Exhibit 99.1

EX-99.1 2 exhibit99-1.htm NI 43-1010 TECHNICAL REPORT DATED FEBRUARY 5, 2008 Exhibit 99.1

NI 43-101 Technical Report
International Royalty Corporation
Las Cruces Project Royalty
Sevilla Province, Spain

Prepared for:

International Royalty Corporation
10 Inverness Drive East
Suite 104
Englewood, CO 80112

SRK Project Number: 151911

Prepared by:

7175 W. Jefferson Ave.
Suite 3000
Lakewood, CO 80235

Effective Date: February 5, 2008
Report Date: February 5, 2008

Contributor:	Endorsed by QP:
Katherine L. Garramone	Dr. Neal Rigby, CEng, MIMMM, PhD

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Table of Contents

1	INTRODUCTION (ITEM 4)			1-1
	1.1	Terms of Reference and Purpose of the Report		1-1
	1.2	Reliance on Other Experts (Item 5)		1-1
		1.2.1	Limitations and Reliance on Information	1-1
	1.3	Effective Data (Item 24)		1-2
	1.4	Qualifications of Consultants (SRK)		1-2
2	PROPERTY DESCRIPTION AND LOCATION (ITEM 6)			2-1
	2.1	Property Description		2-1
	2.2	Property Location		2-1
	2.3	Project Ownership		2-1
	2.4	Mineral Tenure		2-2
	2.5	Surface Land Ownership		2-2
	2.6	Royalties, Agreements and Encumbrances		2-2
	2.7	Environmental Liabilities and Permitting		2-2
3	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY (ITEM 7)
	3.1	Topography, Elevation and Vegetation		3-1
	3.2	Climate and Length of Operating Season		3-1
	3.3	Access to Property		3-1
	3.4	Surface Rights		3-1
	3.5	Local Resources and Infrastructure		3-2
		3.5.1	Power Supply	3-2
		3.5.2	Water Supply	3-2
		3.5.3	Manpower	3-2
4	HISTORY (ITEM 8)			4-1
5	GEOLOGICAL SETTING (ITEM 9)			5-1
	5.1	Regional and Local Geology		5-1
	5.2	Project Geology		5-1
6	DEPOSIT TYPE (ITEM 10)			6-1
7	MINERALIZATION (ITEM 11)			7-1
	7.1	Secondary Sulfide Mineralization		7-1
	7.2	Primary Sulfide Mineralization		7-2
8	EXPLORATION (ITEM 12)			8-1
	8.1	Deposit Exploration		8-1
9	DRILLING (ITEM 13)			9-1
	9.1	Type and Extent of Drilling		9-1
		9.1.1	Core Drilling	9-1
		9.1.2	Reverse Circulation Drilling	9-2
		9.1.3	Interpretation	9-2
10	SAMPLING METHOD AND APPROACH (ITEM 14)			10-1
	10.1	Sampling Methods		10-1
		10.1.1	Core Sampling	10-1


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		10.1.2	Core Recovery	10-2
11	SAMPLE PREPARATION, ANALYSES AND SECURITY (ITEM 15)			11-1
		11.1.1	Reverse Circulation Drilling	11-1
	11.2	Sample Preparation and Assaying Methods		11-1
		11.2.1	Density Testwork	11-2
	11.3	Quality Controls and Quality Assurance		11-3
12	DATA VERIFICATION (ITEM 16)			12-1
	12.1	Assay Checks		12-1
	12.2	PQ Twin Drilling		12-2
	12.3	Other Verification		12-2
	12.4	Drillhole Database Verification		12-2
13	ADJACENT PROPERTIES (ITEM 17)			13-1
14	MINERAL PROCESSING AND METALLURGICAL TESTING (ITEM 18)			14-1
	14.1	Metallurgical Testing Programs		14-1
	14.2	Metallurgical Samples		14-1
	14.3	Batch Metallurgical Testing		14-2
	14.4	Pilot Leaching and Solvent Extraction Testing		14-3
	14.5	Pilot Effluent Testing		14-4
	14.6	Ore Variability		14-4
	14.7	OTG Additional Testing		14-4
	14.8	Copper Recovery Predictions		14-6
	14.9	Copper Production Estimate		14-7
	14.10	PAH Comments		14-7
15	MINERAL RESOURCES AND ORE RESERVE ESTIMATES (ITEM 19)			15-1
		15.1.1	Drillhole Database	15-1
		15.1.2	Compositing	15-5
		15.1.3	Rock Model	15-7
		15.1.4	Grade Models	15-7
		15.1.5	Resource Statement	15-9
		15.1.6	Resource Classification	15-9
		15.1.7	Previous Resource Estimates	15-11
	15.2	Additional 2004/2005 Drilling Results		15-11
	15.3	Additional Exploration Potential		15-12
	15.4	Reserve Development		15-13
		15.4.1	Mine Design	15-13
		15.4.2	Cut-off Grade	15-14
		15.4.3	Pit Design	15-15
		15.4.4	Mining Dilution and Losses	15-17
	15.5	Ore Reserves Statement		15-18
		15.5.1	Effects on Reserves by Other Factors	15-19
16	OTHER RELEVANT DATA AND INFORMATION (ITEM 20)			16-1
17	ADDITIONAL REQUIREMENTS FOR DEVELOPMENT PROPERTIES AND PRODUCTION (ITEM 25)
	17.1	Mining Operations and Method		17-2
	17.2	Ore Processing		17-3


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	17.3	Production Schedule		17-5
		17.3.1	Recoverability	17-6
	17.4	Markets		17-6
	17.5	Contracts		17-8
	17.6	Environmental Considerations		17-8
		17.6.1	Permitting	17-8
		17.6.2	Environmental Restoration	17-11
		17.6.3	Groundwater Management	17-13
		17.6.4	Surface Water Management	17-14
	17.7	Economic Analysis		17-14
18	INTERPRETATION AND CONCLUSIONS (ITEM 21)			18-1
	18.1	Exploration Conclusions		18-1
	18.2	Other Relevant Information		18-1
		18.2.1	Geologic Evaluation Conclusions	18-1
		18.2.2	Resource Estimation Conclusions	18-1
		18.2.3	Reserve Estimation Conclusions	18-2
		18.2.4	Economic Analysis	18-2
19	RECOMMENDATIONS (ITEM 22)			19-1
20	REFERENCES (ITEM 23)			20-1
21	GLOSSARY			21-1
	21.1	Mineral Resources and Reserves		21-1
		21.1.1	Mineral Resources	21-1
		21.1.2	Mineral Reserves	21-1
	21.2	Glossary		21-2

List of Tables
Table	10.1.1.1:	Las Cruces Core Sample Size Distribution	10-1
Table	14.9.1:	Las Cruces Copper Production Schedule	14-7
Table	15.1.1.1:	Las Cruces Sample Copper Data Statistics*	15-4
Table	15.1.1.2:	Las Cruces Sample Density Data Statistics*	15-5
Table	15.1.2.1:	Las Cruces Composite Copper Data Statistics*	15-6
Table	15.1.2.2:	Las Cruces Composite Density Data Statistics*	15-6
Table	15.1.4.1:	Las Cruces Block Model Copper Data Statistics*	15-8
Table	15.1.4.2:	Las Cruces Block Model Density Data Statistics*	15-8
Table	15.1.5.1:	Las Cruces Mineral Resource Summary*	15-9
Table	15.1.6.1:	Las Cruces Resource Confidence Classification Criteria	15-10
Table	15.1.7.1:	Las Cruces Comparison of Historical Resource Estimates*	15-11
Table	15.4.2.1:	Las Cruces Basic Cut-off Calculation	15-15
Table	15.4.3.1:	Las Cruces Design Basis – Mine Operating Cost Assumptions*	15-16


