CDN in 2007. The assay results are summarized in Figures 16.19-16.22, and provided in detail in Appendix 2. There is scatter around the detection limit and higher for the blank samples, however, the values are well below those expected for the mineralization. As per Capstone’s QA/QC procedure, duplicate samples were sent to ALSChemex for analysis. Six hundred and thirteen samples were analyzed for silver, 649 for copper, 608 for lead and 606 for zinc. Figures 16.22-16.25 summarized the analytical results. Detailed plots are presented in Appendix 2. The data show a general linear trend with R2 values of approximately 0.7 and above and correlation coefficients greater than 0.83 (Table 16.6). Table 16.6: Statistical analysis of duplicate chip channel samples. Ag Cu Pb Zn

R2 value 0.7990 0.8672 0.7292 0.6910

Correlation Coefficient 0.8939 0.9313 0.8539 0.8310

In summary, the QA/QC analysis associated with underground chip channel samples show good accuracy and precision and indicate that the sample assay data is suitable for resource estimation.



Figure 16.19: Silver assays of blanks.

Figure 16.20: Copper assays of blanks.



Figure 16.21: Lead assays of blanks.

Figure 16.22: Zinc assays of blanks.



Figure 16.23: Silver assays of check samples.

Figure 16.24: Copper assays of check samples.



Figure 16.25: Lead assays of check samples..

Figure 16.26: Zinc assays of check samples.



17 ADJACENT PROPERTIES The Mala Noche vein is one of several main veins that have been exploited since pre-colonial times in the Zacatecas area (Figure 8.1). The Bote vein has recently been in production but production on the Veta Grande, Panuco, Mala Noche, Cantera and San Rafael veins has been intermittent depending on silver and base metal prices. The average ore grades for the Zacatecas district are reported to be 1.5 g/t gold, 120 g/t silver, 3% lead, 5.1% zinc and 0.16% copper and with total silver production to the end of 1987 estimated to be approximately 750,000,000 ounces (Ponce and Clark, 1988). 18 MINERAL PROCESSING AND METALLURGICAL TESTING 18.1 Introduction The processing operations at the Cozamin mine are contained within one plant which treats a polymetallic ore. The processing of polymetallic ores commenced in August 2006, and production statistics are available for 13 months (September 2007). The plant underwent an expansion which was completed in July 2007 to raise capacity to 2,200 tpd and has treated a maximum rate of 2,530 tpd since commissioning was completed. 18.2 Ore Processing 18.2.1 Ore Types For the purpose of processing, the ores are categorized as low copper-high lead zinc ores and high copper- low lead zinc ores. Low copper ores are contained in the upper part of the ore body. Currently ore feed to the plant is a blend of both ore types, but over time the high copper ores will predominate. 18.2.2 Previous Plant Performance A plant has operated off and on at this site with the last campaign occurring in 1998-1999 when a 750 tonne per day flotation plant was operating owned by Bacis. An examination of historical records indicated that the plant treated approximately 250,000 tonnes of ore grading 1.2% copper, 90 grams per tonnes silver, 1.8% zinc and 0.6% lead. Production to concentrate was 2,370 tonnes of copper, 495,000 ounces of silver, 2,802 tonnes of zinc and 856 tonnes of lead. Three different concentrates were produced; copper, zinc and lead. Recoveries in the mill to specific concentrates were 80% for copper, 65% for zinc and 70% for lead. Total silver recovery to all three concentrates was approximately 68%. Current operations commenced at 1,000 tpd in August 2006 and were expanded in July 2007 to 2,200 tpd. The metallurgical performance from September 2006 to August 2007 is summarized in Table 18.1.



Table 18.1: Plant production data September 2006 to August 2007. Grade Recovery (%) Tonnes Copper

(%) Zinc (%)

Lead (%)

Silver (g/t)

Copper Zinc Lead Silver

Ore 461,933 1.59 1.47 0.56 71 100 100 100 100 Copper Concentrate

27,874 22.66 3.26 2.34 634 86 14.3 26.2 54

Zinc Concentrate

7,157 2.41 42.4 0.92 115 2.5 44.9 2.6 3

Lead Concentrate

2,051 3.33 1.49 64.4 2423 1 0.5 52.6 16

Ore throughput averaged 31,000 tonnes per month during the first three months of operation, and 54,000 tonnes per month in the last three months. Daily grades for copper varied from 1.25 to 2.25% copper, with a corresponding range in zinc. Both recoveries to concentrate and concentrate grades improved throughout the year. 18.2.3 Plant Flow sheet Description Flow sheets for the plant are presented in Figures 18.1 through 18.4. 18.2.3.1 Crushing (Figure 18.1) Although a jaw crusher is being installed underground, ore is presently ore is trucked from the head frame bin to a surface stockpile to allow blending for consistent copper grade feed. The surface stockpile of approximately 10,000 tonnes is reclaimed by a front end loader which feeds material to a 100 tonne bin. Material is withdrawn from the bin by a belt feeder and fed to a 20” x 36” primary jaw crusher. Crusher product is conveyed to a 5’ x 12’ vibrating screen ahead of the 4’ standard secondary crusher. Screen oversize is fed to the crusher with screen undersize combined with secondary crusher product. This material is conveyed to a 6’ x 16’ vibrating screen with oversize material conveyed to the tertiary crusher (1560 Omni cone crusher) and undersize material being conveyed to the fine ore bins. Tertiary crusher product is returned to the 6’ x 16’ 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. 18.2.3.2 Grinding (Figure 18.2) The grinding section is composed of two primary ball mills each 3.65m in diameter by 4.27m long. Feed rate to each mill is approximately 50 tonnes per hour to produce a product sizing of 80% passing 100 mesh. Each ball mill is operated in closed circuit with a cyclone pack composed of 0.66m diameter cyclones. Cyclone under flow is returned to the grinding mill while cyclone overflow from both circuits flows by gravity to a flotation conditioning tank.



18.2.3.3 Flotation (Figure 18.3) Ground slurry flows by gravity to a bank of rougher and scavenger flotation cells (6-OK 16 cells). Rougher flotation product contains copper-lead mineralization and is sent to a two stage cleaning flotation circuit consisting of 4 Wemco 300 cells for first cleaners, and 2 Wemco 300 cells for second cleaner. Rougher flotation tailings are sent to a zinc conditioner tank ahead of zinc rougher flotation. Copper-lead cleaning is performed in bank of 4 Denver 24 cells, which produce a final copper concentrate product and a lead product which is further cleaned in a second bank of 4 Denver 24 cells to produce the final lead concentrate. Final copper and lead products are pumped to their respective concentrate thickeners. Zinc rougher flotation takes place in a bank of 5 OK 16 cells, with the tailings from this bank representing final plant tails and pumped to the tailings dam. Rougher concentrate is subjected to two stages of cleaning (1st stage in a bank of 4 Denver 200 cells, and 2nd stage in 3 OK 5 cells) with tailings recycled back in the circuit. Final stage cleaner concentrate is pumped to a zinc concentrate thickener. 18.2.3.4 Concentrate Dewatering (Figure 18.4) Copper concentrate is pumped from the 16m diameter concentrate thickener and dewatered using a 2m x 4m disc filter. 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 kilometres) and sampled as the material is transferred to the port warehouse and becomes the property of the buyer. Zinc concentrate is pumped from the 8m diameter thickener to the 1.3m diameter x 4m 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 4m diameter thickener to a 1.3m diameter x 2m 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. 18.2.3.5 On Stream Analysis The plant is equipped with an Otokumpo courier on stream analyzer with automatic sampling occurring on ball mill cyclone overflow (head sample) throughout the flotation circuit and on final tailings. Manual sampling and automatic sampling are also done with automatic samplers on the plant feed conveyors and final tailings streams, while manual sampling is done on an hourly basis throughout the flotation circuit. Analysis for copper, lead, zinc and iron are produced every 10 minutes for flotation control.



Figure 18.1: Flow sheet for crushing ore at the Cozamin mine.



Figure 18.2: Flow sheet for grinding ore at the Cozamin mine.



Figure 18.3: Flow sheet showing the flotation process at the Cozamin mine.



Figure 18.4: Flow sheet showing ore thickening and de-watering at the Cozamin mine.



18.2.3.5 On Stream Analysis The plant is equipped with an Otokumpo courier on stream analyzer with automatic sampling occurring on ball mill cyclone overflow (head sample) throughout the flotation circuit and on final tailings. Manual sampling and automatic sampling are also done with automatic samplers on the plant feed conveyors and final tailings streams, while manual sampling is done on an hourly basis throughout the flotation circuit. Analysis for copper, lead, zinc and iron are produced every 10 minutes for flotation control. 18.2.3.6 Reagents The main reagents used in the plant are lime for pH adjustment throughout the circuits, aerophine as the promoter, cyanide to assist in the lead copper separation, zinc sulphate in the copper circuit to depress zinc, and copper sulphate in the zinc circuit to activate zinc. 18.2.3.7 Concentrate Marketing Lead and zinc concentrates are currently sold under contract to Glencore at standard industry terms. Contracts currently in place continue through to the end of 2008. Copper concentrate is sold to both Trafigura and Glencore for 2007 and 2008 and to Trafigura for 2009. These contracts have industry standard terms and do not have any price participation. At present all concentrate is trucked to warehouses at the port of Manzanillo, Mexico for ocean shipment to smelters. 18.3 Laboratory The laboratory is under the supervision of the engineering department. Most control assaying is done by atomic absorption, with concentrates also subject to fire assay for silver and wet chemical analysis for lead, zinc and copper. Approximately 40% of the sample load is mine samples, 50% is plant control and 10% is concentrate analysis. Exploration diamond drill hole analysis is done by an off- site laboratory with checks sent to another off- site laboratory. 18.4. Plant Personnel A plant manager is responsible for the operation and budgeting of the concentrator. The plant operates 24 hours per day, 7 days per week with four operating crews. Each crew consists of a shift foreman and seven operators and two helpers. Plant maintenance is under the supervision of the Maintenance Superintendent and includes two shift mechanics as well as the day crew that includes mechanics, millwrights, electricians and instrumentation engineers. The mill metallurgist is responsible for concentrate quality control and all test work conducted in the metallurgical laboratory. Total number of people in the mill is approximately 72 people including 6 maintenance personnel on contract. 18.5. Plant Operating Costs Operating costs for the period September 1, 2006 to August 31, 2007 are summarized in Table 18.2 below. Operating costs were trending lower at the end of the time period due to the increased plant throughput.



Table 18.2: Plant operating costs September 2006 through August 2007. Area 2006-2007 Actual Costs

$US/tonne 2008 Estimate Costs

$US/tonnes Crushing 1.36 1.20 Grinding 4.11 4.00 Flotation 3.94 3.50

Dewatering 1.05 1.00 Reagent Preparation 0.29 0.20

Tailings 0.21 0.25 Equipment Leases 0.46 0.40

Mill G&A, including Laboratory 0.95 0.75 Maintenance 1.53 1.50

Total 13.90 12.90 19 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES 19.1 2005 Initial Resource Estimate An initial NI 43-101 compliant resource for the Mala Noche vein was estimated by independent consultant Gary Giroux., P.Eng. (Giroux Consultants Ltd.) using data available as of October 5, 2005 (Christopher and Giroux 2005). The database used for the initial resource estimate consisted of 37 exploration surface diamond drill holes numbered CG-04-01 to 33 and CG-05-34 to 37 and 66 underground diamond drill holes numbered CG-05-U01 to U64, U69 and U70 (Figure 19.1). Forty two historic Bacis channel samples were used in the estimation in addition to 6 lines of Capstone channel samples from Level 8 of the San Roberto mine. Capstone geologists determined the entire interval for the Mala Noche vein in each drill hole. The vein interval used for this estimation was the entire width for the vein regardless of the grade tenor. Grades for copper, zinc, lead and gold were estimated using 2m composites and ordinary kriging. Results are summarized below (Table 19.1). Detailed tables are included in the 2005 Christopher and Giroux technical report available on the SEDAR website. Table 19.1: Resources estimated for the Cozamin mine in 2005 reported with a 1% copper cut-off.


Cu (%)

Ag (g/t)

Pb (%)

Zn (%)


Measured 560,000 2.00 85.46 0.70 1.03 0.07 24.7 8.6 12.7 1.5 Indicated 2,430,000 1.86 70.54 0.49 1.05 0.05 99.7 26.3 56.3 5.5 Measured + Indicated 2,990,000 1.89 73.33 0.53 1.05 0.05 124.4 34.9 69.0 7.0 Inferred 3,550,000 1.67 56.99 0.17 1.11 0.04 130.7 13.3 86.9 6.5 Total 6,540,000 1.77 64.46 0.33 1.08 0.05 255.1 48.2 155.9 13.5 * 1 kilogram = 2.2 lbs; 31.103 grams = 1 oz. Troy



Figure 19.1: West-east long section through the Cozamin area showing drill holes used in the 2005 estimate (grey) and additional holes drilled and included in the 2006 (blue) and 2007 estimates (red).



19.2 2006 Updated Resource Estimate In July 2006, an updated resource estimate was completed by Gary Giroux, P.Eng., of Giroux Consultants Ltd (Table 19.2). Details of the resource update are provided in the 2006 technical report by Christopher, Stone and Giroux available on the SEDAR website. Major changes or differences from the October 2005 study included the following: • Data from an additional 159 lines of channel sampling and 46 underground drill holes (Figure 19.1). The additional drill hole data was both infill to and deeper than previous drill holes. • An additional 2,071 specific gravity determinations bringing the total to 2,990. • A reinterpretation by Capstone of the Mala Noche vein breaking it into a sulfide-rich core surrounded by a mineralized envelope. • Three dimensional solids were created for each domain, and each domain was modeled and estimated separately. Table 19.2: Resources estimated for the Cozamin mine in 2006 reported with a 1% copper cut-off.


Cu (%)

Ag (g/t)

Pb (%)

Zn (%)


Measured 550,000 2.58 86.56 0.48 1.07 0.05 31.3 5.8 13.0 1.5

Indicated 2,210,000 2.48 85.13 0.47 1.14 0.05 120.9 22.9 55.6 6.0 Measured + Indicated 2,760,000 2.50 85.41 0.47 1.13 0.05 152.2 28.7 68.6 7.5 Inferred 3,090,000 2.22 79.07 0.27 1.29 0.06 151.3 18.4 87.9 7.9 Total 5,850,000 2.35 82.06 0.37 1.21 0.06 303.5 47.1 157 15.4 * 1 kilogram = 2.2 lbs; 31.103 grams = 1 oz. Troy 19.2 2007 Updated Resource Estimate The Phase IV and V drill programs proposed in the October 2006 technical report (Christopher, Stone and Giroux) and partially reported in the March 2007 technical report (Stone and Giroux) were completed in July 2007. This provided information from 5 additional surface and 69 underground drill holes to update the 2006 resource estimate. An additional 1,125 lines of underground channel sampling were also completed and used in preparation of the revised estimate described below (Figure 19.1). 19.2.1 Database, software and three dimensional models The resource estimate is based on the interpretation of 42 surface drill holes, 183 underground drill holes and 1,272 lines of underground chip channel sampling. The collar position, down hole survey, and assay data for these holes and channels are stored in an Access database. Sample intervals with assays below the minimum detection limit were assigned corresponding metal grades of zero. Similarly, intervals not sampled in the drill holes were assigned zero value grades to minimize potential smearing of grade inside and alongside of the generally thin, mineralized structures.



Channel sample assay grades were reduced by 0.1% copper and 10 g/t silver based on a reconciliation of the channel sample grades assayed at the Cozamin laboratory and the mill throughput (section 19.4.5). When these reductions were applied, the channel sample grades more closely matched those in adjacent drill holes. Three dimensional (3D) models of the copper-dominant mineralization within the Mala Noche vein, and associated hangingwall and footwall mineralization were constructed using drill hole cross sections, and level plans showing the channel samples and the underground mapping. SurpacVision software V. 6.0.1 was used to generate the 3D models and perform the grade estimation. 19.2.2 Domain Modelling The resource estimation is constrained by domains interpreted from geological data. There are three domains at Cozamin based on the generalized bearing and dip of the Mala Noche vein (Figure 19.2 and Table 19.3).

Figure 19.2: West-east long section showing the ore solid and location of the domain boundaries for the 2007 resource update. Table 19.3 Geological domains used in the 2007 resource update. Domain Bearing Dip West 120 45 Central 83 60 East 104 59



19.3.3 Compositing Histograms of sample length for the drill holes and surface channels are presented in Figures 19.3 to 19.5.

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Figure 19.3: Histogram of footwall sample lengths.

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Figure 19.4: Histogram of ore sample lengths.



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Figure 19.5: Histogram of hangingwall sample lengths. Two metre composites were produced from the assay data for each the hangingwall, footwall, and main ore zone. Intervals less than 2m were composited near the contacts where sample length was ≥ 0.02m (1% of the maximum composite length). Table 19.4 summarizes the composite intervals statistics. Table 19.4: Summary of composite intervals statistics. Footwall Ore Hangingwall

Number of composites 1281 3780 3978 Number of drill hole composites 795 816 3212 Number of channel composites 486 2964 766

Maximum composite length 2 2 2 Number of composites = 2m 781 2460 3306

Minimum composite length 0.02 0.02 0.02 Number of composites < 2m 500 1320 672 Number of composites < 1m 292 650 379

Number of composites < 0.5m 133 333 219 Number of samples excluded 2 3 3

The coefficient of variation was checked for the composites in each the footwall, ore and hangingwall. If the coefficient of variation was greater than 1.2, the grade of the specific metal was capped using the 99th percentile value. A summary of the capping statistics for the composites is shown in Table 19.5. The capping and smoothing during formation of composites has reduced the coefficient of variation for the all of the metals. Lead and gold still have high coefficients of variations indicating higher sampling variability.



Table 19.5: Summary of assay statistics for composite samples. GOLD Footwall Ore Hangingwall Number of samples 1281 3780 3978 Minimum 0 0 0 Maximum 0.647 3.510 4.495 Coefficient of Variation 3.609 5.833 8.076 Cut grade 0.242 0.232 0.203 Number cut 13 38 40 Revised Coeffic. of Var. 2.928 3.187 3.227

SILVER Footwall Ore Hangingwall Number of samples 1281 3780 3978 Minimum 0 0 0 Maximum 540.802 781 1182.939 Coefficient of Variation 1.960 0.917 3.038 Cut grade 223.383 NONE 173.706 Number cut 13 0 40 Revised Coeffic. of Var. 1.742 0.917 2.414

COPPER Footwall Ore Hangingwall Number of samples 1281 3780 3978 Minimum 0 0 0 Maximum 7.143 11.398 9.680 Coefficient of Variation 1.894 0.917 2.735 Cut grade 5.120 NONE 4.373 Number cut 13 0 40 Revised Coeffic. of Var. 1.823 0.917 2.484

LEAD Footwall Ore Hangingwall Number of samples 1281 3780 3978 Minimum 0 0 0 Maximum 12.29 20.748 18.879 Coefficient of Variation 4.803 2.832 6.285 Cut grade 3.502 8.714 2.697 Number cut 13 38 40 Revised Coeffic. of Var. 3.414 2.480 3.782

ZINC Footwall Ore Hangingwall Number of samples 1281 3780 3978 Minimum 0 0 0 Maximum 30.770 30.770 24.000 Coefficient of Variation 1.927 1.355 2.679 Cut grade 9.700 9.674 6.168 Number cut 13 38 40 Revised Coeffic. of Var. 1.657 1.234 2.326



19.2.4 Block model The block model parameters are shown below in Table 19.6. Block dimensions in the north-south direction were limited to 2m to try to restrict the across vein smoothing and honour the mineral zonation across the vein. For each block the proportion of the block within the hangingwall, footwall and main ore zone was determined. A rotation was not applied to the model. Table 19.6: Block model extents. Y (m) X (m) Z (m)Model Minimum Extents 2,523,390 746,530 1550 Model Maximum Extents 2,524,580 748,460 2600 Block Size 2 10 5 19.2.5 Density Specific gravity (SG) measurements have been made on 3,734 pieces of half core from exploration holes from Phase I through V. This value represents an increase of 744 determinations over the database used in the July 2006 study. Specific gravity values were interpolated into the model using inverse distance squared and the third pass search ellipsoid applied for the grade interpolation. Blocks not assigned SG values from the inverse distance squared interpolation were assigned the average of the 3,374 measurements, a value of 2.86. A summary of the density values used for the estimation of the hangingwall, footwall and main ore zone SGs are shown in Table 19.7. Table 19.7: Summary of SG sample statistics.

Hanginwall Ore Footwall Number of samples 915 1,938 626

Maximum SG 9.02 6.05 4.95 Minimum SG 1.3 0.78 1.67 Average SG 2.71 2.96 2.78

19.2.6 Block Interpolation The Mala Noche mineralization and surrounding rock was divided into 3 domains based on generalized orientation of the vein. The grade and SG of each domain was modelled individually. Grades were estimated in 3 successive passes (Table 19.8) for each domain using an expanding search ellipse with an anisotropy ratio of 10:1 across strike. A minimum of 3 samples were required for each estimation, to a maximum of 15. If more than 15 samples were found, then only the closest 15 were used in the estimate. The SG was similarly modeled but only using the Pass 3 search ellipse. Table 19.8: Summary of search distances used in the grade and SG estimates.

Search Distance (m) Attribute Estimated Pass 1 30 Au, Ag, Cu, Pb, Zn Pass 2 60 Au, Ag, Cu, Pb, Zn Pass 3 120 Au, Ag, Cu, Pb, Zn, SG



Blocks were estimated below 2,274m elevation (approximately 20m above the current development). When calculating the tonnage for each block the volume of the block was multiplied by the estimated SG and by the proportion of the block within the hangingwall, footwall or ore solid as appropriate. 19.2.7 Classification Based on the study herein reported, delineated mineralization of the Mala Noche vein, Cozamin project is classified as a resource according to the following definition from NI 43-101:

“In this Instrument, the terms "mineral resource", "inferred mineral resource", "indicated mineral resource" and "measured mineral resource" have the meanings ascribed to those terms by the Canadian Institute of Mining, Metallurgy and Petroleum, as the CIM Standards on Mineral Resources and Reserves Definitions and Guidelines adopted by CIM Council on August 20, 2000, as those definitions may be amended from time to time by the Canadian Institute of Mining, Metallurgy, and Petroleum.” “A Mineral Resource is a concentration or occurrence of natural, solid, inorganic or fossilized organic material in or on the Earth's crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.”

The terms Measured, Indicated and Inferred are defined in NI 43-101 as follows:

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



evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes that are spaced closely enough for geological and grade continuity to be reasonably assumed.”

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

Results The geologic continuity of the Mala Noche vein is well established through underground mapping, sampling and mining. Areas below mining stopes have been well tested with diamond drilling with the vein intercepts logged and sampled. Blocks in the footwall and ore zone were classified as follows: Measured: if the minimum distance to a sample point from a model cell was <20m. Indicated: if the minimum distance to a sample point from a model cell was 20-40m. Inferred: if the minimum distance to a sample point from a model cell was 40-80m.

Blocks in the hangingwall were not assigned a resource classification because additional exploration is required to confidently define their continuity and extent. The resulting tonnage from the updated resource estimate was calculated by:

block volume (10m x 5m x 2m) * estimated block SG * proportion of block within each of the hangingwall, footwall, and ore.

These values were adjusted by the amount of material mined, and then summed as reported in the Table 19.9. Significant changes from the 2006 estimate include a change in modelling method from kriged to inverse distance squared, grade estimation search distances and resource classification method. In particular, the resource classification used in preparation of this resource does not include model cells greater than 80m from the estimated cell as compared to the 2006 model which included all estimated cells.



Table 19.9: 2007 Cozamin resource model estimate. Million lbs* Million ozs*

Tonnes Cu (%)

Ag (g/t)

Pb (%)

Zn (%)


Measured 2,591,705 2.48 87.11 0.43 1.18 0.03 141.7 24.6 67.4 7.3 Indicated 2,896,158 2.59 86.37 0.32 1.14 0.04 165.4 20.4 72.8 8.0 Measured + Indicated 5,487,863 2.54 86.56 0.37 1.15 0.04 307.1 45.0 140.2 15.3 Inferred 3,162,838 2.36 80.50 0.18 1.03 0.04 164.6 12.6 71.8 8.2 Total 8,650,702 2.47 84.35 0.30 1.11 0.04 471.7 57.6 212.1 23.5

* 1 kilogram = 2.2 lbs; 31.103 grams = 1 oz. Troy 19.3 Mineral Reserves Estimation 19.3.1 Introduction The Cozamin mineral reserve estimate presented in the following sections is based on the results of the 2007 resource estimate. The operating results over the past year provided a basis for the dilution estimate used in the reserve calculations. The reserve estimate was prepared by Robert Barnes, P.Eng., of Capstone. The mineral reserves are calculated using only resources classified as measured or indicated as required by NI 43-101 (see section 19.3.7). Inferred and unclassified material is not considered. According to the definitions in the CIM Standards on Mineral Resources and Reserves Definitions:

“A Mineral Reserve is 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.”

Proven and probable reserves have been estimated as described in the following sections. The terms Proven and Probable Reserves are defined in NI 43-101 as:

“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. “

19.3.2 Cut-off Grade The mining cut-off grade is based on a combination of economic, metallurgical, mining and geologic factors including:

• metal prices • mining methods • mill recoveries • operating costs • treatment and refining costs • transportation cost • capitalized development

Taking these factors into consideration, the mine production schedule is based on a grade of ore greater in value than $US40 NSR. Mining blocks were assigned an NSR value by applying existing metallurgy (Table 18.1) and terms of existing sales contracts to the block grade. Metal prices used in the NSR calculation were: $8.50/oz silver, $2.25/lb copper, $1.00/lb zinc and $0.60/lb lead. 19.3.3 Recovery and Dilution The resource model is divided into mining blocks based on the mine levels and access. Above Level 11, the sill pillar levels are established every 60m. Below Level 11 the sill pillars are established on 70m intervals. Five metre high horizontal sill pillars (sills) are established between levels and 10m vertical pillars are located as required. Mineral mined in previous years and left in sills and pillars is excluded from the reserve estimate. The mineral left as sills and pillars represents 16% of the mineral reserves. Figure 19.6 is a longitudinal section in the vein showing the mining blocks and the planned sill and pillar locations. Dilution is applied to the mining blocks dependant on the mining method employed. The cut and fill mining method has a dilution for mining and loading that varies between 11% and 15% depending on the mining width. The long holes stopes have a dilution for mining that averages 13% but varies between 9% and 22% depending on width and the amount of inferred resources in the block that are considered waste. The average dilution for the mine is estimated to be 13%.

19.3.4 2007 Cozamin Reserves Measured and indicated resources within the mining blocks have been converted to proven and probable reserves (Table 19.10 through 19.13) by considering the following factors which determine economic viability:



Figure 19.6: West – east long section showing the Cozamin mining blocks, horizontal sills and vertical pillars with development as of August 31, 2007.



• mine plans and production schedules • extraction and process costs • mining recoveries and dilutions • plant recoveries.

Resources below Level 13 and material classified as Inferred are not reported in the reserves because further drilling is needed to demonstrate the economic viability. Table 19.10: Summary of the mineral reserves at the Cozamin mine on August 31, 2007 reported using $US40 NSR/tonne. Classification Tonnes Cu (%) Ag (g/t) Pb (%) Zn (%) Proven 1,809,719 2.32 84 0.45 1.17 Probable 1,915,248 2.42 81 0.34 1.19 Total 3,724,967 2.37 82 0.40 1.18

Table 19.11: Breakdown of the 2007 proven reserves by mining block reported using $US40 NSR/tonne.

Block Tonnes Width Cu (%) Ag (g/t) Pb (%) Zn (%) * 8W 88,373 5.23 1.04 92 0.63 2.24 9W 55,113 5.29 1.28 84 0.47 1.59

10W 40,567 4.69 1.80 98 0.34 1.54 11W 21,066 4.46 2.55 82 0.07 1.21 12W 48,864 6.30 2.23 64 0.05 0.42 13W 15,764 5.04 2.90 90 0.18 1.17 * 9C 45,506 4.80 1.50 82 0.51 2.26 10C 283,431 6.60 1.87 68 0.19 1.38 11C 205,010 5.94 2.76 75 0.12 1.38 12C 54,286 5.05 3.65 83 0.08 1.25 13C 58,531 5.34 3.26 64 0.09 1.31 * 8E 36,235 4.85 1.06 74 0.71 1.47 * 9E 179,771 6.60 1.25 88 0.81 0.96 10E 289,488 8.59 2.05 89 0.79 0.64 11E 175,783 7.41 2.59 88 0.52 0.70 12E 111,538 7.35 3.72 97 0.54 0.99 13E 100,392 6.93 4.12 97 0.20 1.17

TOTAL 1,809,719 6.62 2.32 84 0.45 1.17 * Cut and fill stope.



Table 19.12: Breakdown of the 2007 probable reserves by mining block reported using $US40 NSR/tonne.



TOTAL 1,915,248 6.25 2.42 81 0.34 1.19 * Cut and fill stope. Table 19.13: Breakdown of the 2007 total reserves by mining block reported using $US40 NSR/tonne.



TOTAL 3,724,967 6.43 2.37 82 0.40 1.18 * Cut and fill stope.



19.3.5 Reconciliation Reconciliation of the mine, plant and modelled tonnes and grades is an important tool in the verification of the resource model and the dilution estimate. The mined grades are estimated from channel and face samples with allowances for mining losses and dilution. After correcting for dilution and mining losses the channel sample silver grade is further reduced by 10 g/t and the channel sample copper grade is reduced by 0.1%. Table 19.14 shows the reconciliation between the mine, plant and modelled tonnes and grade. Table 19.14: Mine, plant and model tonnes and grade reconciliation. Tonnes Ag

(g/t) Cu (%) Pb (%) Zn (%)

Mine Estimate 428,294 75 1.66 0.48 1.53 Plant Sampled 431,406 71 1.61 0.59 1.5 Difference 3,116 -4 -0.06 0.1 -0.03 Percent Difference 1% -6% -3% 17% -2% Model Estimate at $US40 NSR 471,132 88 1.74 0.62 1.53 Plant Sampled 431,406 71 1.61 0.59 1.5 Difference -39,726 -17 -0.13 -0.04 -0.03 Percent Difference -9% -24% -8% -6% -2% The model versus actual plant feed over predicts all the tonnage and grade parameters. The database is small at this point in time, and requires another year of data collection to better determine actual comparisons. The reconciliation in future should also include actual sales of concentrate as year end adjustments to concentrate tonnage were positive reducing either the grade differences or increasing plant recovery. Mineral reserves are adjusted annually by the amount mined, by additions and deletions resulting from new geological information and interpretation, in conjunction with changes in operating parameters and metal prices. However, proven and probable mineral reserves are not usually revised in response to short-term fluctuations in the metal markets. Table 19.15 is a reconciliation of the proven and probable mineral reserves at Cozamin to September 30, 2007: Table 19.15: Reconciliation of mineral reserves at Cozamin. Tonnes Opening balance, February 2006* 2,260,000Additions 1,896,373Less Tonnes Milled 431.406Closing balance as of September 30, 2007 3,724,967

*2006 reserves were calculated using metal prices of: $1.25/lb copper, $6.25/oz silver, $0.38/lb lead and $0.50/lb zinc.



19.4 Conclusions The results of the Phase IV and V exploration drill programs were used to update the mineral resources at Cozamin. The updated resource model shows an increase in tonnes in all resource categories and significant increases in the contained metal content. The mineral reserves at Cozamin have been estimated down to Level 13 from the measured and indicated resources. At the current production rate of 2,200 tpd this provides a 5 year mine life.

Infill diamond drilling should continue in the mining blocks to convert Inferred resources to proven and probable reserves. Additional drilling should de done below Level 13 under the mining blocks and also to the east and west of the mining blocks between Levels 8 and 13. The old mine working on the west side of the mine should be dewatered for sampling and also for the construction of underground diamond drill stations.

Operation should be optimized to: improve mine production and plant throughput; increase plant recovery and operating efficiency; and lower mine and plant operating costs. A standby generator should be purchased to provide emergency power during power outages and to facilitate quicker plant restarts when the power comes back on.

A 12% access ramp from surface to Level 12 should be contracted to reduce ore haulage costs out of the mine and to ensure long term production from lower levels 20 OTHER RELEVANT DATA AND INFORMATION The authors are not aware of any other relevant data and information which is relevant to the project.