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Table	15.4.3.2:	Las Cruces Pit Design Parameters	15-17
Table	15.5.1:	Las Cruces Diluted Mineral Reserve* (PAH Report, May 2005)	15-18
Table	15.5.2:	Las Cruces Mineral Reserves (as at December 31, 2006)*	15-19
Table	17.3.1:	Las Cruces Mine and Mill Production Schedule	17-6
Table	17.4.1:	Historic Refined Copper* (t000s)	17-7
Table	17.4.2:	Historic Refined Copper* (t000s)	17-7
Table	17.6.1.1:	Las Cruces Summary of Permit Status	17-10
Table	17.6.2.1:	Las Cruces Annual Restoration Areas and Budgets	17-13
Table	17.7.1:	Royalty Revenue versus Copper Price (1.5% Royalty Rate)	17-15
Table	21.2.1:	Glossary	21-2
Table	21.2.2:	Abbreviations	21-3

List of Figures
Figure	2-1:	Las Cruces General Location Map	2-4
Figure	2-2:	Mining Concession Limits	2-5
Figure	5-1:	General Geologic Section	5-3
Figure	5-2:	General Geologic Plan of Deposit Lenses	5-4
Figure	9-1:	Drillhole Locations with Topography	9-3
Figure	14-1:	Metallurgical Sampling Composites Planview	14-9
Figure	14-2:	Metallurgical Sampling Composites Section	14-10
Figure	14-3:	Copper Extraction vs. Copper Grade	14-11
Figure	15-1:	Model Boundary with Drillholes & Ultimate Pit Outline	15-20
Figure	15-2:	Number of Samples & % Cu above Core Recovery Cut-off*	15-21
Figure	15-3:	Ultimate Pit Design	15-22
Figure	17-1:	Mine Development Plan, Mine Status Year +10 Main Infrastructure Items	17-16
Figure	17-2:	Las Cruces Process Flow Diagram	17-17

List of Appendices

Appendix A


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Certificate of Author

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Summary (Item 3)

SRK Consulting (US), Inc. (SRK), was commissioned by International Royalty Corporation (IRC) in December 2007 to prepare a Canadian Securities Administrators (CSA) National Instrument 43-101 (NI 43-101) compliant Technical Report on its 1.5% Sales Revenue (SR) royalty on all sales from the Las Cruces Project (Las Cruces or the Project) owned by Inmet Mining Corporation (Inmet) and MK Resources in the Sevilla Province of Spain.

This report has been prepared for IRC in support of a royalty interest (not direct ownership) on the Project. This report has been prepared in accordance with the guidelines provided in NI 43-101 Standards of Disclosure for Mineral Projects.

However, this report has been prepared for a company, which holds a royalty interest (not direct ownership) on the property. Mining companies are not (typically) required and, as a matter of practice, do not normally disclose detailed information to companies which hold a royalty interest in their operations unless legally mandated. IRC therefore, is limited in the amount of information and details it can disclose to that which is available in the public domain. This technical report, therefore, relies primarily upon the NI 43-101 Technical Report prepared by Pincock Allen & Holt (PAH) effective as of May 27, 2005 the (PAH Report), as well as general information available in the public domain.

The information contained in this report is effective as of February 5, 2008.

Property Description and Location

The Las Cruces Project is currently under development and copper production is scheduled to begin in 2008. The development of the Project was planned as an open pit mine, with copper cathode to be produced on site by applying Outokumpu atmospheric leaching and conventional solvent extraction electrowinning (SX/EW) technology, and now includes some underground mining to access additional tonnes. Las Cruces is expected to produce approximately 72,000t of cathode copper per year over a mine life of 15 years. (PAH Report, May 2005 and Inmet AIF, March 2007)

Ownership

The following sections are excerpted from the PAH Report (May 27, 2005). Changes to table, figure numbers, section numbers and standardizations have been made to suit the format of this report. Updates to the text made to reflect current tense, data and/or information are annotated by the use of brackets .

Under the terms of the agreement, Inmet [has] the authority to oversee the development and operation of the Las Cruces Project, subject to certain protective minority rights of Leucadia. Inmet and Leucadia [committed] to the project development costs and provide guarantees of project financing, in proportion to their respective interests in the project. Inmet [assumed] its share of development expenditures for the Las Cruces Project as of April 1, 2005.


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Royalty Interest

IRC acquired a 1.5% SR royalty interest from Rio Tinto PLC on December 21, 2007, as presented in IRC’s December 21, 2007 news release. The royalty was established by Rio Tinto PLC in 1999 when it sold its interest in CLC to MK Resources.

The description of the royalty is as follows:

A 1.5% SR royalty is to be paid to IRC by CLC in respect of all sales realized in any calendar month in which the official monthly average LME Cash Settlement price for Grade A copper is equivalent to or greater than US$0.80/lb-Cu; and
Allowable deductions in calculating sales revenue include 1) freight, transportation, and insurance to the point of sale; 2) government royalties; and 3) sales expenses.

Geology and Mineralization

The Las Cruces deposit occurs near the eastern end of the Iberian Pyrite Belt, a 250km long and 40km wide geologic belt that extends eastward from southern Portugal into southern Spain. The belt is host to more than 100 mineral deposits, some of which were exploited for metals as long ago as pre-Roman times. Mineralization at Las Cruces, as in most other Iberian Pyrite Belt deposits, consists of syngenetic massive sulfides containing polymetallic mineralization. The massive sulfide is hosted by late Devonian to early Carboniferous Period volcanic and sedimentary rocks, deposited in a submarine setting within a narrow and relatively shallow intra-continental sea, characterized by bimodal volcanism and sedimentation.

The Las Cruces massive sulfide is located within a host sequence of volcanic and sedimentary rocks, with a shale unit generally present as a 10 to 20m wide envelope around the massive sulfides. Mineralization occurs in one massive sulfide horizon that has a general strike to the east and a dip to the north at about a 35º angle, with a gradual change to the west to a northwesterly dip 30º. The overall dimensions of the massive sulfide deposit are 1,000m along strike, 500m or more down dip (still not completely drill delineated), and up to 100m thick (averages 30 to 40m). To the west, the massive sulfide mineralization thickens and is then truncated by a roughly north-south trending fault; with attempts to locate the offset mineralization not having been successful so far. Along strike to the northeast, the deposit is not completely drill delineated, however the grade of the mineralization is relatively low. Due to supergene enrichment processes in pre-Miocene time, secondary sulfide mineralization formed a generally horizontal zone in the up-dip part of the primary massive sulfide.