21 INTERPRETATION AND CONCLUSIONS The Cozamin mine is located approximately 4 kilometres from the city of Zacatecas in the state of Zacatecas, Mexico. Capstone restarted the mine in July 2006 after a successful exploration and development program. Operating statistics from September 1, 2006 to August 31, 2007 were available. The resources are contained in the Male Noche vein with lower copper grades and higher lead zinc grades in the upper areas of the mine, and stronger copper grades with lower lead – zinc in the lower areas of the mine. The first year of operation has provided an accurate data base including operating costs, metallurgical recovery and the impact of an expanded operating rate.



The resource estimate is prepared in accordance to the requirements of NI 43-101and CIM guidelines. The model was based on individual zones for each of the main zone, footwall zone and hanging wall zone. This methodology prevents smearing of grade between zones and in particular into areas of no mineralization between the main zone and the hangingwall zone. The database is fully functional and provides easy and secure access to the data upon which resource estimation is based. The database is composed of diamond drill holes and channel samples in the areas developed for mining. Channel sample grades were reconciled to plant feed grades and values for copper and silver were reduced to enable matching between predicted mine grade and actual plant grade. This methodology is conservative due to the inherent upsets in a plant in the first year of operation, and a larger data base should be used before correction factors are introduced to geological data as a normal procedure. The resources calculation shows a high degree of conversion of inferred resources in the June 2006 report to measured and indicated resources in the current report due to infill drilling and actual mining. The methodology used to calculate the mining reserves at Cozamin are simple and transparent. The cut off grade applied takes into account economic, mining and geological factors. Minimum mining widths, development, access, stoping methods and previous mining are considered. Factors for recovery and dilution are a combination of actual experience and estimates related to particular stoping methods. The mine access is provided by one vertical shaft and one ramp from surface at a gradient of 15% until Level 9 and 12% thereafter. Two stoping methods are used but the most prominent is the longhole method. Dilution in the longhole has been highlighted as a potential issue but has been reasonably addressed in the reserve calculation. The processing operations have continued to improve during the year, but still are subject to frequent interruptions from power failures (utility). The introduction of an on stream analyzer in mid 2007 has begun to allow better control in the plant. Steady state operations have been hindered in the first year of operation due to constant expansion modifications in the plant during operations. The increase in operating rate from 350,000 tonnes per year to 750,000 tonnes per year was accomplished with minimal downtime and has managed to operate at 96% of design capacity in the first three months of expanded tonnage. The tailings dam is a water retaining embankment dam. Tailings are deposited into the tailings dam to form a beach area and a clear pond area for water reclaim. Mine water is also pumped to the tailings dam, and mixed with pond water to improve the overall quality of reclaimed water.



Environmentally, the main challenge to the mine will be to ensure tailings remain oxygen deficient to ensure no acid generation and subsequent metal dissolution. Tailings are sent to the dam at a ph of approximately 10 (leaving the zinc rougher flotation circuit) and reclaim water ranges from pH 7-8 as a result of a combination of rain water (pH 6) mine water pH 7.5 and process water pH 10. The mine has removed historic spills downstream of the tailings dam and installed a small grouted catchment dam to ensure any seepage is returned to the dam and not allowed to percolate through old tailings. There are no waterways downstream of the mine property, and only small local agricultural water ponds exist, limiting any potential impact. The company has purchase the surface land needed for the ultimate tailings dam (+10 million tonne capacity) as well as a buffer zone to limit impact to any surrounding communities and to provide space between the growing city of Zacatecas and the mine site. The company has filed a closure plan with the Mexican authorities, and this plan is updated and filed on an annual basis. The plan is designed with progressive closure activities during the life of the mine, and final closure at the end of the mine life. Monitoring of both water quality, geotechnical stability of the tailings dam, and successful re-vegetation lasts for 5 years after closure. Progressive closure is unlikely to occur as the major historic clean ups required have been completed under the expansion in 2007. Mine plan modifications have resulted in no waste being transported to surface removing the necessity of periodic waste dump closures. These changes will be addressed in the 2007 year end filings. The mine has prepared a 5 year mine plan based on the reserves presented in this report. The mine plan exploits only a portion of the measured and indicated resources and none of the inferred resources. Capital associated with the mine plan includes the necessary access to deeper resources to ensure that short term planning does not result in an impact to a longer mine life. The Capstone financial model provides an indication of the likely pre-tax operating returns, however there can be no assurances that the assumptions made in preparing these cash flow projections will prove accurate, and actual results may be materially greater or less than those contained in such projections. The Capstone financial model indicates the Net Present Value of the Cozamin mine at a discount rate of 8% as $US192.2M. This is based on average copper prices of $US 2.25/lb, average zinc prices of $US1.00/lb, average lead prices of $US0.60/lb and average silver prices of $US8.50/ounce. The authors consider the project robust. 22 RECOMMENDATIONS In order for the company to truly evaluate a long term mining plan (+10 years) it is highly recommended to complete another phase of infill drilling especially in the planned mining blocks to convert inferred resources to measured and indicated



resources and to potentially increase reserves. Infill drilling below Level 13 is also recommended as the ore handling and transportation to surface will become a critical operational and cost component over time. An internal ore handling study should be completed after the next phase of drilling. Power interruptions are frequent, and result in sanding of flotation cells. Normal time necessary to get the plant back to steady state averages two hours after a power interruption of 10 or more minutes. It is recommended that a standby generator be sized, purchased and installed to allow critical agitators and ventilation equipment to continue operating during power outages to minimize operating time with poor metallurgical recoveries and increase overall mine safety.



23 REFERENCES AND SOURCES OF INFORMATION Cardenas Vargas, J., Carrasco Centeno, M., Sanenz Reyes, R., and Macedo Palencia, R., 1993. Monografia Geologico-Minero del Estado de Durango. For Consejo de Recursos Minerales, Publication M-10e. Cardenas Vargas, J., Paraga Perez, J. de J., Merida Montiel, R., Macedo Palencia, R., and Rodriguez Salinas, J. de J., 1992. Geological-Mining Monograph of the State of Zacatecas for Consejo de Recursos Minerales, Publication M-2e. Christopher, P.A., 2003. Technical Report on Proposed Exploration Ventanas Gold-Silver Property, Durango State, Mexico. For Capstone Gold Corporation, dated May 9, 2003. Christopher, P.A., 2003. Technical Report on Proposed Exploration Cozamin project, Zacatecas State, Mexico. For Capstone Gold Corporation, dated November 25, 2003. Christopher, P.A., 2004. Technical Report on Phase 1 & 2 Exploration Cozamin project, Zacatecas State, Mexico. For Capstone Gold Corporation, dated December 14, 2004 (Revised and Corrected). Christopher, P.A and Giroux, G., 2005. Technical Report on Initial Resource Estimates (and Exploration Update) Cozamin project, Zacatecas State, Mexico. For Capstone Gold Corporation, dated October 31, 2005. Enriquez, E. 2003. Cozamin project. Executive Summary prepared for Capstone. Hawthorn, G., 2004. Review of Cozamin mine processing plant Zacatecas, Zacatecas State, Mexico. Prepared by Westcoast Mineral Testing Inc. for Capstone Gold Corporation, dated March 22, 2004. Johnson, B., Montante, A., Kearvell, G., Janzen, J., and Scammell, R., 1999. Geology and Exploration of San Nicolas Polymetallic (Zn-Cu-Au-Ag) Volcanogenic Massive Sulphide Deposit. In VMS and Carbonate-Hosted Pollymetallic Deposits of Central Mexico, ed. J.L. Jambor, B.C. and Yukon Chamber of Mines Cordilleran Roundup, January 1999. Moore, R., 1999. Geology and Development of the La Colorada Ag-Pb-Zn Deposit, Zacatecas State, Mexico. In VMS and Carbonate-Hosted Polymetallic Deposits of Central Mexico, ed. J.L. Jambor, B.C. and Yukon Chamber of Mines Cordilleran Roundup, January 1999. Ponce S., B.F., and Clark, K.F., 1988. The Zacatecas Mining District: A Tertiary Caldera Complex, Associated with Precious and Base Metal Mineralization. Econ. Geol. V. 83, No. 8, pp. 1668-1682.



Rodger, R., 2006. Feasibility Report on the Cozamin project. Prepared for Capstone Mining Corporation, dated February 17, 2006. SGS Lakefield Research Limited. An Investigation into Metallurgical Testing of Cozamin Ores. Prepared for Capstone Gold Corporation, dated March 10, 2005. Sinclair, A.J., 1974. “Applications of probability graphs in mineral exploration”, Spec. v. Association of Exploration Geochemists, 95 pages Smalwood, R. and Enriquez, E., 2003. San Dimas District, Durango, Mexico. Abstract for presentation at Cordilleran Roundup, pp.17-19. Smee, B.W., 2004. Results of an Audit of Reno and Elko Laboratories Nevada, USA. Prepared by Smee and Associates Consulting Ltd., Audit jointly funded by Capstone Gold Corporation and other Canadian listed companies, dated September 2004. Stone, J.G., 1956. Geology and Ore Deposits of the Cantera Mine, Zacatecas, Mexico. Econ. Geol., Vol. 51, No 1, pp. 80-95. Tan, G. 2005. Flotation tests on two sulphide composites from the Cozamin project, San Roberto Mine. Prepared by Process Research Associates Ltd. for Capstone Gold Corporation.



24 SIGNATURE, STAMP AND DATE Signed and stamped at Vancouver, B.C., on the 31std day of October 2007. “/s/ Michelle Stone” _______________________________ Michelle S. Stone, Ph.D., P.Geo. “/s/ Robert Barnes” _______________________________ Robert Bannister Barnes B.Sc., M.B.A., P.Eng. “/s/ Jenna Hardy” _______________________________ Jenna Hardy, M.Sc., M.B.A., P.Geo.



25 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES 25.1 Mining Operations 25.1.1 Introduction The Cozamin mine has completed 12 months of commercial operation since commissioning in June 2006 and ramping up to 1,000 tpd by September 2006. Between September 2006 and September 2007 Cozamin initiated a mine and plant expansion to 2,000 tpd by August 2007. The eventual capacity of the expansion designed at 2,200 tpd throughput was reached in October 2007. In the year between September 2006 and August 2007 the mine produced a mill feed of 431,406 tonnes of ore grading 1.50% copper, 1.61% zinc, 0.59% lead and 71 g/t silver. The average production rate was approximately 1,400 tpd during that period. During the same year, 5,300 m of development (ramps, drifts and raises) were completed including development to support stope mining, development to prepare crosscuts for Level 8 diamond drill stations and development to extend the ramp and shaft down to Level 11 (located 60 metres below Level 10). During the year a diamond drill program was conducted from the Level 8 drill stations using 4 underground diamond drill rigs completing approximately 21,000m of drilling. The information obtained from the drill program was used to update the resource estimate. The Cozamin mine operates on a two-10 hr shift per day basis 7 days per week using Cozamin employees and contractors. In the mine department, there are 74 Cozamin staff employees, 191 Cozamin hourly employees and 95 contactor employees. Three crews work the mine on a basis of 10 days of work and 5 days off rotation. Access to the Cozamin mine is by a 3.5m x 4m ramp from surface down to Level 11 (380 m below surface). The grade of the ramp is 15%. Personnel and equipment enter the mine by this ramp. A 2m x 4m vertical shaft extending 360m to loading pockets located below Level 10. The Level 10 hoisting capacity is 1,800 tpd using a 400 hp hoist with 2 skips of 4.2 tonnes per skip. The head frame and 400 tonne capacity surface ore bins are located 900 metres from the mill stockpile. Ore is transported by contractor 18 tonne trucks from the surface ore bins to the crusher coarse ore stockpile. Six 10 tonne trucks haul 400 tonnes of ore per day up the access ramp to the plant, as well as, waste for stope backfill and ore from the stopes to the shaft ore pockets. All waste produced from development is used as fill in cut and fill stopes. Waste rock is not hoisted out of the mine. The main air intake to the mine for ventilation is the San Roberto shaft. A 3m x 3m raise (280m deep) located on the eastern side of the active stoping area is used as ventilation exhaust. A 150 hp 72” exhaust fan extracts 120,000 cfm from the mine.



A second raise on the west side of the mine is being widened to 3m x 3m and will become a second exhaust with a 150 hp fan installed. Mine services including compressed air, water, electrical power and dewatering pump water lines are installed in a 1.5m Robbins service raise. Four Gardner Denver stationary air compressors with a capacity of 750 cfm each are installed on surface near the shaft and supply compressed air underground for drilling and pump operation. The mine has operating stopes on Levels 8, 9 and 10. Stopes on Level 11 are being developed. Transfer raises near the shaft transfer all ore to the Level 10 skip measuring and loading pocket. The dump pocket and grizzly on Level 10 is 42 cm x 42 cm and has a hydraulic hammer to help break the ore small enough to pass through the grizzly and to fit in the hoisting skips. The mine has a small shop on Level 8.5 and a larger shop on surface. A new larger shop is being constructed on Level 11. 25.1.2 Mining Method All of the mining reserves at the Cozamin mine occur within a 1.4 km section of the Mala Noche vein, which has an identified strike of more than 7 km. Prior to the re-commissioning, the mine was operated until 1998 by various companies. Previous mining methods included open stoping, mechanized cut and fill, shrinkage stoping and long hole stoping. Since commissioning, Cozamin has operated mechanized cut and fill stoping and long hole stoping mining methods. The stope widths average between 4 and 6 metres, although some sections approach 10m in width. The operating levels (8, 9, 10 and 11) are spaced 60m apart with the vein dipping an average of 60º to the north. Access to both the mechanized cut and fill stopes and the long hole stopes is from 4m x 4m footwall ramps, footwall drifts and crosscuts (+/-20%) from the footwall ramps. Figure 25.1 is a stope mining schematic showing stope blocks, access ramps, levels and ventilations raises. 25.1.2.1 Mechanized Cut and Fill Stoping During 2007, four stopes were operated as mechanized cut and fill stopes, two stopes on Level 8 and two stopes on Level 9. The stopes are accessed from ramps driven on the footwall of the vein. Crosscuts access the vein from the ramps. These crosscuts are driven at a grade of minus 20% and then pivoted to a plus 20% grade for the successive cuts in the cut and fill stope. The cuts are 1.5 to 3.0 metres in height. Single boom jumbos or jackleg drills are used for horizontal breast down cuts (4.2m wide by 3.0m advance with jumbo drills or 1.5m advance with jackleg drills) in some stopes. In other stopes, the cut is taken with jumbo or jackleg drills by drilling vertically into the back. Load haul dump (LHD) equipment will haul mineral to ore passes or load into trucks to haul to ore passes. Development waste or non economic ore is used for back fill in the cut and fill stopes. There are 3 waste passes from surface that can be



used if additional backfill if needed. A typical cut and fill stoping sequence and a cut and fill production cycle are shown in Figures 25.2 and 25.3.

Figure 25.1: Stope mining schematic.

Figure 25.2: Typical cut and fill stoping sequence.



Figure 25.3: Cut and fill production cycle. 25.1.2.2 Long Hole Stope Mining This is the main mining method used at the mine. It is considered when the width is more than 4 m, the ground very stable and the vein is very continuous. Eighty percent of the reserves will be mined by this method. In each stope area a 4m x 4m footwall ramp is driven to provide access by crosscut to the sublevels in the stope. Each stope has two sublevels which are used for drilling. Raises are driven (usually at the extremities of the stopes) to provide ventilation and access for air, water and power. Ore loading access is established by driving footwall haulage drifts on the sill level and then crosscuts ever 20 metres to access the vein. A drift is driven in the vein on the sill so complete access to loading level is established. Drilling using elector-hydraulic long hole drilling machines is done from sublevels with 12 cm diameter blast holes on a 1.5m x 1.8m pattern and up to 18m in length. A slot raise is driven using the same blast hole machine to establish a free face for blasting the long holes. ANFO and Nonels are used for blasting. Cable bolts and/or shotcrete are used when additional ground support is needed. The ore is loaded by LHDs which have remote controls for access into stopes. The LHDs haul to the ore passes or load trucks that haul to the ore passes or out of the mine. Five metre high sills and 10m vertical pillars are left for ground support. Development waste will also be used to backfill part of the mined out long hole stopes.



Figure 25.4: Long hole stoping – drilling process.

Figure 25.5: Long hole drilling, blasting and extraction process.

25.1.3 Mine Equipment The major mine equipment that is used to complete development, backfill and mining at Cozamin are shown in Table 25.1.