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Exploration

The Las Cruces deposit was discovered in 1994 by Riomin Exploraciones, S.A. (Riomin), a wholly owned subsidiary of Rio Tinto. The discovery was the result of drilling on a gravimetric survey anomaly. Rio Tinto drilled the deposit between 1994 and 1999. A total of 277 core holes, for 82,352m, were drilled by Rio Tinto for exploration and deposit definition purposes and were included in the drillhole database for resource estimation (CR series of drillholes). Another 106 holes, for 24,279m, were drilled for geotechnical, hydrogeological and metallurgical purposes. The details of the drilling programs are discussed more completely in the Section 9 on drilling.

More recently in late 2004/early 2005, CLC conducted an additional infill drilling program with the objective of providing more detailed information for short-term mine planning and this data will also serve to validate the resource model. At the time of this report, the results were not complete and a comprehensive review will be conducted in the future when all the results are available.

Development and Operations

According to Inmet’s 2007 AIF, the following updates have been reported since the May 27, 2005 PAH Report.

Inmet adjusted the original mine plan from the May 2004 feasibility study for the project to include some underground mining, which will be accessible beginning in 2012 from a ramp in the pit. The underground mine should provide significant benefits because it will reduce waste material by approximately 21Mt, and gives Inmet access to approximately 40,000 additional tonnes of contained copper. The adjusted mine plan calls for 738,000t of ore to be extracted from underground, mainly by drift and fill methods, over a period of four years from 2013 to 2016;
Pre-stripping commenced in March 2006 and by the end of the year, one-third or 9Mm³ of the overburden had been removed. Inmet will continue to remove overburden in 2007, and expect to reach ore in the fourth quarter. We expect Las Cruces to mine between 0.6 and 1.3Mt of ore and up to 30Mt of waste each year at an overall strip ratio of 12.7:1. While this ratio is unusually high, it is offset by the very high ore grade. During years 5 through 9 of production (2012 – 2016), we plan to supplement ore from a small underground operation which will mine approximately 740,000t of high grade ore not accessible from the open pit;
The mine will produce copper cathode on site, using technology developed by Outokumpu Technology Oy. The process involves crushing and grinding the ore and then leaching it in the presence of ferric sulphate, sulphuric acid and oxygen in agitated tanks at atmospheric pressure. Once in solution, the copper will be recovered by standard solvent extraction and electrowinning technology. Recoveries of copper are expected to exceed 90% and the plant is designed for a production rate of 72,000t of copper cathode


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per year. The production of cathode sheets rather than concentrate will diversify Inmet’s copper production, which has been historically in the form of concentrate, and reduce our exposure to smelter and refining treatment charges, which can be volatile. Because cathode copper can be transported by truck, train and ocean-going vessels to plants in Europe that produce wire rod and other copper products, it also lowers freight risk;

Inmet completed basic engineering in April 2006, and a project budget of €380 million was approved at that time to complete the engineering design, procurement, construction and commissioning of the mine and a hydrometallurgical plant capable of annually producing 72,000t of copper as cathodes. A subsidiary of SNC Lavalin Group Inc. has been appointed as lead engineer for the engineering, procurement and construction management of the plant and associated infrastructure. Outokumpu Technology Oy has been contracted to provide grinding, leaching, solvent extraction and electrowinning technology; and
By the end of December 2006, Inmet had completed 55% of the detailed engineering and made commitments to procure €124 million of equipment and services. Total expenditures in 2006 were €65 million. Physical construction of facilities commenced in November 2006 with the building of temporary offices. Civil works on the plant itself began in January 2007

Conclusions and Recommendations

SRK notes that some of the information residing in the public domain generated internally by Inmet, especially Mineral Resources and Ore Reserves, require NI 43-101 compliance for public disclosure, the estimations and conversions are reported as CIM compliant.

SRK is not aware of any issues that have not been otherwise disclosed in this report which would materially affect the Project and makes no further recommendations in regard to the Project or the royalty holder.


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1 Introduction (Item 4)

1.1 Terms of Reference and Purpose of the Report

The information contained in this report is effective as of February 5, 2008.

1.2 Reliance on Other Experts (Item 5)

NI 43-101 contains certain requirements relating to disclosure of technical information in respect of mineral projects. The information contained herein with respect to the Project is primarily extracted from the PAH Report and general information available in the public domain, referred to in Section 20 and provided by IRC throughout the course of SRK’s investigations. SRK did not conduct a site visit, did not independently sample and assay portions of the deposit and did not review the following items prescribed by NI 43-101: (i) geological investigations, reconciliation studies, independent check assaying and independent audits; (ii) estimates and classification of Mineral Resources and Ore Reserves, including the methodologies applied by the mining company in determining such estimates and classifications, such as check calculations; or (iii) LoM plan and supporting documentation and the associated technical-economic parameters, including assumptions regarding future operating costs, capital expenditures and saleable metal for the mining asset.

Also, SRK did not independently sample and assay portions of the deposit because it was not permitted to access the relevant material and data.

1.2.1 Limitations and Reliance on Information

The royalty holder is not entitled to detailed or confidential information regarding the Project. Due to the royalty holder’s lack of legal rights to obtain this data, SRK was unable to conduct detailed, thorough and independent assessments. Therefore the data available for the preparation of this report was significantly limited, especially in consideration of the requisite reporting requirements of NI 43-101.


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This report includes technical information, which requires subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, SRK does not consider them to be material to the findings and use of this Technical Report.

The achievability of LoM plans, budgets and forecasts are inherently uncertain. Consequently, actual results may be significantly more or less favorable.

SRK was unable to conduct an in-depth review of mineral title and ownership; consequently, no opinion will be expressed by SRK on this subject.

Pursuant to Part 9.2(1) of NI 43-101, SRK is not required to perform an onsite visit of the Project site, nor is it required to complete those items under Form 43-101F1 that require data verification, inspection of documents, or personal inspection of the property. The royalty holder is relying on the exemption available under Part 9 of NI 43-101, as it has requested but was denied access to the necessary data from Inmet and is not able to obtain the necessary information from the public domain. SRK notes that some of the information residing in the public domain generated internally by Inmet, especially Mineral Resources and Ore Reserves, require NI 43-101 compliance for public disclosure, and as such are assumed to be NI 43-101 compliant.

Inmet has not reviewed this report and takes no responsibility nor assumes any liability for the statements in this report. No express or implied representation or warranty has been made by Inmet that the contents of this report are verified, accurate, suitably qualified, reasonable or free from errors, omissions or other defects.

1.3 Effective Data (Item 24)

Unless otherwise specifically noted, the information contained in this report is effective as of February 5, 2008.

1.4 Qualifications of Consultants (SRK)

The SRK Group comprises 750 staff, offering expertise in a wide range of resource engineering disciplines. The SRK Group’s independence is ensured by the fact that it holds no equity in any project and that its ownership rests solely with its staff. This permits SRK to provide its clients with conflict-free and objective recommendations on crucial judgment issues. SRK has demonstrated track record in undertaking independent assessments of Mineral Resources and Ore Reserves, project evaluations and audits, technical reports and independent feasibility evaluations to bankable standards on behalf of exploration and mining companies and financial institutions worldwide. The SRK Group has also worked with a large number of major international mining companies and their projects, providing mining industry consultancy service inputs.

Neither SRK nor any of its employees and associates employed in the preparation of this report has any beneficial interest in the royalty holder or in the assets of the royalty holder. SRK will be paid a fee for this work in accordance with normal professional consulting practice.

Listed below are the individuals, based in SRK Group’s Denver, Colorado office, who have provided input to this technical report:

Dr. Neal Rigby, CEng, MIMMM, PhD; and


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Las Cruces Project Royalty	NI 43-101 Technical Report

Katherine L. Garramone, Technician.

Dr. Neal Rigby is the Qualified Persons (QP) responsible for the content, compilation, and editing of all sections of this Technical Report. The Certificate of Author is provided in Appendix A.


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2 Property Description and Location (Item 6)

2.1 Property Description

2.2 Property Location

The Las Cruces Project is located in the Sevilla Province of southern Spain, about 20km northwest of the Province capital city of Seville. This province is part of the Andalusia region. A location map is presented in Figure 2-1. Access to the project site is excellent by well-maintained all-weather paved roads. Rail service is available in Seville, as is an international airport with connections throughout Europe. Port facilities are available in Huelva approximately 80km to the southwest. Las Cruces is situated in the Guadalquivir River basin. The Guadalquivir basin is relatively flat and extends approximately 420km from the Alcaraz and Segura mountain systems in the east to the Costa de la Luz on the Atlantic Coast in the west. The Guadalquivir basin is set between a long range of low mountains to the north (Sierra Morena) and a group of formations making up the Cordilleras Beticas mountain range to the south. The regional climate of the area is characterized as Mediterranean, with gently rolling hills of arable land. The geographic coordinates of the property are N37°30’, W06°06’.

2.3 Project Ownership

Inmet announced on May 3, 2005 that it had entered into an agreement with MK Resources and its majority shareholder, Leucadia National Corporation, to acquire a 70% interest in Cobre Las Cruces S.A. (CLC), the owner of the Las Cruces Project. Leucadia owns 72% of MK Resources. Immediately following the closing of the agreement, MK Resources [became] a wholly-owned subsidiary of Leucadia, which then [sold] a 70% interest in the Las Cruces Project to Inmet in exchange for the issuance by Inmet to MK Resources of 5.6 million common shares of Inmet. Leucadia [retains] a 30% interest in the Las Cruces Project through MK Resources Under the terms of the agreement, Inmet [has] the authority to oversee the development and operation of the Las Cruces Project, subject to certain protective minority rights of Leucadia. Inmet and Leucadia [committed] to the project development costs and provide guarantees of project financing, in proportion to their respective interests in the project. Inmet [assumed] its share of development expenditures for the Las Cruces Project as of April 1, 2005.


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2.4 Mineral Tenure

CLC has been granted a mining concession for subsurface minerals through Mining Concession No. 7532, granted by the Regional Ministry for Employment and Technological Development of the Province of Andalusia. The Mining Concession was granted August 6, 2003, following CLC’s 2001 publication of an Environmental Impact Study and the subsequent May 9, 2002 issuance of a Declaration of Environmental Impact. Figure 2-2 shows the mining concession location relative to the project area. Other permits and surface ownership rights are addressed separately.

2.5 Surface Land Ownership

Land ownership within Spain is procedurally clear, and is established via survey. Mineral rights and surface land are separately transferable under Spanish law. CLC [purchased] the surface rights to the land above the deposit and to adjacent lands needed for infrastructure and right of ways, which are necessary to develop and operate the project.

Development of the project [required] the acquisition of 895ha of land (Phases 1 and 2) from 52 separate landowners, together with the leasing of an additional 50ha for three years, through the end of the first year of operation. The leased area comprises part of the north mine waste dump; it will be formed and reclaimed early in the Project life and then returned to the property owner. Additional land might be necessary to accommodate operations in later years.

Through March 2005, CLC [had] entered into agreements to purchase more than 99% of the land (Phase 1), to enable project construction and development. CLC has stated that all of the required land will be purchased.

CLC has purchased the surface rights to the land above the deposit and to the adjacent lands needed for infrastructure and is in the process of finalizing all of the rights of ways that are necessary to construct and develop the project. The project also covers land in the public domain for which applicable permits have been obtained. (Inmet AIF, March 2007) The ore deposit is located on the eastern side of the property, at a depth of approximately 120m.

2.6 Royalties, Agreements and Encumbrances

The description of the royalty is as follows:

A 1.5% SR royalty is to be paid to IRC by CLC in respect of all sales realized in any calendar month in which the official monthly average LME Cash Settlement price for Grade A copper is equivalent to or greater than US$0.80/lb-Cu; and
Allowable deductions in calculating sales revenue include 1) freight, transportation, and insurance to the point of sale; 2) government royalties; and 3) sales expenses.

2.7 Environmental Liabilities and Permitting

Las Cruces is a Greenfields operation, comprised of agricultural land well away from populated areas. No existing liabilities are known to exist. As the project is developed, [Inmet] will incur reclamation liabilities that are clearly identified in the planning, with appropriate allowances


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within the cash flow. Permitting for the Las Cruces project will [includes] both environmental approvals for operation of the mine, processing plant and mine waste disposal as well as land use permits and approvals. A discussion of the environmental permitting status [as of March 2005] is presented in Section 17. Land use permits include obtaining authorization for construction of crossings of the water supply pipeline over railways, highways, and a water supply canal, as well as various crossing of power line, telephone line and underground utility line rights-of-way. Works and activity permits [are] required from local community councils.

Based on the public domain information and data reviewed, SRK is not aware of any existing environmental liabilities on the Project.


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Figure 2-1: Las Cruces General Location Map


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Figure 2-2: Mining Concession Limits


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3 Accessibility, Climate, Local Resources, Infrastructure and Physiography (Item 7)

3.1 Topography, Elevation and Vegetation

The topography around the project is characterized by gently rolling hills of arable land. Elevations at the site range from 15 to 45masl.

3.2 Climate and Length of Operating Season

The regional climate is Mediterranean or semi-arid with hot, dry summers (May through September) and warm, wet winters. Prolonged droughts periodically occur. Mining operations will be possible year-round.

Wind is generally from the southwest and brings hot and humid air into the area that generates autumn and winter rainstorms, and intense spring storms. Summer maximum temperatures average about 30ºC, with the maximum occasionally reaching 40 ºC in July and August. Winter maximum temperatures average about 15ºC and the minimum temperature may drop to as low as –5 ºC in December and January.

Mean annual rainfall is about 550mm and ranges from about 300 to 1,000mm. The heaviest one-day rainfall on record was 100mm.

The site is characterized as having a moderate seismic activity. A seismic potential of a magnitude 5.4 (Richter Scale) event with a recurrence interval of 250 years has been defined for the Seville area by the Instituto Geografic Nacional and the Center Internacional de Sismologia.

3.3 Access to Property

Access to the project site is excellent by well-maintained all-weather paved roads. Highway N-630 presently is in service to the east side of the property, while Highway SE-520 crosses the southern side of the property. The A66 freeway is presently under construction.

Rail service is available in Seville, with high-speed passenger rail service to Madrid. Seville offers an international airport with connections throughout Europe within 30 minutes of the site. Port Facilities are available in Huelva approximately 80km to the southwest.

Seville (Sevilla, pop. 700,000), a major Spanish city, is located approximately 20km to the south from the property, with a large and well-educated population. The village of Gerena (pop. 5,600) is located approximately 4km to the northwest of the property.