Table 25.1: Major mining equipment used at Cozamin. Equipment Number Specifications Load Haul Dump 4 yard 3 Toro 005 Load Haul Dump 6 yard 3 Toro 007 Jumbo Drills Single Boom 2 Tamrock - 14 ft booms Long hole drill 1 Simba 4.5" pneumatic hammer Long hole drill 1 Stopemate pneumatic top hammer Electrical High Pressure Air Compressor 2 Air compressors for long hole drills Haul trucks 6 Single axel conventional -10 tonne Service Vehicles 12 Various trucks and 4wd pickup trucks Backhoe and skid steer loader 2 JD backhoe with hydraulic hammer with Bobcat Jackleg drills 10 Pneumatic drill Main Ventilation Fan 2 150 Hp 120,000 cfm main fans De-watering pumps 10 Pumps various size 30 to 125 Hp Jaw Crusher 1 30 x 42 inch jaw crusher

25.1.4 Ore and Waste Handling Mined ore is transported to transfer raises from operating stopes directly by LHDs or by loading the ore into trucks and then hauling to the transfer raises. The transfer raises connect to the Level 10 loading pocket at the shaft. The Level 10 pocket consists of a dump pocket, grizzly with openings 42 cm x 42 cm, hydraulic rock breaking hammer to clear the grizzly, a 500 tonne storage bin, skip measuring bins and skip loading chutes. The measuring bins and skip loading is done with hydraulic cylinders which open and close chute gates. Figure 25.6 show a schematic of the loading pockets on Level 10. The production requirement for mill feed is 2,200 tpd. The hoist has a capacity limit of 1800 tpd, and 400 tpd are hauled up the ramp by 10 tonne haul trucks. All the development waste from operations is transferred by LHD or by truck to use as backfill in the cut and fill stopes. Excess waste will be dumped in empty long holes stopes. If more waste is need for backfill waste will transferred underground from surface by 3 Robbins raises located on the east, west and central parts of the mine. 25.1.5 Mine Ventilation One main ventilation fan is installed on the east side of the mine over a ventilation raise slashed to 3m x 3m. The fan is 150 hp with the fan blades set for 120,000 cfm which is exhausted from the mine. The ventilation raise extends to Level 9 and will eventually be extended to Level 12. A second exhaust fan of the same size will be installed on the west side of the mine. Air intake is by the main San Roberto shaft, Robbins raise and by the main ramp. Auxiliary fans are used for ventilation in drifts, ramps and raises.



Figure 25.6: Schematic diagram showing the Level 10.5 loading pocket (not to scale). 25.1.6 Pumping The main pump station is established on Level 8 with three 100 hp pumps (1 operating and 2 on standby). The pumps are in a clean water sump located adjacent to a primary decant sump where solids can be removed by LHDs. The pumping rate is approximately 20 l/s. Pump stations from Level 10 transfer water to this main pumps station. Auxiliary pumps in other part of the mine are also pumped to the main pump station or Level 10 pump station. Waters from the main pump stations are pumped to surface and are transferred to the tailing pond to be used as plant make-up waters. These waters are also used as mine drill water from a tank on surface. 25.1.7 Operations Expansion to 2,200 tpd Capital Project The expansion of operation to 2,200 tpd anticipates that the following projects will be completed in 2008: the central ramp down to Level 11.5, the extension of the shaft to



Level 11.5, and the construction of a crusher station and loading pockets. A pump stations and a maintenance shop on Level 11 are also in progress. A ramp on the west side of the mine to Level 9 is also in progress. 25.1.6 Milling The mill flowsheet consists of three-stage crushing, ball mill grinding and selective floatation of the ore to produce copper, lead and zinc concentrates, followed by thickening and filtering of the concentrates. Throughout 2006/7 daily treatment capacity was from 1,000 tpd to 2,200 tpd, with the mill operating 7 days a week. In 2006/7, the concentrator plant processed approximately 431,000 tonnes of ore and is expected to process 750,000 tonnes of ore in 2007/8. In 2006/7, Cozamin generated positive cash flow and income, with recorded cash costs of $0.54/lb copper net of by-product credits and including smelting, refining, transportation and all site costs. 25.2 Production and Development Schedules Mine production and development has been scheduled for a five year period commencing from January 1, 2008. The reserves were reconciled at the end of August 2007 and estimated production for September through December 2007 has been subtracted from the reserve base. The schedule is based on the mining reserves outlined in section 19.4 with detailed scheduling for each stope. The development estimate includes the development needed to advance the stopes, an exploration development allotment, capital development and ongoing general mine development. Waste is considered to be backfilled in cut and fill stopes or placed in empty long holes stopes. The mine schedule is designed to deliver 750,000 tonnes per year to the plant with a mix from more that 6 stopes to provide a smooth transition of the grades of copper, lead, zinc and silver. A summary of the production and development schedule is shown in Table 25.2 below.



Table 25.2: Summary of the production and development schedule: 2007-2012.

Year Width

(m) Tonnes

(t) Ag

(g/t)Cu (%)

Pb (%)

Zn (%)

Development Mining (m)

Development Capital (m)

2007

6.29 187,500

78

1.69

0.45

1.40 1,138

362

2008

6.30 750,000

79

1.72

0.46

1.39

4,764

1,236

2009

6.80 750,000

81

2.00

0.48

1.18

5,300

700

2010

6.61 750,000

84

2.25

0.45

1.06

5,100

900

2011

6.02 750,000

85

2.83

0.30

1.06

5,100

1,000

2012

6.17 537,468

85

3.57

0.23

1.14

1,871

900

Total 3,724,966 82 2.37 0.40 1.18 25,109 5,098 25.3 Mining Costs 25.3.1 Mine Operating Costs (in $US) The mine operating costs for the period September 1, 2006 to August 31, 2007 are summarized in Table 25.3 below. Operating costs were trending lower due increased mine production. Estimated 2008 and onward mine operating costs are also summarized. Table 25.3: Mine operating costs ($US/tonnes mined).

Area Sept to Aug 2007 $/tonne 2008 estimated $/tonne Stoping and Mine Development 11.86 10.10 General Development 0.59 0.32 Hoisting 0.78 0.83 Haulage to Plant 0.87 0.90 Mine Services 1.16 1.30 Mine Supervision/General 2.92 2.12 General Maintenance 1.46 1.47 Total 19.63 17.05 25.3.2 Capital Cost (in $US) The 2,200 tpd Expansion Project will be completed in 2008 with the following mine projects (Table 25.4).



Table 25.4: Mine operating costs ($US/tonnes mined). Area 2008 estimated expenditure

Ramp to Level 11.5, raises, extend shaft and sumps $500,000 Loading pockets and cable extension in shaft $315,000 Underground Crusher installation $130,000 Mine Shop excavation tools concrete electrical $250,000 Level 11 pumps station and pumps $120,000 West ramp to Level 9 $100,000 Total $1,415,000 25.3.3 Mill Operating Costs Mill operating costs are shown in section 18.5, Table 18.2. 25.3.4 General and Administrative Operating Costs The general and administrative operating costs for the period September 1, 2006 to August 31, 2007 are summarized in Table 25.5 below. The general and administrative operating costs were trending lower due increased mine production. Estimated 2008 and onward general and administrative operating costs are also summarized. Table 25.5: General and administrative operating costs ($US/tonne mined).

Area Sept to Aug 2007 $/tonne 2008 estimated $/tonne General Management 0.26 0.21 Engineering and Planning 0.36 0.34 Geology 0.62 0.52 Laboratory 0.45 0.39 Environmental and Security 0.98 0.87 Administration 1.20 0.72 Accounting and Sales 0.52 0.49 Total 4.39 3.39

25.4 General and Administrative - Infrastructure 25.4.1 Site Access The Cozamin mine is located 3.8 km from the city of Zacatecas and is accessible by an all weather gravel roads to the property. Large truck loads into site and shipments off- site are by 10 km of paved road from Veta Grande which is on major Mexican highway route. 25.4.2 Personnel The majority of mine and plant personnel and employees are from Zacatecas or one of the towns surrounding around the mine. Cozamin pays for food and housing to about 25 employees who are from outside of the area.



A general organizational chart for the mine is shown in Figure 25.7. The support staff works an 8 hour shift for 5.5 days per week. 25.4.3 Safety and Environmental Department The Safety Superintendent is the head of the Safety and Environmental Department He reports directly to the General Manager of Cozamin. The department includes 3 hourly employees, 9 security guards, 1 trainer/safety engineer, 1 environmental coordinator, 1 paramedic and 1 nurse. Cozamin has a mine rescue team which trains two times per month. The rescue team also trains with other mine rescue teams in the area and is prepared to provide assistance as needed at Cozamin and other mines. Conversely, Cozamin can receive assistance from other mine rescue teams if required. The mine has contracted with a private hospital located 8 kilometres from the mine site to provide the best medical assistance available in case of an incident. 25.4.4 Power Supply and Distribution Electrical energy is supplied to Cozamin by the Federal Electrical Commission, via a 13,200v line from Zacatecas. Power is distributed throughout the property at 13,200v. There are two main substations - one at the plant and one near the mine shaft. The plant substation services the process plant, crushing plant and the laboratory. The motors in the plant are primarily 480 volts except the two 14 ft x 12 ft primary grinding mills and the blowers for the flotation circuit which have 4,160 volt motors. The mine substation on surface supplies power to the hoist at 2,300v, 480v to the mine compressors and pumps to plant make up water tanks and 220/110v for offices and surface infrastructure. Power is distributed underground at 13,200v to two main substations on Levels 8 and 10. The main distribution voltage underground is 4,160v. The Level 8 substation has 13,200/4160v transformers and 4,160/480v transformers which supply power at 480v to operating equipment. The Level 10 substation is supplied at 13,200v, which is then transformed and supplied to the primary for distribution by a 1,500 kva 13,200/4160v transformer, and then to remote 500 kva substations at 4,160/440v transformers. The 13,200v supply line feeds the process water return pumps, the fresh water pump in the San Rafael mine and the main ventilation fans through 13,200/480v transformers. The total load for the mine and plant operating at 2,200 tpd is 4,300 kW.



Figure 25.7: Organization chart for the Cozamin mine.



25.4.5 Water Supply 25.4.5.1 Process Water The main fresh water supply for the plant is obtained from the San Roberto and San Rafael mines. Reclaimed water is also recycled from the tailings disposal area. There is a power line and a 4” steel waterline from the San Rafael mine to the process plant. Any water infiltrating from the toe of the tailing dam is returned to the dam and recycled to the plant. 25.4.5.2 Domestic Water Mine potable water is presently supplied by truck from Zacatecas. A water line is being constructed from a potable water line passing near the property from Zacatecas to Veta Grande. Two head tanks will be installed to supply potable water by gravity to all areas of the property. Mine drill water is supplied from the water settling tank near the San Roberto shaft. 25.5 Environmental Considerations This summary of the environmental impact and permitting requirements is based on the work undertaken for Capstone under the supervision of Nimbus Management Ltd., J. L. Hardy, P.Geo., Principal. The Cozamin project site 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. Prior to Capstone’s option of 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. With a view to re-opening the mine and expanding tonnage throughput to 1,000 tpd, Capstone completed the following to support permitting and regulatory approvals:

• 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 expansion 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 newly generated waste rock will be used as underground back fill, so over the longer term, mining will not generate new surface waste dumps, and will reduce significantly the volume of existing dumps. Additional mitigation measures involved both engineering design and operational approaches. Planning for closure design may include covering and encapsulation of the tailings impoundment if monitoring indicates this is required. 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. 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 has the general objective of leaving the land in a useful, stable and safe condition capable of supporting native plant life, provided appropriate wildlife habitat, maintaining watershed function and supported limited livestock grazing. 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. The preliminary closure cost estimate of $2.21M will be revised and updated to reflect evolving knowledge of specific site conditions and an understanding of the success of ongoing progressive rehabilitation, reclamation and closure activities, as well as current site costs. 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 14-Feb-06) for Phases 1 and 2 of the tailings dam expansion. Capstone subsequently filed application for a modification to this MIA, to include additional lands leased from the neighboring ejido for tailings dam expansion. This application was accepted on April 20, 2006. 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 MIA for the expanded operation was approved by SEMARNAT on 18 April 2007, and also has an operational term of ten years. As tailings dam expansion proceeds, one or more additional ETJ will be required to allow inclusion of additional lands for tailings dam and water management structures. The neighboring



ejido has approved the necessary purchase and lease of community lands, and no significant issues have been identified which would preclude approval of one or more additional ETJ. 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 (Procuraduria Federal de Proteccion 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 Unica Ambiental (LAU). An application is underway for modification of the LAU for the expanded tonnage. 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 the modification to the LAU. 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 Operacion Anual). An environmental management and monitoring program is currently underway and will be ongoing. Data collected will be 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 Auditoria 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. It also regularly consults industry representatives about possible changes to the program. 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 process consists of three stages: planning, assessment, and post-audit activities. The audit results in an action plan that is included in the Environmental Compliance Agreement to be signed by PROFEPA and the company concerned. Under this process, the Cozamin mine will be able to receive certification under the Clean Industry Program. As a first step, the Cozamin operation is to undergo a rigourous 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 PROPEPA, the Company will identify areas for improvement, and develop a detailed Action Plan (with estimated costing) to achieve compliance over a period which may extend up to two years. 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 certificate is valid for two years and can be re-authenticated and renewed by an additional audit. 25.6 Financial Analysis The operating cost estimates used for the 5 year Capstone mine plan have been derived from the 2007 actual operating costs with particular emphasis on the last quarter of operation when the plant operated at 2,200 tpd. 25.6.1 Mine Operating Costs The mining operating costs are detailed in section 25.3.1. 25.6.2 Plant Operating Costs Processing costs are detailed in section 18.5. 25.6.3 General and Administrative Costs These costs are detailed in section 25.3.3.



25.6.4 Total Operating Costs The total operating costs used in the five year plan are shown in Table 25.6 which includes average costs through the last 12 months of operation and projected costs going forward based on 2,200 tpd. Table 25.6: Operating costs - actual and estimated. Actual $US/tonne Estimated $US/tonne Mining 19.63 17.05 Processing 13.90 12.90 G&A 4.39 3.39 Total 37.92 33.34 Mine capital development requirements are included in mining costs with the exception of projects outlined in section 25.3.2 which include the last items which need to be completed and formed part of the capital budget for the expansion to 2,200 tpd. Total costs are $1.415M to be spent by the end of 2008. An allowance of $0.5M has been included as sustaining capital on an annual basis. 25.7 Capstone 5 Year Mine Plan Capstone has prepared a reserve statement based on measured and indicated resources as presented in section 19.4. This result is restated in Table 25.7 below and formed the basis of the financial model. Table 25.7: Cozamin 2007 proven and probable reserves. Classification Tonnes Cu (%) Ag (g/t) Pb (%) Zn (%) Proven 1,809,719 2.32 84 0.45 1.17 Probable 1,915,248 2.42 81 0.34 1.19 Total 3,724,967 2.37 82 0.40 1.18 Capstone has prepared a detailed stoping and development schedule to achieve and sustain the planned production as shown in Table 25.8.



Table 25.8: Summary of the production and development schedule: 2007-2012.

Year Width

(m) Tonnes

(t) Ag

(g/t)Cu (%)

Pb (%)

Zn (%)

Development Mining (m)

Development Capital (m)

2007

6.29 187,500

78

1.69

0.45

1.40 1,138

362

2008

6.30 750,000

79

1.72

0.46

1.39

4,764

1,236

2009

6.80 750,000

81

2.00

0.48

1.18

5,300

700

2010

6.61 750,000

84

2.25

0.45

1.06

5,100

900

2011

6.02 750,000

85

2.83

0.30

1.06

5,100

1,000

2012

6.17 537,468

85

3.57

0.23

1.14

1,871

900

Total 3,724,966 82 2.37 0.40 1.18 25,109 5,098 25.7.1 Processing Issues Recovery to date has been less than planned with copper currently averaging 87.5% recovery versus a planned 89%. Concentrate grade ranges from 22-23% copper, also below planned grades of 24-25%. The plant is examining water quality issues in reclaim water to increase both recovery and concentrate grade. Test work in the laboratory indicates recovery and grade can be produced at better than plan on potable water. The financial model assumes only the average metallurgical performance to date and does not include proposed improvements. 25.7.2 Capital Expenditures During 2006/7, capital expenditures were approximately $16.2M and consisted of $5.0M for surface and underground diamond drilling, $11.2M for mine and mill expansion from 1,000 tpd to 2,200 tpd. Capstone has budgeted $2.0M for 2008 capital expenditures at the Cozamin mine (Table 25.9) consisting of $1.0M for mine development and equipment, $0.5M for expansion of the tailings dam, and $0.5M for mill upgrades. Mine development capital has been included in the mine operating cost estimate with the exception of $1.4Mto be spent by the end of 2008 related to the expansion. Sustaining capital has been estimated at $0.5M per year. Due to the short life of the plan based on the current proven and probable reserves, the only other major capital requirement is for tailings dam capacity increase to accommodate the total tonnage.



Table 25.9: Annual capital expenditures ($USM). 2007 2008 2009 2010 2011 2012 Mine 0.4 1.0 Tailings 0.5 1.0 Sustaining 0.2 0.5 0.5 0.5 0.5 0.5 Total 0.6 2.0 1.5 0.5 0.5 0.5 25.7.3 Marketing In 2006/7, zinc and lead concentrates were sold to Glencore International AG under a contract which is valid through 2008. Copper concentrate was sold 75% to Glencore International AG and 25% to Trafigura Ltd. In 2008, concentrates will be split 50%/50% between the two metal traders. In 2009, 100% of production has been sold to Trafigura Ltd. Payable silver contained in all concentrate was sold to Silverstone on April 11, 2007 for the following 10 years. Capstone received $US44M on signing, and an ongoing payment of $US4.00 for every ounce delivered. The copper concentrate produced is highly marketable in Mexico due to the low arsenic and antimony impurity levels. The lead concentrate is also highly marketable with high lead and silver content and low levels of impurities. The zinc concentrate is a low grade concentrate with high iron content but to date Capstone has not been able to secure contracts for its sale. 25.7.4 Mine Plan A financial model has been created for a 5 year plan based on reserves statements, mine production and development schedules, metallurgical recoveries, operating costs, and capital expenditures as outlined in this report. The model provides an indication of the likely pre tax operating returns, however there can be no assurances that the assumptions made in preparing these cash flow projections will prove accurate, and actual results maybe materially greater or less than those contained in such projections. Metal prices used in the 5 year plan were the same as those used in the cut off calculations, and are shown below in table 25.10: Table 25.10: Metal prices.

Metal 2007-2012 Silver $US per ounce 8.50

Copper $US per pound 2.25 Lead $US per pound 0.60 Zinc $US per pound 1.00

The salient details of the financial model are given in Table 25.11.



Table 25.11: Results of financial modelling ($US). 2007-2012 Production Tonnes Processed 3,724,966 Copper Produced (t) 79,400 Zinc Produced (t) 22,700 Lead Produced (t) 9,350 Silver Produced (oz) 7,300,000 NSR Revenue Total Revenue 385,700,000 Operating Costs Mining Costs 63,500,000 Processing Costs 48,000,000 G&A Costs 12,600,000 Total Operating Costs 124,100,000 Royalty 3 % NSR 11,550,000 Net Operating Profit 250,000,000 Capital Costs Total Capital Costs 5,600,000 Pre Tax Free Cash Flow 244,400,000 The model has been prepared on an all equity basis. The base case Net Present Value (NPV) at various discount rates is shown in table 25.12. The model does not discount 2007 costs and revenues, but all other cash flows are discounted to the beginning of 2008. Table 25.12: Capstone base case financial model NPV. Discount Rate NPV $USM

0% 244.4 8% 192.2 15% 159.5

25.7.5 Sensitivity Analysis Simplistic sensitivity analysis was performed with copper price and operating costs both rising and falling. Since capital expenditures for construction and expansion have been completed, no sensitivity to capital costs was performed. Table 25.13 illustrates the analysis. Table 25.13: NPV sensitivity analysis ($USM). NPV Base Case Copper Price

+ 10% Copper Price

– 10% Operating

Costs + 10% Operating

Costs – 10%0% 244.4 280.9 207.8 231.9 256.8 8% 192.2 221.1 163.2 182.0 202.2

15% 159.5 183.6 135.4 150.8 168.1 As expected, the project is most sensitive to the copper price.



26 CERTIFICATES



CERTIFICATE To accompany the report entitled

“Technical Report on the Cozamin project, Zacatecas State, Mexico” dated 31 October 2007.

I, Michelle S. Stone, of 1988 Rivergrove Place, North Vancouver, British Columbia, do hereby certify that:

1) I am a Senior Geologist with Capstone Mining Corp. 1980-1055 West Hastings

Street, Vancouver, British Columbia.

2) I hold a B.Sc. (1994) from McMaster University (Ontario), an M.S. (1996) from the University of Alabama (Alabama), and a Ph.D. (2005) from the University of Western Australia (Australia).

3) I am a Professional Geoscientist and a member in good standing of the

Association of Professional Engineers and Geoscientists of British Columbia since 2006 (registered #30601).

4) I have practiced my profession continuously since 1994.

5) I have read the definition of “qualified person” set out in National Instrument

43-101 and certify that by reason of education, experience, and affiliation with a professional association, I meet the requirements of a Qualified Person (not independent) as defined in National Policy 43-101.

6) This report titled “Technical Report on the Cozamin Project, Zacatecas State,

Mexico” and dated 31 October 2007 (the “Technical Report”) is based on a study of the data and literature available on the Mala Noche vein, Cozamin project and twelve (12) site visits from February 2006 through October 2007. I am responsible for sections 1 - 19.2, 22 and 23 of the Technical Report.

7) As of the date of this certificate, to the best of my knowledge, information and

belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report, not misleading.

8) I have read National Instrument 43-101 and Form 43-101F1, and the Technical

Report has been prepared in compliance with that instrument and form.

9) I consent to the filing of the Technical Report by Capstone Mining Corp. with any stock exchange and other regulatory authority, and any publication of the Technical Report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible to the public.



Dated at Vancouver, British Columbia, the 31st day of October 2007. “/s/ Michelle Stone” _______________________________ Michelle S. Stone, Ph.D., P.Geo.




“Technical Report on the Cozamin Project, Zacatecas State, Mexico” dated 31 October 2007.

I, Robert Bannister Barnes, of 1220 Fifth St., Spearfish, South Dakota, USA, 57783 do hereby

certify that:

1) I am Vice President of Operations for Capstone Mining Corp., located at 1980 - 1055 West Hastings Street, Vancouver, British Columbia.

2) I graduated from the Montana College of Mineral Science and Technology in 1970

with a B.Sc. Honours in Mining Engineering. In 1994, I graduated with an M.B.A. from the University of South Dakota.

3) I am a member in good standing of the Association of Professional Engineers and

Geoscientists of the Province of British Columbia.

4) I have practised my profession continuously since 1970. I have had over 30 years experience in engineering, management and supervision in all aspects mine design and operation of base and precious metal deposits.

5) I have read the definition of “qualified person” as set out in National Instrument 43-

101 and certify that by reason of education, experience, and affiliation with a professional association, I meet the requirements of a Qualified Person (not independent) as defined in National Instrument 43-101.

6) This report titled “Technical Report on the Cozamin Project, Zacatecas State,

Mexico” and dated 31 October 2007 (the “Technical Report”) is based on a study of the data and literature available on the Mala Noche Vein. I am responsible for sections 19.3 - 25.4 and 25.6 - 25.7 of the Technical Report. I have been on site frequently over the past 3 years.

7) As of the date of this certificate, to the best of my knowledge, information and

belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

8) I have read National Instrument 43-101 and Form 43-101F1, and the Technical

Report has been prepared in compliance with that instrument and form. 9) I consent to the filing of the Technical Report by Capstone Mining Corp. with

any stock exchange and other regulatory authority, and any publication of the Technical Report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible to the public.



Dated at Vancouver, British Columbia, the 31st day of October 2007.

“/s/ Robert Barnes” _______________________________ Robert Bannister Barnes B.Sc., M.B.A., P.Eng.




“Technical Report on the Cozamin project, Zacatecas State, Mexico” dated 31 October 2007.

I, Jenna L. Hardy, , of 535 East Tenth St, North Vancouver, British Columbia, Canada, V7L 2E7, do hereby certify that:

1) I am a Principal of Nimbus Management Ltd, of the above address, a firm employed to carry out environmental consulting on behalf of clients in the mining and exploration sectors.

2) I graduated from the University of Toronto with a B.Sc. - Specialist in Geology (1974) and an M.Sc. (1978), and hold an M.B.A. (1988) from Simon Fraser University, British Columbia.

3) I am a Professional Geoscientist and a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#109549).

4) I have practised my profession continuously since 1978, and have over twenty five years of experience in geological and environmental aspects of mineral exploration, project development and mine operations. I have read the definition of “qualified person” set out in National Instrument 43-101 and certify that by reason of education, experience, and affiliation with a professional association, I meet the requirements of a Qualified Person as defined in National Policy 43-101.

5) I am responsible for the preparation of the sub-section of section 25.5 entitled “Environmental Considerations” of the Technical Report titled “Technical Report on the Cozamin Project, Zacatecas State, Mexico” and dated 31 October 2007 (“Technical Report”). This text is based on observations and discussions during site visits to the Cozamin mine in Zacatecas, Mexico carried out over the period 2004 to the present, as well as environmental investigations carried out by Clifton Associates Ltd., Natural Environmental S.C., and mine environmental staff.

6) 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.

7) I am independent of the issuer applying all the tests in section 1.4 of the National Instrument 43-101.

8) I have read National Instrument 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

9) I consent to the filing of the Technical Report by Capstone Mining Corp. with any stock exchange and other regulatory authority, and any publication of the Technical Report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible to the public.



Dated at Vancouver, British Columbia, the 31st day of October 2007. “/s/ Jenna Hardy” _______________________________ Jenna Hardy, M.Sc., M.B.A., P.Geo.



VOLUME I

APPENDIX FOR:

TECHNICAL REPORT ON THE COZAMIN PROJECT, ZACATECAS STATE, MEXICO

BY:

MICHELLE S. STONE, Ph.D., P.Geo.

AND

ROBERT B. BARNES, MASc., M.B.A., P.Eng.

AND

JENNA HARDY, M.Sc, M.B.A., P.Geo.

DATED OCTOBER 31st, 2007



APPENDIX 1 QA/QC ANALYSIS PHASES IV AND V

SHOWING

BLANKS FOR GOLD, SILVER, COPPER, LEAD AND ZINC

COMPANY STANDARDS: LOW GRADE STANDARD 4757

LOW GRADE STANDARD 4787-2 MEDIUM GRADE STANDARD 4759

ALSCHEMEX GOLD, SILVER, COPPER, LEAD AND ZINC STANDARDS: BM-44 CCu-1c

GBM302-10 GBM303-1

GBM306-12 GBM398-4c GBM399-5 GBM997.6c

IGS 42 OXQ47 PB-105 SP27 SQ27

PREPARATION AND ANALYTICAL CHECKS

Graphs in this appendix show up to 3 sets of lines: Red: maximum and minimum expected value Blue: the mean of the sample data plus and/or minus 1 standard deviation Yellow: the mean of the sample data plus and/or minus 2 standard deviations



LIST OF STANDARDS AND ACCEPTED VALUES

Au (ppb) Ag (g/t) Cu (%) Pb (%) Zn (%) Cozamin LOW GRADE STANDARD 4759

Lower Limit 189.5 21.68 0.52 0.00 0.63 Mean 200.3 24.42 0.55 0.01 0.65

Upper Limit 211.1 27.15 0.58 0.03 0.68 LOW GRADE STANDARD 4787-2 Lower Limit 14.96 28.88 0.62 0.11 0.62

Mean 98.50 35.55 0.66 0.12 0.68 Upper Limit 182.04 42.21 0.70 0.13 0.74 MEDIUM GRADE STANDARD 4757 Lower Limit 61.0 52.58 1.25 0.02 0.82

Mean 70.2 60.04 1.31 0.03 0.86 Upper Limit 79.5 67.50 1.37 0.05 0.89 HIGH GRADE STANDARD 4787 Lower Limit 92.9 188.13 3.29 0.15 2.65

Mean 109.4 212.46 3.45 0.17 2.78 Upper Limit 126.0 236.79 3.61 0.19 2.91

Au (ppb) Ag (g/t) Cu (%) Pb (%) Zn (%) ALSChemex BM-44

Lower Limit 0.06 <0.01 <0.01 Mean 0.07 <0.01 0.01

Upper Limit 0.08 <0.01 0.02 CCu-1c

Lower Limit 24.72 0.337 3.85 Mean 25.62 0.353 3.99

Upper Limit 26.52 0.369 4.13 GBM302-10

Lower Limit 0.41 5.38 13.75 Mean 0.44 5.59 14.28

Upper Limit 0.46 5.79 14.80 GBM303-1

Lower Limit 0.041 22.8 2.76 Mean 0.044 23.65 2.88

Upper Limit 0.046 24.5 2.99 GBM306-12

Lower Limit 4 1.425 2.58 1.97 Mean 5 1.480 2.68 2.05

Upper Limit 6 1.535 2.78 2.13



Au (ppb) Ag (g/t) Cu (%) Pb (%) Zn (%) ALSChemex GBM398-4c

Lower Limit 46 0.374 1.12 0.48 Mean 48.5 0.389 1.17 0.51

Upper Limit 51 0.404 1.22 0.54 GBM399-5

Lower Limit 2.83 2.03 0.91 Mean 2.95 2.12 0.95

Upper Limit 3.06 2.20 0.99 GBM997-6c

Lower Limit 0.3664 24.0337 15.6256 Mean 0.38181 24.9095 16.1944

Upper Limit 0.3972 25.7853 16.7632 IGS42

Lower Limit 72.96 Mean 74.84

Upper Limit 76.72 OXQ47

Lower Limit 46.3 132 Mean 49.8 147.5

Upper Limit 53.3 163 PB-105

Lower Limit 0.60 3.50 5.44 Mean 0.63 3.64 5.65

Upper Limit 0.66 3.78 5.86 SP27

Lower Limit 16.80 49 Mean 18.1 58

Upper Limit 19.40 67 SQ27

Lower Limit 46.8 137 Mean 50.4 152.5

Upper Limit 54.0 168



COZAMIN BLANKS

INTERNAL BLANKS - AU

0

0.01

0.02

0.03

0.04

0.05

0.06

0.07

0 20 40 60 80 100

Number of Samples

Au(g

/t)

INTERNAL BLANKS - AG

0

2

4

6

8

10

12

0 20 40 60 80 100

Number of Samples

Ag(g

/t)

INTERNAL BLANKS - CU

0

0.02

0.04

0.06

0.08

0.1

0.12

0.14

0.16

0.18

0.2

0 20 40 60 80 100

Number of Samples

Cu(%

)



INTERNAL BLANKS - PB

0

0.005

0.01

0.015

0.02

0.025

0 20 40 60 80 100

Number of Samples

Pb(%

)

INTERNAL BLANKS - ZN

0

0.01

0.02

0.03

0.04

0.05

0.06

0.07

0.08

0.09

0 20 40 60 80 100

Number of Samples

Zn (%



ALSCHEMEX BLANKS

ALSCHEMEX BLANKS - AU

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 200 400 600 800 1000 1200 1400

Number of Samples

Au(g

/t)

INTERNAL BLANKS - AG

0

1

2

3

4

5

6

7

8

9

0 200 400 600 800 1000 1200 1400

Number of Samples

Ag(g

/t)

ALSCHEMEX BLANKS - CU

0

0.0005

0.001

0.0015

0.002

0.0025

0 20 40 60 80 100 120 140

Number of Samples

Cu(%

)



ALSCHEMEX BLANKS - PB

0

0.005

0.01

0.015

0.02

0.025

0 20 40 60 80 100 120 140

Number of Samples

Pb(%

)

ALSCHEMEX BLANKS - ZN

0

0.002

0.004

0.006

0.008

0.01

0.012

0 20 40 60 80 100 120 140

Number of Samples

Zn(%

)



STANDARD 4757

CZ STD 4757 - AU

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0.40

0.45

0.50

0 5 10 15 20 25 30

Number of Samples

Au(g

/t)

CZ STD 4759 - AG

0

20

40

60

80

100

120

140

160

180

0 5 10 15 20 25 30

Number of Samples

Ag(g

/t)

CZ STD 4757 - CU

0.0

0.5

1.0

1.5

2.0

2.5

0 5 10 15 20 25 30

Number of Samples

Cu(%

)



CZ STD 4757 - PB

0

0.01

0.02

0.03

0.04

0.05

0.06

0 5 10 15 20 25 30

Number of Samples

Pb(%

)