3.4 Surface Rights

CLC had made a determined effort to minimize the land requirements for the property, and developed a detailed plan for the land use at the property including adequate room for waste dumps, tailings ponds and pit slopes.


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[It was PAH’s opinion] that sufficient room [was] available for the plans as presented [as of March 2005].

3.5 Local Resources and Infrastructure 3.5.1 Power Supply

The power supply during construction will be provided through connections to the 15kV overhead lines that end at the sites of the Render Sur factory and the Seroncillo farm, with additional capacity from a 400V diesel generator with a 1MW capacity. During operation, the electrical power will be supplied through a new 220kV branch line with a length of approximately 4.2km, to be constructed for the Las Cruces project. On the Las Cruces project site a switchyard and a 40MVA transformer will distribute the electric power to the different project locations.

3.5.2 Water Supply

The plant will require 60L/s of process water, 10% of which can be water supplied from contact water (water that has been exposed to sulfide minerals) and the remainder of which must be supplied from elsewhere. The remaining required process water will be pumped 18.6km from the San Jerónimo sewage treatment plant, located to the northwest of Seville. The process water will be pumped through a pipeline to the primary supply pond in the west of the Las Cruces project area. Water from sewage treatment plant will not be available during the dry season; therefore, the primary supply pond will require sufficient storage capacity (approximately 1.3Mm ³) to store water during the five-month dry season.

Potable water will come from a well drilled into the Posadas-Niebla aquifer. CLC has received authorization from the government authority to use this source. Plans are to drill a 150m deep well.

3.5.3 Manpower

The Project is located approximately 20km from Seville and surrounded by four smaller communities. As the area is host to surrounding established and developing mine sites, skilled mining personnel are available for employment. According to Inmet, the Project has total of 87 employees and 1,658 contractors. (www.inmet-mining.com)


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4 History (Item 8)

The Las Cruces deposit was original discovered in 1994, by Riomin Exploraciones, S.A. (Riomin), a wholly owned subsidiary of Rio Tinto. The discovery was the result of the drilling on a gravimetric survey anomaly. Rio Tinto drilled the deposit between 1994 and 1999 and prepared a feasibility study in 1998.

In 1999, MK Resources (formerly known as MK Gold Company) acquired a 100% interest in the project from Rio Tinto and established CLC as the local Spanish subsidiary company. Leucadia National Corporation, a diversified investment and holding company, had a 72% share in MK Resources. In 2001, Bechtel, an international engineering and construction company, completed an independent feasibility study for the Las Cruces project. The Bechtel feasibility study incorporated results of an environmental impact study completed by FRASA Ingenieros Consultores, a team of national and international environmental engineering experts based in Madrid.

The Mining Concession was granted August 6, 2003, following CLC’s 2001 publication of an Environmental Impact Study and the subsequent May 9, 2002 issuance of a Declaration of Environmental Impact.

In November 2003, DMT-Montana Consulting (DMT-MC) prepared a new independent feasibility study based upon Outokumpu technology, followed by the Feasibility Study Addendum I in May 2004. The DMT-MC feasibility study incorporates the requirements from the Declaration of Environmental Impact, the mining concession, and various water permits into the development plan for the Las Cruces project. The DMT-MC feasibility study and addendum has been reviewed by PAH, an independent mine engineering consulting company. In its audit report, PAH stated that it did not identify any deficiencies that would preclude the Las Cruces project from meeting the designated production and cost objectives presented in the DMT-MC feasibility study.

On May 3, 2005, Inmet announced that it had entered into an agreement with MK Resources and its majority shareholder, Leucadia National Corporation, to acquire a 70% interest in the Las Cruces Project. Closing of the transaction [was] subject to the terms and conditions set out in the agreement.

In August 2005, Inmet completed the acquisition of a 70% interest in Las Cruces.


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5 Geological Setting (Item 9)

5.1 Regional and Local Geology

Subsequent tectonism during the Hercynian Orogeny of the late Paleozoic Era resulted in the general uplift of the region, with variable deformation and faulting. This was followed by weathering and erosion of the overlying host rocks during an indeterminate period between the Hercynian Orogeny and the Miocene Epoch. Late in the period of erosion and weathering, the upper part of the original massive sulfide deposit was exposed. Near surface oxidation of the sulfide minerals to an iron-oxide gossan occurred, along with the downward transport of copper leached from the oxidized gossan zone. The transported copper was precipitated and replaced unoxidized primary massive sulfide at depth, forming a secondary enrichment of the sulfide deposit. Relatively immobile gold and silver metals remained in the oxidized gossan, with local enrichments.

Subsequent transgression by a shallow sea in the middle Tertiary Period (Miocene Epoch) interrupted the weathering and oxidation process on the deposit. Submarine deposition buried the deposit under 100 to 150m of sandstone and calcareous mudstone (marl). Subsequent marine regression in Pliocene times again resulted in the subaerial exposure and erosion of the region. The current surface of the project area is on the calcareous mudstone sequence, with the Las Cruces deposit buried at depth below this sequence.

The Miocene age sandstone unit that occurs at the base of the younger sedimentary sequence overlying the paleoerosion surface is an important regional aquifer (Niebla-Posada Aquifer). This aquifer consists of a basal conglomerate overlain by semi-consolidated sandstone and averages approximately 5m thick over the deposit area.

5.2 Project Geology


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truncated by a roughly north-south trending fault; with attempts to locate the offset mineralization not having been successful so far. Along strike to the northeast, the deposit is not completely drill delineated, however the grade of the mineralization is relatively low. Due to supergene enrichment processes in pre-Miocene time, secondary sulfide mineralization formed a generally horizontal zone in the up-dip part of the primary massive sulfide. Figure 5-1 shows a geology-drillhole cross section looking west and reflects the closely spaced drilling of a local statistical cross. Figure 5-2 shows a plan view of deposit. Deposit mineralization is discussed in Section 7.


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6 Deposit Type (Item 10)

The Las Cruces deposit occurs near the eastern end of the Iberian Pyrite Belt, a 250km long and 40km wide geologic belt that extends eastward from southern Portugal into southern Spain. The belt is host to more than 100 mineral deposits, some of which were exploited for metals as long ago as pre-Roman times. Mineralization at Las Cruces, as in most other Iberian Pyrite Belt deposits, consists of syngenetic massive sulfides containing polymetallic mineralization. The massive sulfide is hosted by late Devonian to Early Carboniferous Period volcanic and sedimentary rocks, deposited in a submarine setting within a narrow and relatively shallow intracontinental sea, characterized by bimodal volcanism and sedimentation.

Volcanogenic massive sulfides are deposited as submarine hot spring exhalative metal sediment accumulations, typically in tectonic areas of active submarine volcanism, including rift spreading centers and island arc subduction zones. In these geologic environments the volcanic rocks may be interlayered to a variable degree with contemporaneous volcanic sediments. Massive sulfide bodies are deposited as metallic sulfide sediments, with a variable component of volcanic sediment and silica, in conformable layers blanketing the area surrounding the exhalative center. Massive sulfide layers consist predominantly of pyrite and/or pyrrhotite, with variable base metal sulfides commonly including chalcopyrite, sphalerite, and galena. Subsequent tectonism may result in variable uplift, folding, and/or faulting of the original massive sulfide deposit.

The Las Cruces massive sulfide was subjected to post-depositional tectonism during the Hercynian Orogeny of the late Paleozoic Era resulting in the general uplift of the region, with variable deformation and faulting. This was followed by weathering and erosion of the overlying host rocks during an indeterminate period between the Hercynian Orogeny and the Miocene Epoch. Late in the period of erosion and weathering, the upper part of the original massive sulfide deposit was exposed. Near surface oxidation of the sulfide minerals to an iron-oxide gossan occurred, along with the downward transport of copper leached from the oxidized gossan zone. The transported copper was precipitated and replaced unoxidized primary massive sulfide at depth, forming a secondary enrichment of the sulfide deposit. Relatively immobile gold and silver metals remained in the oxidized gossan, with local enrichments.

The massive sulfide body and copper enriched zones was subsequently buried under 100 to 150m of Miocene age sediments.


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7 Mineralization (Item 11)

7.1 Secondary Sulfide Mineralization

The pre-Miocene paleosurface that cuts the up-dip part of the massive sulfide deposit is planar, with minor irregularities and a gentle dip to the south. The central and western part of the deposit, below the paleosurface, was most affected by the pre-Miocene weathering and oxidation. The uppermost part of the massive sulfide mineralization has been completely oxidized to an iron oxide gossan, with local enrichment of gold and silver. The gossan is also enriched in lead, reportedly as the mineral galena. Infilling of the porous gossan with variable carbonate occurred with the subsequent sedimentary marl deposition. PAH’s review found that the gossan is generally 10 to 20m in vertical thickness, with grades from all gossan drillhole samples averaging 5.9gpt gold (up to 353gpt Au) and 98gpt silver (up to 1,733gpt Ag). Average copper grade in the gossan is 0.20% . Strong silicification of the host rock occurs in the hangingwall immediately above the gossan, with erratic gold and silver enrichment. While there is potential for economic value within the gossan, limited studies have been completed. The project does not consider the gossan zone to have economic value and is not considered as part of the current resource estimate.

Underlying the iron-oxide gossan is the secondary sulfide zone in which the primary massive sulfide mineralization has been secondarily enriched by the downward migration and precipitation of copper leached from the gossan zone. The secondary sulfide mineralization is the most important economically, and the focus of the DMT-MC 2003 Feasibility Study. Secondary sulfide enrichment occurs in a roughly horizontal zone within the up dip part of the original primary massive sulfide deposit. The zone of secondary sulfide enrichment averages about 40m thick below the gossan, but in areas of intense fracturing secondary sulfide may extend up to 60m thick. The secondary sulfide mineralization is gradational downward into the primary sulfide zone. The footwall contact of the secondary mineralization with the underlying host rocks is usually sharp, as a result of the clay altered shales and volcanics being relatively impermeable. This permeability change tended to force secondary mineralization along the base of the massive sulfides, locally producing lenses of secondary mineralization that follow the contact zone down dip into the primary sulfides.

Secondary sulfide mineralization has been superimposed upon and has partially replaced the original primary sulfide mineralization below the gossan zone, increasing the quantity of copper minerals present. Secondary sulfide mineralization consisted predominantly of the deposition of chalcocite and to a lesser extent bornite and covellite by partial replacement of pyrite and other sulfides, and by infilling into any open space resulting from fracturing or original porosity. Chalcocite ranges from very fine grained (sooty) dark gray coatings and relatively unconsolidated intergrowths to disseminated consolidated grains to consolidated veins and replacement bands. Chalcopyrite content is minor. PAH’s review found that grades from all secondary sulfide drillhole samples averaged 7.0% copper (up to 39.4% Cu), 0.2% zinc (up to 36.9% Zn), 0.6% lead (up to 40.0% Pb), 0.5gpt gold (up to 62.9gpt Au), and 26gpt silver (up to 1,472gpt Ag).


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The secondary sulfide mineralization was divided into three main sub-lenses, and three minor sublenses. The boundaries between the six types are complex and gradational. The High Copper – High Density (HCH) lens consists of secondary enrichment of dense primary massive sulfides. The High Copper – Low Density (HCL) lens consists of secondary enrichment of primary semi-massive sulfides, footwall stockworks, or footwall lithologies of lesser density. The High Copper lens Number 4 (HC4) is a spatially distinct lens of HCH secondary mineralization that follows the contact zone down dip into the primary sulfide zone. In addition, three minor sub-lenses occur: the High Copper Intrusive (HCI), the High Copper Footwall (HCF), and the Medium Copper Low Density (MCL); all three of which were grouped with the main lenses for resource estimation purposes.

7.2 Primary Sulfide Mineralization

Original primary massive sulfide mineralization (CZ lens) contains massive to semi-massive sulfide minerals (generally more than 80% sulfide). Pyrite is the predominate sulfide mineral, with lesser finely intergrown sphalerite, galena, and chalcopyrite, as well as minor enargite, tennantite, and tetrahedrite. PAH’s review found that grades from all primary sulfide drillhole samples averaged 3.2% copper (up to 18.9% Cu), 1.1% zinc (up to 11.9% Zn), 0.4% lead (up to 5.4% Pb), 0.5gpt gold (up to 1.63gpt Au), and 20gpt silver (up to 174gpt Ag). Most of the primary massive sulfide mineralization occurs down-dip to the northwest and are not considered for the current feasibility study. In the footwall below the primary massive sulfide is a stockwork of interconnected pyrite veins and veinlets, with local higher grades of copper and zinc. The primary sulfide (CZ) is not considered as part of the current resource estimate.

SRK did not verify Project mineralization data including; length, width, depth and continuity, as well as, type character and distribution of the Las Cruces project due to the lack of access to specific Project data.


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8 Exploration (Item 12)

8.1 Deposit Exploration

SRK did not conduct an in-depth review of the surveys, procedures or methodologies used in the exploration of the Las Cruces Project. Consequently, an interpretation of whether industry standards were used and the results of the exploration will not be expressed by SRK, due to the lack of access to specific Project data.


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9 Drilling (Item 13)

9.1 Type and Extent of Drilling

9.1.1 Core Drilling

A total of 277 core holes, for 82,352m, were drilled by Rio Tinto between 1994 and 1999, for exploration and deposit definition purposes and were included in the drillhole database for resource estimation (CR series of drillholes). Another 106 holes, for 24,279m, were drilled for geotechnical, hydrogeological and metallurgical purposes. Drilling of the CR series of holes was conducted by three different companies: Almaden (about 55% of holes), Insersa (about 40% of holes), and Iberica (about 5% of holes). Most of these holes were drilled vertically (70% of holes), with the remainder drilled at an angle (30%). Hole collar locations have been surveyed by a professional surveyor using a theodolite. All core holes were surveyed down-hole using a multi-shot or single-shot instrument.

Drilling of the CR holes by all companies consisted of rotary tricone drilling through the marl and sandstone overlying the deposit, followed by core drilling in the underlying host rock sequence. Core drilling was generally conducted to obtain larger diameter core for analysis, metallurgical purposes, and for reference. Core samples were taken at nominal 1m lengths (+50%) based on geological or mineralogical sampling breaks. The core was logged for geologic and mineralogic information and the percent core recovery recorded.