CZ STD 4757 - ZN

0.74

0.76

0.78

0.8

0.82

0.84

0.86

0.88

0.9

0.92

0 5 10 15 20 25 30

Number of Samples

Zn(%

)



STANDARD 4787-2

CZ STD 4787-2 - AU

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

0.4

0 10 20 30 40 50 60Number of Samples

Au(g

/t)

CZ STD 4787-2 - AG

0

10

20

30

40

50

60

0 10 20 30 40 50 60

Number of Samples

Ag(g

/t)

CZ STD 4787-2 - CU

0.00

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0.80

0 10 20 30 40 50 60

Number of Samples

Cu(%

)



CZ STD 4787-2 - PB

0

0.02

0.04

0.06

0.08

0.1

0.12

0.14

0 10 20 30 40 50 60

Number of Samples

Pb(%

)

CZ STD 4787-2 - ZN

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

0 10 20 30 40 50 60

Number of Samples

Zn(%

)



STANDARD 4759

CZ STD 4759 - AU

0

0.05

0.1

0.15

0.2

0.25

0.3

0.35

0.4

0.45

0.5

0 2 4 6 8 10 12

Number of Samples

Au(g

/t)

CZ STD 4759 - AG

0

5

10

15

20

25

30

35

40

45

0 2 4 6 8 10 12

Number of Samples

Ag(g

/t)

CZ STD 4759 - CU

0.51

0.52

0.53

0.54

0.55

0.56

0.57

0.58

0.59

0 2 4 6 8 10 12

Number of Samples

Cu(%

)



CZ STD 4759 - PB

0

0.02

0.04

0.06

0.08

0.1

0.12

0 2 4 6 8 10 12

Number of Samples

Pb(%

)

CZ STD 4759 - ZN

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0 2 4 6 8 10 12

Number of Samples

Zn(%

)



BM-44

ALS STD BM44 - CU

0

0.01

0.02

0.03

0.04

0.05

0.06

0.07

0.08

0.09

0 1 2 3 4

Number of Samples

Cu

(%

ALS STD BM44 - ZN

0

0.005

0.01

0.015

0.02

0.025

0 1 2 3 4

Number of Samples

Zn (%



CCu-1c

ALS STD CCU-1c - CU

23.5

24

24.5

25

25.5

26

26.5

27

0 1 2 3

Number of Samples

Cu (%

ALS STD CCU-1c - PB

0.335

0.34

0.345

0.35

0.355

0.36

0.365

0.37

0.375

0 1 2 3

Number of Samples

Pb (%

ALS STD CCU-1c - ZN

3.8

3.85

3.9

3.95

4

4.05

4.1

4.15

0 1 2 3

Number of Samples

Zn (%



GBM302-10 ALS STD GBM302-10 - CU

0.415

0.42

0.425

0.43

0.435

0.44

0.445

0.45

0.455

0 10 20 30 40 50

Number of Samples

Cu

(%

ALS STD GBM302-10 - PB

5.35

5.4

5.45

5.5

5.55

5.6

5.65

5.7

5.75

5.8

5.85

0 10 20 30 40 50

Number of Samples

Pb (%

ALS STD GBM302-10 - ZN

13.6

13.8

14

14.2

14.4

14.6

14.8

15

15.2

0 10 20 30 40 50

Number of Samples

Zn (%



GBM303-1

ALS STD GBM303-1 - CU

0.03

0.035

0.04

0.045

0.05

0 1 2 3 4 5

Number of Samples

Cu

(%


22.6

22.8

23

23.2

23.4

23.6

23.8

24

24.2

24.4

24.6

0 1 2 3 4 5

Number of Samples

Pb

(%


2.55

2.6

2.65

2.7

2.75

2.8

2.85

2.9

2.95

3

3.05

0 1 2 3 4 5

Number of Samples

Zn (%



GBM306-12


1.4

1.45

1.5

1.55

1.6

0 10 20 30 40 50 60 70

Number of Samples

Cu (%


2.5

2.55

2.6

2.65

2.7

2.75

2.8

0 10 20 30 40 50 60 70

Number of Samples

Pb (%


1.96

1.98

2

2.02

2.04

2.06

2.08

2.1

2.12

2.14

0 10 20 30 40 50 60 70

Number of Samples

Zn (%



GBM398-4c

ALS STD GBM398-4c - CU

0.3

0.35

0.4

0.45

0 10 20 30

Number of Samples

Cu (%

ALS STD GBM398-4c - PB

1.1

1.12

1.14

1.16

1.18

1.2

1.22

1.24

0 10 20 30

Number of Samples

Pb (%

ALS STD GBM398-4c - ZN

0.47

0.48

0.49

0.5

0.51

0.52

0.53

0.54

0.55

0 10 20 30

Number of Samples

Zn (%



GBM399-5


2.75

2.8

2.85

2.9

2.95

3

3.05

3.1

0 10 20 30 40 50 60 70

Number of Samples

Cu (%


2

2.05

2.1

2.15

2.2

2.25

2.3

2.35

2.4

0 10 20 30 40 50 60 70

Number of Samples

Pb

(%


0.9

0.91

0.92

0.93

0.94

0.95

0.96

0.97

0.98

0.99

1

0 10 20 30 40 50 60 70

Number of Samples

Zn (%



GBM997-6c

ALS STD GBM997-6c - CU

0.03

0.08

0.13

0.18

0.23

0.28

0.33

0.38

0.43

0 1 2 3 4 5

Number of Samples

Cu

(%

ALS STD GBM997-6c - PB

23

23.5

24

24.5

25

25.5

26

0 1 2 3 4 5

Number of Samples

Pb

(%

ALS STD GBM997-6c - ZN

15.2

15.4

15.6

15.8

16

16.2

16.4

16.6

16.8

17

0 1 2 3 4 5

Number of Samples

Zn (%



IGS 42

ALS STD IGS 42 - PB

72.5

73

73.5

74

74.5

75

75.5

76

76.5

77

0 1 2 3 4 5

Number of Samples

Pb

(%



OXQ47

ALS STD OXQ47 - AU

46

47

48

49

50

51

52

53

54

55

0 10 20 30 40 50 60 70 80 90 100 110 120 130 140 150 160 170 180 190 200

Number of Samples

Au

(g/t

ALS STD OXQ47 - AG

125

130

135

140

145

150

155

160

165

170

0 10 20 30 40 50 60 70 80 90 100 110 120 130 140 150 160 170 180 190 200

Number of Samples

Ag

(g/t



PB-105

ALS STD PB-105 - CU

0.55

0.575

0.6

0.625

0.65

0.675

0.7

0 5 10 15 20

Number of Samples

Cu

(%

ALS STD PB-105 - PB

3

3.1

3.2

3.3

3.4

3.5

3.6

3.7

3.8

3.9

4

0 5 10 15 20

Number of Samples

Pb

(%

ALS STD PB-105 - ZN

5

5.1

5.2

5.3

5.4

5.5

5.6

5.7

5.8

5.9

6

0 5 10 15 20

Number of Samples

Zn (%



SP27

ALS STD SP27 - AU

16

17

18

19

20

0 50 100 150 200 250 300 350 400 450

Number of Samples

Au

(g/t

ALS STD SP27 - AG

40

45

50

55

60

65

70

75

80

0 50 100 150 200 250 300 350 400 450

Number of Samples

Ag

(g/t



SQ27

ALS STD SQ27 - AU

46

47

48

49

50

51

52

53

54

55

0 10 20 30

Number of Samples

Au

(g/t



CORE DUPLICATES

CORE DUPLICATES - AUR2 = 0.5384

0.0

0.2

0.4

0.6

0.8

1.0

1.2

0.0 0.2 0.4 0.6 0.8 1.0 1.2 1.4 1.6

Primary (ALS) sample Au (g/t)

Dupl

icate

(SGS

)Au

(g/t)

CORE DUPLICATES - AGR2 = 0.915

0

500

1000

1500

2000

2500

3000

0 200 400 600 800 1000 1200 1400 1600

Primary (ALS) sample Ag (g/t)

Dupl

icate

(SGS

)Ag

(g/t)

CORE DUPLICATES - CUR2 = 0.9338

0

2

4

6

8

10

12

14

16

0 2 4 6 8 10 12 14

Primary (ALS) sample Cu (g/t)

Dupl

icat

e(SG

S)Cu

(%)



CORE DUPLICATES - PBR2 = 0.9634

0

5

10

15

20

25

0 5 10 15 20 25 30

Primary (ALS) sample Pb (g/t)

Dupl

icat

e(SG

S)Pb

(%)

CORE DUPLICATES - ZN

R2 = 0.9308

0

2

4

6

8

10

12

14

16

18

20

0 2 4 6 8 10 12 14 16 18 20

Primary (ALS) sample Zn (g/t)

Dupl

icat

e(S

GS)Z

n(%

)



PULP PREPARATION DUPLICATES

CZ PULP DUPLICATES - AU

0

0.1

0.20.3

0.4

0.5

0.6

0.7

0.8

0.9

1

0 20 40 60 80 100Number of Samples

Au(g

/t)

CZ PULP DUPLICATES - AUR2 = 0.9942

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 0.9

Original Au (g/t)

Dup

licat

e A

u (g



CZ PULP DUPLICATES - AG

0

200

400

600

800

1000

1200

1400

1600

0 20 40 60 80 100

Number of Samples

Ag(g

/t)

CZ PULP DUPLICATES - AGR2 = 0.9973

0

200

400

600

800

1000

1200

1400

0 200 400 600 800 1000 1200 1400

Original Ag (g/t)

Dup

licat

e A

g (g



CZ PULP DUPLICATES - CU

0

5

10

15

20

25

0 20 40 60 80 100

Number of Samples

Cu(%

)

CZ PULP DUPLICATES - CUR2 = 0.9841

0

2

4

6

8

10

12

0 2 4 6 8 10 12

Original Cu (%)

Dup

licat

e C

u (



CZ PULP DUPLICATES - PB

0

5

10

15

20

25

30

0 20 40 60 80 100

Number of Samples

Pb(%

)

CZ PULP DUPLICATES - PBR2 = 0.9707

0

5

10

15

20

25

30

0 5 10 15 20 25 30

Original Pb (%)

Dup

licat

e P



CZ PULP DUPLICATES - ZN

0

5

10

15

20

25

30

0 20 40 60 80 100Number of Samples

Zn(%

)

CZ PULP DUPLICATES - ZNR2 = 0.9167

0

2

4

6

8

10

12

14

16

0 2 4 6 8 10 12 14 16

Original Zn (%)

Dup

licat

e Zn

(



ANALYTICAL LAB PULP CHECKS

ALSCHEMEX LAB CHECKS - GOLDR2 = 0.974

0

0.1

0.2

0.3

0.4

0.5

0.6

0.7

0.8

0.9

0 0.05 0.1 0.15 0.2 0.25 0.3 0.35 0.4 0.45 0.5Original Au (g/t)

Dupl

icat

eAu

(g/t)

ALS LAB CHECKS - AU

0

0.1

0.2

0.3

0.4

0.5

0.6

0 10 20 30 40 50 60

Number of Samples

Au(g

/t)



ALSCHEMEX LAB CHECKS - SILVERR2 = 0.9792

0

200

400

600

800

1000

1200

1400

0 50 100 150 200 250

Original Ag (g/t)

Dupl

icate

Ag(g

/t)

ALS CHECKS - AG

0

50

100

150

200

250

0 10 20 30 40 50 60

Number of Samples

Ag(g

/t)



ALSCHEMEX LAB CHECKS - COPPERR2 = 0.9998

0

1

2

3

4

5

6

7

8

9

10

0 1 2 3 4 5 6 7 8 9 10

Original Cu (%)

Dupl

icat

eCu

(%)

ALS LAB CHECKS - CU

0

1

2

3

4

5

6

7

8

9

10

0 10 20 30 40 50 60

Number of Samples

Cu(%

)



ALSCHEMEX LAB CHECKS - LEADR2 = 1

0

5

10

15

20

25

30

0 2 4 6 8 10 12 14 16

Original Pb (%)

Dupl

icat

ePb

ALS LAB CHECKS - PB

0

2

4

6

8

10

12

14

16

0 10 20 30 40 50 60

Number of Samples

Pb(%

)



ALSCHEMEX LAB CHECKS - ZINCR2 = 0.9998

0

2

4

6

8

10

12

0 1 2 3 4 5 6 7 8 9 10

Original Zn (%)

Dupl

icat

eZn

(%)

ALS LAB CHECKS - ZN

0

2

4

6

8

10

12

0 10 20 30 40 50 60

Number of Samples

Zn(%

)

ALSCHEMEX LAB CHECKS - ZINCR2 = 0.9998

0

2

4

6

8

10

12

0 1 2 3 4 5 6 7 8 9 10

Original Zn (%)

Dupl

icat

eZn

(%)



ANALYTICAL REPEATS

ALSCHEMEX REPEAT ASSAY - GOLDR2 = 1

0

10

20

30

40

50

0 10 20 30 40 50

Original Au (g/t)

Dup

licat

e A

u (g

ALSCHEMEX REPEAT ASSAY SILVER

R2 = 0.9972

0

200

400

600

800

1000

1200

1400

1600

0 200 400 600 800 1000 1200 1400 1600

Original Ag (g/t)

Dup

licat

e A

g (g

.



APPENDIX 2 QA/QC ANALYSIS CHIP CHANNEL SAMPLES

SHOWING

BLANKS FOR SILVER, COPPER, LEAD AND ZINC

COMPANY STANDARDS: LOW GRADE STANDARD 4759

LOW GRADE STANDARD 4787-2 MEDIUM GRADE STANDARD 4757-2

CDN STANDARDS SILVER, COPPER, LEAD AND ZINC STANDARDS:

CDN1 CDN2

Graphs in this appendix show up to 3 sets of lines: Red: maximum and minimum expected value Blue: the mean of the sample data plus and/or minus 1 standard deviation Yellow: the mean of the sample data plus and/or minus 2 standard deviations



LIST OF STANDARDS AND ACCEPTED VALUES

Au (ppb) Ag (g/t) Cu (%) Pb (%) Zn (%) Cozamin LOW GRADE STANDARD 4759

Lower Limit 189.5 21.68 0.52 0.00 0.63 Mean 200.3 24.42 0.55 0.01 0.65

Upper Limit 211.1 27.15 0.58 0.03 0.68 LOW GRADE STANDARD 4787-2 Lower Limit 14.96 28.88 0.62 0.11 0.62

Mean 98.50 35.55 0.66 0.12 0.68 Upper Limit 182.04 42.21 0.70 0.13 0.74 MEDIUM GRADE STANDARD 4757-2 Lower Limit

Mean Upper Limit HIGH GRADE STANDARD 4787 Lower Limit 92.9 188.13 3.29 0.15 2.65

Mean 109.4 212.46 3.45 0.17 2.78 Upper Limit 126.0 236.79 3.61 0.19 2.91

Au (g/t) Ag (g/t) Cu (%) Pb (%) Zn (%)

CDN CDN1 (CDN-HLLC)

Lower Limit 0.71 58.4 1.43 0.26 2.84

Mean 0.83 65.1 1.49 0.29 3.01

Upper Limit 0.95 71.8 1.55 0.32 3.18

CDN2 (CDN-HLHC)

Lower Limit 1.75 102.4 4.80 0.16 2.24

Mean 1.97 111.0 5.07 0.17 2.35

Upper Limit 2.19 119.6 5.34 0.18 2.46



COZAMIN BLANKS

CZ CDN BLANK - AG

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

4.50

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Ag

(g/t

CZ CDN BLANK - CU

0.00

0.02

0.04

0.06

0.08

0.10

0.12

0.14

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Cu

(%



CZ CDN BLANK - PB

0

0.05

0.1

0.15

0.2

0.25

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Pb (%

CZ CDN BLANK - ZN

0

0.02

0.04

0.06

0.08

0.1

0.12

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Zn (%



CDN LABORATORY BLANKS

CZ CDN BLANKS - AG

0

1

2

3

4

5

6

7

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Ag

(g/t

CZ CDN BLANKS - CU

0

0.02

0.04

0.06

0.08

0.1

0.12

0.14

0.16

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Cu

(%



CZ CDN BLANKS - PB

0

0.05

0.1

0.15

0.2

0.25

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Pb (%

CZ CDN BLANKS - ZN

0

0.02

0.04

0.06

0.08

0.1

0.12

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Zn (%



COMPANY STANDARD: LOW GRADE STANDARD 4759

CZ STD 4759 - AG

0.00

20.00

40.00

60.00

80.00

100.00

120.00

140.00

0 10 20 30 40 50 60 70

Number of Samples

Ag

(g/t

CZ STD 4759 - CU

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

4.50

5.00

0 10 20 30 40 50 60 70

Number of Samples

Cu

(%



CZ STD 4759 - PB

0.00

0.05

0.10

0.15

0.20

0.25

0 10 20 30 40 50 60 70

Number of Samples

Pb (%

CZ STD 4759 - ZN

0.00

0.50

1.00

1.50

2.00

2.50

0 10 20 30 40 50 60 70

Number of Samples

Zn (%



COMPANY STANDARD: LOW GRADE STANDARD 4787-2

CZ STD 4787-2 - AG

0.00

5.00

10.00

15.00

20.00

25.00

30.00

35.00

40.00

45.00

0 20 40 60 80 100 120

Number of Samples

Ag

(g/t

CZ STD 4787-2 - CU

0.00

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0.80

0.90

0 20 40 60 80 100 120

Number of Samples

Cu

(%



CZ STD 4787-2 - PB

0.00

0.05

0.10

0.15

0.20

0.25

0.30

0.35

0 20 40 60 80 100 120

Number of Samples

Pb (%

CZ STD 4787-2 - ZN

0.00

0.20

0.40

0.60

0.80

1.00

1.20

1.40

1.60

0 20 40 60 80 100 120

Number of Samples

Zn (%



COMPANY STANDARD: MEDIUM GRADE STANDARD 4757-2

CZ STD 4757-2 - AG

0.00

50.00

100.00

150.00

200.00

250.00

300.00

350.00

400.00

0 20 40 60 80 100 120 140 160 180

Number of Samples

Ag

(g/t

CZ STD 4757-2 - CU

0

1

2

3

4

5

6

0 20 40 60 80 100 120 140 160 180

Number of Samples

Cu

(%



CZ STD 4757-2 - PB

-0.50

0.00

0.50

1.00

1.50

2.00

2.50

0 20 40 60 80 100 120 140 160 180

Number of Samples

Pb (%

CZ STD 4757-2 - ZN

0

1

2

3

4

5

6

7

8

0 20 40 60 80 100 120 140 160 180

Number of Samples

Zn (%



CDN LABORATORIES STANDARD: CDN1

CZ STD CDN1 - AG

0.00

10.00

20.00

30.00

40.00

50.00

60.00

70.00

80.00

90.00

100.00

0 5 10 15 20 25 30

Number of Samples

Ag

(g/t

CZ STD CDN1 - CU

0.00

0.20

0.40

0.60

0.80

1.00

1.20

1.40

1.60

1.80

0 5 10 15 20 25 30

Number of Samples

Cu

(%



CZ STD CDN1 - PB

0.00

0.10

0.20

0.30

0.40

0.50

0.60

0.70

0 5 10 15 20 25 30

Number of Samples

Pb (%

CZ STD CDN1 - ZN

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

0 5 10 15 20 25 30

Number of Samples

Zn (%



CDN LABORATORIES STANDARD: CDN2

CZ STD CDN2 - AG

0.00

20.00

40.00

60.00

80.00

100.00

120.00

140.00

160.00

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Ag

(g/t

CZ STD CDN2 - CU

0.00

1.00

2.00

3.00

4.00

5.00

6.00

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Cu

(%



CZ STD CDN2 - PB

0.00

0.05

0.10

0.15

0.20

0.25

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Pb (%

CZ STD CDN2 - ZN

0.00

0.50

1.00

1.50

2.00

2.50

3.00

3.50

4.00

4.50

0 2 4 6 8 10 12 14 16 18 20

Number of Samples

Zn (%



COZAMIN PREPARATION CHECKS

Ag Duplicate Assays

0

100

200

300

400

500

600

700

800

900

0 100 200 300 400 500 600 700

ID

Ag

(g/t CZ

ALS

Silver Duplicate Assays

R2 = 0.799

0

100

200

300

400

500

600

700

800

0 100 200 300 400 500 600 700 800 900

Cozamin Lab Ag (g/t)

ALS

Che

mex

Ag

(g



Cu Duplicate Assays

0

2

4

6

8

10

12

14

0 100 200 300 400 500 600 700

ID

Cu

(% CZALS

Copper Duplicate AssaysR2 = 0.8672

0

2

4

6

8

10

12

0 2 4 6 8 10 12

Cozamin Lab Cu (%)

ALSC

hem

ex C

u (



Pb Duplicate Assays

0.00

1.00

2.00

3.00

4.00

5.00

6.00

7.00

0 100 200 300 400 500 600

ID

Pb

(% CZALS

Lead Duplicate Assays

R2 = 0.7292

0

5

10

15

20

25

30

35

40

0 2 4 6 8 10 12 14 16 18 20

Cozamin Lab Pb (%)

ALSC

hem

ex P

b (



Zn Duplicate Assays

0

5

10

15

20

25

30

35

0 100 200 300 400 500 600 700

ID

Zn (% CZ

ALS

Zinc Duplicate AssaysR2 = 0.691

0

5

10

15

20

25

30

35

0 5 10 15 20 25 30 35

Cozamin Lab Zn (%)

ALS

Che

mex

Zn

(



APPENDIX 3 SPECIFIC GRAVITY DETERMINATIONS PHASE IV AND V DRILL PROGRAMS

Hole Depth From Depth To SG Hole Depth From Depth To SG S38 761.50 762.00 2.62 S41 786.00 787.00 9.02S38 762.50 763.00 2.79 S41 787.00 788.00 1.64S38 763.30 763.80 2.58 S41 788.75 789.50 2.51S38 764.30 764.80 2.62 S41 790.00 790.50 2.62S38 765.30 765.60 2.57 S41 791.00 791.50 2.55S38 765.60 766.30 2.56 S41 792.00 792.50 2.59S38 768.50 768.75 0.96 S41 792.50 793.00 2.55S38 769.00 769.25 2.47 S41 794.00 794.50 2.58S38 769.50 769.75 2.57 S41 795.00 795.50 2.53S38 770.00 770.25 2.52 S41 796.00 796.50 2.53S38 770.25 771.00 2.52 S41 797.00 797.50 2.47S38 771.00 771.75 2.55 S41 798.00 798.50 2.54S38 772.00 772.25 3.01 S41 799.00 799.50 2.51S38 772.50 772.75 2.60 S41 800.00 800.55 2.49S38 773.00 773.25 2.40 S42 781.00 781.50 2.54S38 773.50 773.75 2.55 S42 783.00 783.50 2.54S39 742.00 742.50 2.52 S42 783.50 784.00 2.57S39 742.50 743.00 2.58 S42 784.00 784.50 2.56S39 743.00 743.50 2.51 S42 784.50 785.00 2.54S39 745.00 745.50 2.52 S42 785.00 785.50 2.52S39 746.00 746.50 2.58 S42 785.50 786.00 2.61S39 746.50 746.75 2.63 S42 786.00 786.50 2.54S39 747.00 747.50 2.53 S42 789.50 791.50 2.55S39 747.50 748.00 2.51 S42 792.50 793.00 3.41S39 748.50 749.00 2.44 S42 793.50 794.00 2.56S39 749.50 750.00 2.55 S39 750.50 751.00 2.47 S40 600.00 600.50 2.42 S40 656.50 657.00 2.51 S40 657.50 658.00 2.54 S40 658.00 658.50 2.53 S40 659.00 659.50 2.40 S40 661.00 661.50 2.45 S40 662.00 662.50 2.33 S40 663.00 663.50 2.55 S40 664.00 664.50 2.49 S40 665.00 665.50 2.49 S40 665.50 666.00 2.46 S40 666.50 667.00 2.53 S40 667.50 668.00 2.62 S40 668.00 668.50 2.60 S40 669.00 669.50 2.45 S40 670.00 670.50 2.48



Hole Depth From Depth To SG Hole Depth From Depth To SG U115 64.25 64.75 0.26 U119 338.60 338.85 2.75U115 142.00 142.50 2.82 U119 351.70 351.95 2.57U115 147.50 148.00 2.84 U119 352.45 352.95 2.70U115 150.50 151.00 2.58 U119 392.00 392.50 2.58U115 151.00 151.50 2.62 U119 393.00 393.50 2.66U115 152.00 152.50 3.27 U120 311.00 311.50 2.59U115 152.50 153.00 2.65 U120 312.00 312.50 2.79U116 132.40 132.60 2.48 U120 312.75 313.25 2.76U116 132.60 132.90 2.55 U120 314.75 315.25 2.75U116 132.90 133.15 2.22 U120 316.00 316.50 2.78U116 133.15 133.40 2.32 U120 317.00 317.50 2.89U116 133.40 133.65 2.31 U120 317.75 318.00 2.70U116 133.65 133.90 1.30 U120 318.00 318.50 3.22U116 133.90 134.40 2.51 U120 319.00 319.50 2.63U116 134.40 134.90 2.51 U120 320.00 320.50 2.72U116 134.90 135.40 2.56 U120 320.75 321.25 2.96U117 56.50 56.75 2.51 U120 321.75 322.00 2.39U117 56.75 57.00 3.03 U120 322.50 323.00 2.66U117 57.00 57.75 2.58 U120 323.50 324.00 2.58U117 57.75 58.25 3.16 U120 324.50 325.00 2.67U117 58.25 58.75 3.31 U120 325.50 326.00 2.86U117 58.75 59.00 2.59 U120 326.50 327.00 3.01U117 153.15 154.15 2.50 U120 327.50 328.00 2.95U117 154.15 154.65 3.05 U120 328.00 329.00 2.68U117 154.65 155.15 2.78 U121 282.20 282.50 2.53U117 168.00 168.50 2.51 U121 283.00 283.50 3.51U117 168.50 169.00 2.56 U121 284.00 284.50 3.58U117 169.00 169.75 2.46 U121 285.00 285.50 3.37U117 169.75 170.00 2.53 U121 317.00 317.50 2.57U117 170.00 170.50 2.52 U121 318.00 318.25 2.94U117 170.50 171.00 2.48 U121 318.50 319.00 2.95U117 171.00 171.50 2.46 U121 319.00 319.50 3.15U117 171.50 172.00 2.49 U121 320.00 320.25 2.75U118 396.50 397.25 2.48 U121 320.50 321.00 3.39U118 397.85 398.00 2.50 U121 321.25 321.50 3.30U118 398.00 398.45 2.50 U121 321.50 322.00 3.48U118 398.45 399.45 2.59 U121 322.50 323.00 3.09U118 399.45 399.95 2.64 U118 399.95 400.85 2.66 U118 400.85 401.35 27.10 U118 411.00 412.00 2.56 U118 412.00 412.50 2.59 U118 412.50 413.10 2.68 U118 413.10 413.40 2.82 U118 414.10 414.40 2.76 U118 415.40 416.30 2.83 U118 416.30 417.30 2.64



Hole Depth From Depth To SG Hole Depth From Depth To SG U122 374.50 375.00 2.72 U126 186.70 187.20 2.57U122 375.00 375.50 3.04 U126 187.70 188.00 3.42U122 375.75 376.00 2.58 U126 188.00 188.25 2.64U122 376.25 376.50 2.86 U126 196.00 196.50 2.78U122 409.00 409.50 2.62 U126 197.00 198.00 2.67U122 410.00 410.50 2.66 U126 199.00 199.50 2.67U122 411.00 411.65 2.59 U126 200.10 200.50 2.67U122 412.05 412.30 2.82 U126 200.75 201.00 2.49U122 412.75 413.25 3.27 U126 201.25 202.75 3.11U122 413.50 414.00 2.82 U126 203.00 203.50 2.79U122 414.25 414.50 2.79 U126 203.80 204.50 2.68U122 414.50 415.00 3.04 U126 204.80 205.50 2.60U122 415.50 415.75 2.82 U127 249.50 250.00 2.94U122 416.25 416.50 2.70 U127 250.00 250.50 3.41U122 417.00 417.50 2.65 U127 251.00 251.50 3.03U122 418.00 418.50 2.73 U127 252.00 252.50 3.14U122 419.00 420.00 4.95 U127 253.00 253.50 2.80U123 390.15 390.50 2.64 U128 170.00 170.50 2.61U123 390.50 390.75 3.05 U128 171.00 171.50 2.70U123 391.00 391.50 3.10 U128 171.50 172.00 2.69U123 392.00 392.50 3.05 U128 172.50 173.00 2.79U123 393.00 393.50 2.60 U129 285.50 286.00 2.53U123 394.00 394.25 2.63 U129 286.50 286.75 2.67U124 301.50 302.00 2.96 U129 286.75 287.25 2.83U124 302.50 303.00 2.68 U129 287.50 288.00 2.85U124 303.50 304.20 2.67 U129 288.50 289.00 2.68U124 304.50 305.00 3.16 U129 289.50 290.00 2.77U124 305.50 305.75 2.86 U129 290.50 291.00 3.02U124 306.00 306.50 3.03 U129 291.00 291.50 3.03U124 307.00 307.25 3.41 U130 330.00 332.75 2.67U124 307.50 308.00 2.97 U130 333.50 334.00 3.19U124 308.50 308.75 2.93 U130 334.50 335.00 2.74U124 308.75 309.25 2.96 U130 335.50 336.00 2.60U124 309.50 310.00 2.58 U130 336.50 337.25 2.55U124 310.50 311.25 2.65 U130 337.75 338.00 2.76U124 311.75 312.00 2.51 U131 188.20 188.50 2.70U125 288.75 289.25 2.78 U131 189.00 189.50 0.78U125 289.50 290.00 3.03 U131 190.00 190.50 2.85U125 290.50 291.00 2.83 U131 191.00 191.50 2.62U125 291.00 291.50 2.77 U131 191.75 192.00 2.64U125 291.75 292.25 3.12 U131 193.00 193.20 2.51U125 292.50 293.00 3.06 U131 193.50 194.00 2.70U125 293.00 293.25 2.78 U131 194.50 195.00 2.67 U131 195.50 196.00 2.52 U131 196.50 197.00 2.65 U131 197.50 198.00 2.56 U131 198.50 199.00 2.63



Hole Depth From Depth To SG Hole Depth From Depth To SG U132 327.25 327.75 2.60 U138 342.50 343.00 2.50U132 328.00 328.50 2.58 U138 343.50 344.00 2.47U132 329.00 329.50 2.64 U138 344.50 345.00 2.54U132 330.00 330.50 2.58 U138 345.50 346.00 2.52U132 331.00 331.50 2.46 U138 346.50 347.00 2.60U132 332.00 332.50 2.57 U138 347.50 348.00 2.59U132 333.00 333.50 2.79 U138 348.00 348.50 2.46U132 333.50 334.00 3.12 U138 349.00 349.50 2.58U132 334.50 335.00 2.87 U138 350.00 350.50 2.61U132 335.00 336.00 2.64 U139 219.50 220.00 2.95U132 337.00 337.50 2.58 U139 220.00 220.50 3.11U133 319.55 320.00 2.66 U139 221.00 221.50 2.93U133 320.50 321.00 2.59 U139 222.00 222.25 3.04U133 321.50 322.00 2.53 U139 222.50 223.00 2.63U133 322.50 323.00 2.59 U140 363.25 363.50 2.57U134 231.00 231.50 2.48 U140 364.00 364.50 2.47U134 232.00 232.50 2.62 U140 365.00 365.50 2.46U134 232.75 233.00 3.74 U140 365.50 366.00 2.38U134 234.00 234.50 2.64 U140 366.50 367.00 2.37U134 235.00 235.50 2.57 U140 367.50 368.00 2.50U134 236.00 236.50 2.66 U141 400.00 400.75 2.59U134 237.00 237.50 5.21 U141 401.00 401.50 2.50U134 238.00 238.50 2.55 U141 402.00 402.50 2.97U134 239.00 239.50 2.56 U141 402.50 403.00 2.69U134 240.00 240.50 2.58 U141 403.50 404.00 2.52U134 241.00 241.50 2.56 U141 404.50 405.00 2.58U134 242.00 243.00 2.65 U141 405.25 405.50 2.56U134 244.00 245.00 2.93 U142 214.00 214.50 2.73U134 245.50 246.00 2.58 U142 215.00 215.75 2.54U134 246.50 247.00 2.83 U142 216.50 217.00 2.54U134 247.50 248.00 2.72 U142 217.50 218.00 2.55U134 248.50 249.00 2.83 U142 218.65 219.00 2.52U134 249.50 250.00 2.98 U143 338.00 338.50 2.47U134 250.50 251.00 2.91 U143 338.50 339.00 2.52U134 251.50 252.00 2.68 U143 339.50 340.00 2.78U134 252.50 253.00 2.69 U143 340.25 340.50 2.52U134 253.50 254.10 2.62 U143 341.00 341.50 2.64U135 342.40 343.00 2.64 U143 341.50 342.00 2.52U135 343.00 343.50 2.52 U143 342.50 343.00 2.66U135 344.00 344.50 2.54 U143 343.50 344.00 2.61U136 346.40 347.00 2.90 U143 344.50 345.00 2.57U136 347.00 347.50 2.71 U136 347.70 348.00 2.56 U136 348.50 348.90 2.59 U137 263.50 263.70 2.44 U137 264.00 264.50 2.51 U137 264.50 265.00 2.53 U137 265.50 266.20 2.53