Drilling of the secondary sulfide deposits is on a variable grid with a drillhole spacing of 50m x 50m. Two statistical crosses were drilled, a complete cross in the western part of the deposit and a less complete cross in the central part of the deposit, with a nominal 12.5m spacing between these holes and inclined at a 60º angle. PAH [believed] that this drilling [provided] adequate coverage of the deposit, with some in-fill recommended near holes in which core recovery was an issue or were wider spacing occurs. Independent Mining Consultants (IMC) expressed a similar opinion in the 2001 Bechtel Feasibility and CLC concurs with these recommendations. Figure 9-1 shows the drillhole locations with some of the drillholes marked to indicate: 1) holes for which low core recovery from the secondary sulfide (<80%) excluded the entire drillhole from being used in the resource modeling; 2) holes for which low core recovery from the secondary sulfide (80%) excluded more than one half of the secondary sulfide; and 3) initial holes for which core recovery from the secondary sulfide was not recorded so a 100% core recovery was assigned.

Drilling of the secondary sulfides consisted of 88% PQ size core (85mm core diameter), 9% HQ size core (64mm core diameter), and 3% NQ size core (48mm core diameter). Average copper grade for the three core sizes shows that the larger PQ core samples average about 15% higher in copper than the smaller HQ sized core and that the HQ sized core averages about 65% higher in copper than the smaller NQ sized core. Because of the increasing core surface area/core volume with decreasing core size, this [provided] evidence of the preferential loss of copper minerals from the core surface.


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Subsequent to feasibility resource modeling and reserve estimation work, additional infill drilling was conducted by CLC from late 2004 to early 2005, with the objective of providing more detailed information for short-term mine planning. At the time of this report, the results were not complete. This additional drilling is discussed at the end of the resource section (Section 15), otherwise the sample data analysis and resource modeling results presented in this report do not reflect this new information.

9.1.2 Reverse Circulation Drilling

Eighteen reverse circulation holes (311mm diameter) were drilled in 1998 as twins of original PQ sized core holes as checks on the sampling and to provide material for metallurgical testwork. The reverse circulation sample results compared poorly on a sample-by-sample basis with the original core hole grades and were generally lower grade. On average, the reverse circulation samples were 43% lower in grade than the adjacent core hole (4.46% Cu compared to 6.41% Cu in the core holes). It is believed by CLC that the differences are attributable to short-range lateral variations in grade and also that significant amounts of fine-grained chalcocite were lost during the reverse circulation drilling and as a result the grades are not representative. These holes have not been included in the drillhole database for resource estimation and so these holes are not considered further in the resource modeling and estimation process.

9.1.3 Interpretation

A summary of the interpretation of the drilling results will not be expressed by SRK, due to the lack of access to specific Project data.


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10 Sampling Method and Approach (Item 14)

10.1 Sampling Methods

10.1.1 Core Sampling

Core drilling was generally conducted to obtain larger diameter core for analysis, metallurgical purposes, and for reference. Core samples were taken at nominal 1m lengths (+50%) based on geological or mineralogical sampling breaks. The core was logged for geologic and mineralogic information and the percent core recovery recorded. The core (PQ sized core for the majority of samples) was sawed into quarters, with one quarter used for the chemical analysis, while one quarter was stored in the core box for reference. The remaining material was stored in a commercial freezer in Seville to minimize sample oxidation, for use in subsequent metallurgical testing.

Drilling of the secondary sulfides consisted of 88% PQ size core (85mm core diameter), 9% HQ size core (64mm core diameter), and 3% NQ size core (48mm core diameter). Average copper grade for the three core sizes shows that the larger PQ core samples average about 15% higher in copper than the smaller HQ sized core and that the HQ sized core averages about 65% higher in copper than the smaller NQ sized core. Because of the increasing core surface area/core volume with decreasing core size, this provides evidence of the preferential loss of copper minerals from the core surface.

HQ diameter core was drilled from hole CR01 to CR028 and was largely halved for an analytical sample. Holes from CR029 to CR075 were generally drilled with larger PQ sized core and halved for an analytical sample. Holes from CR076 were largely drilled with larger PQ sized core and quartered for an analytical sample. As shown in Table 10.1.1.1, of the total of 4,407 secondary sulfide core samples, of which 231 were sampled in their entirety (5%), 632 samples consisted of a half of the core (14%), and 3,544 samples consisted of a quarter of the core (81%).

Table 10.1.1.1: Las Cruces Core Sample Size Distribution


Sample Size	NQ Core (48mm dia.)	HQ Core (64mm dia.)	PQ Core (85mm dia.)	All Core Sizes
Entire Core	0	0	231	231
Half Core	88	347	197	632
Quarter Core	26	52	3,466	3,544
All Sample Sizes	114	399	3,894	4,407
Note: Results are for the drillhole database provided to PAH.

As a result, the predominant sample by far was PQ sized core for which the sample consisted of a quarter of the core (3,466 samples or 79% of all of the secondary sulfide core). PAH [found] that the core sample used generally provides an adequate sized sample. PAH [noted] that one quarter of the larger PQ sized core ideally provides a sample comparable in size to one half of the smaller HQ sized core. PAH [questioned] the comparability of quarter samples of NQ sized core, as quarter samples of NQ sized core ideally only provides about 30% of the material of


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quarter samples of PQ sized core. PAH [noted], however, that there are a relatively small number of this sample size.

10.1.2 Core Recovery

Core recovery from the main mineralized lenses (HCH, HCL, and HC4) ranges from 0 to 100%, with an average of about 80%. With the application of an 80% core recovery cut-off, the average recovery was raised to about 93%. Geologic observation during the logging of the Las Cruces core provides good evidence of the variable loss of core material due to the abrasion and erosion of friable and sooty, chalcocite-bearing mineralization. The apparent preferential loss of this material during drilling and sampling has tended to decrease the relative content of copper minerals, with a proportionally less decrease in the relative content of the other associated minerals, including pyrite, quartz, sphalerite, and/or galena. As a result of this, the copper grade in the actual analyzed sample of the recovered core tends to be less than in the copper grade in the original in-place location. For an independent opinion, CLC contracted Dr. Stephen Henley of IMC (UK) to evaluate the issue and he came to a similar conclusion. IMC, however, concluded in the 2001 feasibility study that higher-grade copper occurs preferentially in more competent rocks, whereby chalcocite acts as cement holding fragments together and improving recovery. Previously, Rio Tinto in their 1998 feasibility study, came to a similar conclusion to that of CLC and DMT in the current feasibility.

PAH [found] that the majority of the geologic and statistical data provide evidence for the preferential loss of chalcocite, although reduction in copper grade is not a uniform systematic statistical relationship, but is somewhat irregular due to chalcocite distribution both as more durable masses and elsewhere as friable aggregates. It appears that both high core recovery and low core recovery zones have been subject to variable chalcocite loss. As such, core recovery itself may provide a good general indication that original in-place copper grades are higher than reported, but are not an absolute indication on a sample-by-sample basis that the original in-place copper grade was higher. The effect of the preferential loss of chalcocite tended to increase the difference between original in-place copper grade and the actual assayed copper grade at progressively lower core recoveries. This is significant in terms of generating a reasonable resource model, resulting in the question of what sample core recoveries result in reliable enough copper grades (see subsequent discussion in resource modeling subsection).

SRK was unable to complete a detailed review of the sampling method and approach due to the lack of access to specific Project data.