Hole Depth From Depth To SG Hole Depth From Depth To SG

U144 242.50 243.00 2.72 U149 195.50 196.00 2.89U144 243.25 243.50 3.30 U149 198.00 198.50 2.65U144 243.50 244.00 3.37 U149 199.00 199.50 3.82U144 244.50 245.00 3.06 U149 201.00 201.50 2.75U144 245.50 246.00 2.97 U149 204.00 204.50 2.64U144 246.50 247.00 2.93 U149 204.50 205.00 3.80U144 247.50 248.00 2.48 U149 205.00 205.50 4.70U144 248.50 249.00 2.55 U149 205.50 205.75 4.31U144 249.50 250.00 3.03 U149 245.00 245.50 2.66U144 250.50 251.00 2.45 U149 246.00 246.50 2.69U145 257.70 258.00 2.59 U149 247.00 247.50 2.66U145 258.00 258.50 3.14 U149 247.50 248.00 2.72U145 258.50 259.00 3.00 U149 248.50 249.00 2.93U145 259.00 259.50 2.84 U149 249.00 249.55 2.84U145 259.75 260.00 2.58 U149 250.00 250.50 3.57U146 179.00 179.50 2.50 U150 335.25 336.00 2.57U146 179.50 180.00 2.42 U150 337.00 338.00 2.52U146 180.50 181.00 2.62 U150 338.25 338.50 2.41U146 191.50 192.25 2.92 U150 339.00 339.50 2.54U146 192.50 193.00 2.62 U150 339.50 340.00 2.60U146 193.00 193.50 2.65 U150 340.50 341.00 2.48U146 194.00 194.50 2.60 U150 341.50 341.75 2.43U147 351.50 352.50 2.59 U150 341.75 342.25 2.51U147 352.50 353.00 2.60 U150 342.75 343.00 2.46U147 353.00 354.00 2.54 U147 354.75 355.00 2.49 U147 355.00 355.25 2.69 U147 355.50 356.00 2.95 U147 356.00 356.50 3.37 U147 357.00 357.50 3.04 U147 357.50 358.00 2.74 U147 358.50 359.00 2.63 U148 389.00 390.00 3.15 U148 390.25 390.50 2.44 U148 395.50 396.00 2.57 U148 396.50 397.00 2.56 U148 397.00 397.50 2.57 U148 398.00 398.50 2.72 U148 399.00 399.50 2.65 U148 400.00 400.50 2.68 U148 401.30 402.00 2.58 U148 402.50 403.00 2.72




U151 193.50 194.00 2.80 U152 348.50 349.00 2.95U151 195.00 196.00 2.64 U152 349.50 350.00 2.47U151 196.50 197.00 3.04 U152 350.75 351.00 2.51U151 197.00 197.50 2.95 U152 351.25 351.75 2.71U151 198.00 198.50 3.40 U152 351.75 352.25 2.85U151 199.00 199.50 2.50 U152 352.50 353.00 3.00U151 200.00 200.50 2.63 U152 353.50 354.00 3.09U151 201.00 202.00 2.58 U152 354.50 355.00 2.81U151 220.00 221.00 2.49 U152 355.50 356.00 2.89U151 222.00 223.00 2.26 U152 356.50 357.00 2.64U151 223.50 224.00 2.60 U152 357.50 358.00 2.74U151 225.70 226.00 3.11 U152 358.50 359.25 2.67U151 227.35 228.00 3.66 U152 359.75 360.00 2.82U151 228.50 229.00 3.50 U152 360.50 361.00 3.15U151 232.00 232.50 2.99 U152 361.00 361.50 2.62U151 232.95 233.50 3.53 U152 362.00 362.50 2.83U151 233.50 234.00 3.63 U152 363.00 363.50 2.46U151 234.00 234.50 3.25 U152 363.50 365.00 2.51U151 235.00 235.50 3.22 U151 235.50 236.00 3.12 U151 236.50 237.00 3.19 U151 237.50 238.00 2.87 U151 238.30 238.75 2.63 U151 240.00 241.00 2.64 U151 243.50 244.00 2.65 U151 244.50 245.10 2.68 U151 245.50 246.00 3.27 U151 246.30 247.00 3.45 U151 247.50 248.00 3.34 U151 248.50 249.00 3.11 U151 249.50 250.00 3.49 U151 250.50 251.00 3.61 U151 251.50 252.00 3.19 U151 252.40 253.00 2.72




U153 271.50 272.00 2.63 U156 145.80 146.00 2.45U153 273.75 274.00 2.75 U156 146.50 147.00 2.87U153 275.25 275.75 2.65 U156 147.50 148.00 2.75U153 277.80 278.35 2.83 U156 148.00 148.50 3.02U153 278.35 279.00 2.74 U156 149.00 149.50 3.21U153 279.50 280.00 2.68 U156 149.50 150.20 3.19U153 280.00 280.60 3.17 U157 364.00 364.50 2.58U153 281.00 282.00 2.49 U157 365.00 365.50 2.51U153 283.00 283.75 2.63 U157 365.50 366.00 2.72U153 286.00 286.50 2.74 U157 366.50 367.00 2.69U153 289.00 289.50 2.63 U157 367.50 368.25 2.51U153 290.25 290.50 3.25 U157 371.30 374.35 2.69U153 290.50 291.00 3.36 U158 170.00 170.25 2.70U153 291.50 292.00 2.95 U158 170.50 171.00 3.35U153 292.50 292.80 2.77 U158 171.50 172.00 2.79U153 292.80 293.25 2.69 U158 172.50 172.75 2.51U153 294.25 294.75 2.71 U158 173.00 173.50 3.10U153 295.50 296.00 2.64 U158 173.50 174.00 2.88U153 298.00 298.50 2.55 U158 174.50 175.00 2.55U153 299.50 300.00 2.57 U159 201.50 202.00 3.29U153 300.60 301.00 2.86 U159 202.00 202.50 2.72U153 301.50 302.00 2.98 U159 203.00 203.50 2.64U153 302.00 302.50 3.09 U159 204.00 204.50 2.55U153 302.50 303.00 2.83 U159 205.00 205.50 2.54U153 303.50 304.00 3.16 U159 206.00 206.50 2.52U153 304.50 304.75 3.16 U159 207.00 207.50 3.53U153 304.75 305.00 2.29 U159 208.30 209.00 2.57U153 305.00 305.50 3.03 U160 363.10 363.60 2.61U153 306.00 306.50 3.27 U160 364.10 364.50 2.48U153 306.50 307.00 3.15 U160 364.50 364.75 2.96U153 308.25 308.50 2.96 U160 365.00 365.30 2.54U153 308.50 309.00 3.13 U153 310.00 310.50 3.35 U153 311.00 311.50 2.97 U153 312.00 312.50 3.03 U153 312.50 313.00 3.78 U153 313.00 313.50 3.36 U153 314.00 314.60 3.21 U154 159.85 160.50 2.62 U154 161.00 161.50 2.84 U154 161.75 162.50 2.54 U155 134.50 135.00 2.92 U155 135.00 135.50 2.93 U155 136.00 136.50 2.76 U155 137.00 137.50 3.00 U155 137.50 138.00 3.20




U161 172.50 173.00 2.34 U165 117.00 117.50 2.57U161 173.35 173.75 3.25 U165 117.50 118.00 2.84U161 174.25 174.50 3.19 U165 118.50 119.00 2.70U161 175.00 175.50 3.71 U165 120.50 121.00 2.58U161 175.50 176.00 3.85 U165 121.50 122.00 2.55U161 176.50 177.00 2.96 U165 122.50 123.50 2.54U161 177.50 178.00 2.62 U165 124.50 125.00 2.46U161 178.00 178.20 2.82 U165 125.00 125.50 2.57U161 178.80 179.50 2.60 U165 125.50 126.00 2.24U162 209.65 210.00 2.83 U165 254.00 255.00 2.65U162 210.05 211.00 2.88 U165 255.50 256.00 2.54U162 211.50 212.00 3.34 U165 257.50 258.00 2.58U162 212.50 213.00 3.35 U165 258.50 259.00 2.60U162 213.50 214.00 3.04 U165 260.00 260.50 2.79U162 214.50 215.00 3.00 U165 260.50 261.00 2.63U162 215.50 216.00 2.63 U165 261.30 261.50 2.66U162 217.00 218.00 2.57 U165 262.00 263.00 2.67U162 219.50 220.00 2.75 U166 171.50 172.00 2.68U162 220.50 221.00 2.87 U166 172.00 172.50 2.59U163 349.00 349.85 2.70 U166 172.75 173.00 2.62U163 350.10 350.50 1.60 U166 173.50 174.00 2.74U163 351.00 352.00 2.57 U166 174.50 175.00 2.91U163 354.50 355.00 4.93 U166 175.50 176.00 2.19U163 358.00 359.00 2.79 U166 176.00 176.50 3.46U163 360.50 360.60 2.83 U166 177.00 177.25 3.01U163 361.00 361.50 2.67 U167 180.50 181.00 2.71U163 362.00 363.00 2.76 U167 182.00 183.00 2.61U163 364.00 365.00 2.65 U167 186.00 186.50 2.77U163 367.75 368.00 3.20 U167 187.00 187.50 2.90U163 368.59 369.00 2.86 U167 188.00 188.50 2.58U163 370.50 371.00 2.60 U167 189.85 190.25 3.36U163 372.00 372.50 2.68 U167 191.00 191.50 3.02U164 370.00 370.50 2.43 U167 191.50 192.00 3.18U164 371.00 372.00 1.89 U167 192.00 192.50 3.12U164 372.50 373.00 2.51 U167 193.00 193.50 3.13U164 373.50 374.00 2.54 U167 194.00 194.50 3.23U164 374.00 375.00 2.61 U167 194.50 195.00 3.21U164 375.50 376.00 2.48 U167 195.50 196.00 3.43U164 376.50 377.00 2.55 U167 197.00 197.50 3.13U164 377.25 377.50 2.44 U167 198.00 198.50 2.54U164 378.00 378.50 2.70 U167 199.00 199.50 4.22U164 379.00 379.50 2.66 U167 200.50 201.00 3.20U164 380.00 380.50 2.34 U167 201.50 202.00 3.41U164 380.50 381.00 2.49 U167 202.50 202.75 3.10U164 381.50 382.00 2.74 U167 202.75 203.00 3.04 U167 203.00 203.50 3.24



Hole Depth From Depth To SG Hole Depth From Depth To SG U168 285.65 286.00 2.63 U173 219.00 219.70 2.74U168 286.50 287.00 2.32 U173 219.70 220.25 2.84U168 287.00 287.50 2.57 U173 220.75 221.50 2.61U168 288.00 288.50 2.61 U173 222.20 222.50 3.03U168 289.00 289.25 2.57 U173 223.25 223.50 2.93U169 267.30 268.00 2.61 U173 224.00 224.50 2.78U169 268.50 268.75 2.63 U173 224.50 224.75 2.54U169 269.00 269.50 2.72 U173 225.25 226.50 2.56U169 269.50 270.00 2.98 U174 265.10 265.50 2.62U169 270.50 271.00 2.51 U174 266.00 266.50 3.07U170 185.00 186.00 2.82 U174 266.50 267.00 2.83U170 187.00 187.50 2.66 U175 321.15 321.50 2.66U170 188.50 189.00 2.61 U175 322.00 322.50 2.68U170 189.00 189.50 2.75 U175 322.50 323.00 2.63U170 190.50 191.00 2.66 U175 323.50 323.70 2.61U170 191.00 191.50 2.72 U175 324.00 324.50 2.75U170 192.15 192.50 2.83 U175 325.00 326.00 2.57U170 194.00 194.50 3.24 U176 202.85 203.25 3.10U170 195.00 195.50 3.05 U176 203.50 204.00 2.97U170 196.00 196.50 3.51 U176 204.25 204.75 2.74U170 196.50 197.00 3.74 U176 205.00 205.50 2.71U170 198.00 198.50 2.66 U176 206.00 206.50 2.68U170 199.50 200.00 2.40 U176 206.50 207.00 2.76U171 303.00 303.50 2.93 U176 207.50 208.00 2.62U171 304.00 304.50 3.10 U176 208.50 209.00 2.82U171 304.50 305.00 2.78 U176 209.50 210.00 2.70U171 305.50 306.00 2.59 U176 211.00 211.65 2.53U171 306.25 307.00 2.59 U176 212.00 212.25 2.53U171 307.50 308.00 2.64 U176 218.50 219.00 2.70U171 308.00 308.50 2.94 U176 219.50 220.00 2.72U172 344.50 345.00 2.64 U177 241.00 241.50 3.67U172 347.50 348.00 2.63 U177 242.00 242.35 3.66U172 348.50 349.00 2.60 U177 290.70 291.00 2.54U172 351.00 351.50 2.55 U177 291.50 292.00 2.60U172 352.00 352.50 2.53 U177 292.50 292.75 2.96U172 353.00 353.50 2.56 U178 195.35 196.00 3.03U172 354.00 354.50 2.48 U178 196.50 197.00 3.04U172 355.00 355.50 2.48 U178 197.00 198.00 2.91U172 356.00 356.50 2.60 U178 198.00 198.50 3.34U172 357.00 357.50 2.66 U178 199.00 199.50 3.35U172 358.00 358.50 2.54 U178 200.00 200.25 3.48U172 359.00 359.50 2.56 U172 360.00 360.50 2.54 U172 361.00 361.50 2.49 U172 362.00 362.50 2.48 U172 363.00 363.50 2.47 U172 364.50 365.50 2.82



Hole Depth From

Depth To SG Hole

Depth From

Depth To SG

U179 172.45 173.00 2.90 U183 347.00 347.50 2.65 U179 173.50 174.00 3.43 U183 348.00 348.50 2.58 U179 174.00 174.25 2.65 U183 348.95 349.25 2.72 U179 174.75 175.00 2.43 U183 349.75 350.25 2.68 U179 175.50 176.00 2.58 U183 350.25 350.50 2.70 U179 176.50 177.00 3.22 U183 351.00 351.50 2.71 U179 177.50 178.00 2.89 U183 352.00 352.50 2.83 U179 190.65 191.00 3.08 U183 353.00 353.50 2.77 U179 191.50 192.00 3.25 U183 354.00 354.25 3.12 U179 192.50 193.00 2.99 U183 354.50 355.00 3.00 U179 193.50 194.00 2.66 U183 355.00 355.35 3.09 U179 194.50 195.15 2.91 U183 356.00 357.00 2.52 U179 202.75 203.00 2.56 U179 203.50 204.00 2.81 U179 204.50 205.00 2.80 U179 205.00 205.50 2.54 U180 218.50 219.50 2.61 U180 220.00 220.50 2.63 U180 221.00 221.25 2.72 U180 221.50 222.00 2.84 U180 222.00 222.50 3.35 U180 223.50 224.00 2.61 U180 224.35 225.00 3.07 U180 225.50 225.75 2.80 U180 226.25 226.70 2.59 U181 301.40 302.00 2.71 U181 302.50 303.00 2.69 U181 303.35 304.00 2.57 U181 304.50 305.00 2.59 U182 172.50 173.00 2.76 U182 174.00 174.50 2.70 U182 176.00 176.50 2.72 U182 176.85 177.50 3.03 U182 178.00 178.50 3.40 U182 178.50 179.00 3.50 U182 179.00 179.25 2.60 U182 179.70 180.00 1.67 U182 182.00 182.50 2.55

TECHNICAL REPORT ON THE COZAMIN PROJECT, ZACATECAS … · 23.11.2007 · Technical Report on the...

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