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11 Sample Preparation, Analyses and Security (Item 15)

11.1.1 Reverse Circulation Drilling

11.2 Sample Preparation and Assaying Methods

Sample preparation for up to hole CR128 was conducted at Anamet Laboratory in Avonmouth, United Kingdom. After this, a preparation facility was set up on site, with samples from CR129 to CR277 prepared on site under Anamet supervision, with the pulp then sent for analysis. For routine chemical analysis, the quartered core sample (PQ sized core for the majority of samples) was crushed to minus 2m size and then split into 250g quantities. The 250g samples were then pulverized to approximately to minus 200 mesh (-74µ) for analysis. Coarse reject material was also stored in a freezer.

Most of the assaying was done at the Anamet Laboratory (Rio Tinto) in Avonmouth, United Kingdom. From hole CR257 to CR277 (the last 20 holes), OMAC Laboratory was used as the primary laboratory because the Anamet Laboratory was no longer available. Analysis for copper, lead, zinc, silver was by digestion of 0.5g of pulverized material in a four-acid solution (HCl, nitric, HF, and perchloric) with an atomic absorption spectophotometery determination. Gold analysis was generally by fire assay on 20g of pulverized material. For drillholes CR001 to CR081 and CR082 to CR106, a fire assay was conducted only if an initial atomic absorption spectophotometery determination indicated gold above 0.5gpt or 0.25gpt, respectively. Sulfur was analyzed by a Leco furnace procedure. Other trace elements were also analyzed, including bismuth, cadmium, tin, arsenic, antimony, barium, and mercury. PAH [believed] that the analytical procedures are generally reasonable, but that for gold it is more standard to use a 30g pulp for fire assay in order to provide more representative results. For potentially coarser gold particles present in the gossan zone, there is some evidence that the analytical protocol may underestimate the gold grade, which is being evaluated further by CLC. All logging and sample data were entered into an Access database, with analytical data imported from electronic files from the laboratory.


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11.2.1 Density Testwork

Density tests were generally conducted by Rio Tinto on every assay sample in the Paleozoic age host rocks and mineral deposit, for a total of 15,787 tests. Density was measured on whole core prior to splitting, by measuring the sample weight in air and in water, with the density given by the weight in air divided by the difference between the weight in air and in water. This method works well for competent pieces of core, but is difficult for fragmented pieces. Where this was the case, a representative sub-sample was taken and shaped in such a way as to not allow trapped air to be caught under it. To insure quality control, a standard weight was measured and no weighing bias was found by Rio Tinto.

PAH [noted] that the samples were not sealed to preserve natural voids and the samples were not dried prior to measurement, as is standard practice. Due to one or two day delays in measuring core density, DMT has stated that the bulk of the contained moisture was evaporated and that the cores samples were considered to effectively be dry, except for water in sealed pores which will have a minimum effect of resource calculations. Rio Tinto checked the effect of moisture content for 219 pieces of core by measuring dry core density (naturally dry from two year storage) and then measuring wet core density (submerged in water for three days). The difference between dry and wet density was about a 1% difference, which PAH [concurred] with DMT that it does not believe is very significant.

A potentially greater source of error is related to open vugs and pore space, which fills with water immediately upon submersion during testing. Geologists and hydrologists working on the project have visually estimated that permeable porosity in the mineralization is about 3%, which is factored into the resource tonne totals. Larger vugs were reported by Rio Tinto to be insignificant and they estimated the open vug space to be a few percent, which is not factored into the resource tonne totals. For the HCH/HCL lens, the average core density was 3.62t/m ³, which was subsequently factored during the resource tabulation using a 3% porosity factor (average adjusted density of 3.51t/m ³).

The iron-oxide rich gossan directly overlying the secondary sulfide mineralization is described as having a crumbly nature that has produced much broken core and core loss. DMT states that it is likely that a certain amount of void space is present, but is quite variable and difficult to estimate. For the gossan, the average core density was 2.46t/m ³, which included an average 5.1% moisture (average dry density of 2.34t/m ³).

For the Miocene age marl and sandstone overlying the deposit, density and moisture content was conducted on geotechnical core. The core was sealed in wax to preserve the moisture content of the sample. Moisture content was determined and then the dry density was determined. The average core density of the marl was 2.02t/m ³, which included an average 27% moisture (average dry density of 1.58t/m ³) based on 88 samples. The average core density of the sandstone was 2.27t/m ³, which included an average of 19% moisture (dry density of 1.74t/m ³) based on three samples.

PAH [found] that the density determinations have been conducted in an acceptable manner and that the results are representative of the various rock types present.


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11.3 Quality Controls and Quality Assurance

PAH [understood] that reasonable and customary procedures were employed for the security of the samples. Samples were maintained under the control of the site geologists, shipping agents, and the analytical laboratory.

SRK was unable to complete a detailed review of sample preparation, analyses and security due to the lack of access to specific Project data.


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12 Data Verification (Item 16)

12.1 Assay Checks

Routine quality control procedures were used by the Anamet Laboratory throughout the analytical program, including the use of standard samples, duplicate checks on coarse reject materials, and duplicate checks on pulverized material. Check assays were conducted by OMAC for copper, lead, zinc, silver, and gold. The analytical procedure used a two-acid solution (HCl and nitric) for copper, lead, and zinc, which [PAH noted as] not as strong of an acid. For silver, an aqua-regia digestion was used. For gold a 30g pulp was used. All other analytical procedures were the same as Anamet Laboratory. Two reference standard samples consisted of a high grade and low-grade material from Las Cruces and were included in the sample submissions to Anamet Laboratories. The data from the quality control checks did not indicate any significant bias or quality control issues. PAH [noted] that a visit was not made to the Anamet Laboratory to see the operation firsthand, nor is PAH familiar with the general historical performance of the facility.

A second core sample by Rio Tinto was submitted to Anamet Laboratories for about one out of 20 samples (5%) as a check of sample grade reproducibility (1,223 check samples). The checks were conducted on a second quarter of the drill core (PQ sized core for the majority of samples) and as such reflects a combination of natural grade variability across the core, sample preparation error, and analytical variance. The results showed no significant copper grade bias between the original and the check; however, a fair amount of variability existed, believed to largely be the result of the natural irregular character of the chalcocite enrichment.

A round robin program, consisting of the analysis of 10 samples by five different laboratories, was conducted by Rio Tinto to assess variability between the labs. Some variability was observed in copper grades between labs, but in general, the results show that the Anamet copper results are comparable to those from the other laboratories. An independent check was conducted in 2000 by IMC, as part of the previous work for the 2001 Bechtel Feasibility Study. Fifty samples were randomly selected for copper and gold analysis from remaining core stored in the Sevilla refrigeration facility. The core was crushed to minus 2cm and a 1,200g split sent to Actlabs-Skyline Laboratories in Tucson, Arizona. The results for copper showed these check samples to compare reasonably with original grade values. The results for gold were less consistent, with Actlabs-Skyline Laboratory consistently higher than the Rio Tinto values, generally on the order of 30% or so.

PAH [found] that Rio Tinto had a reasonable sampling and assaying protocol in place for the analysis of drillhole samples, complete with a reasonable quality control program. PAH [noted] that the quality control program did not incorporate “blank” samples known to have no grade; however, this is considered a minor issue. Check assays indicate reasonable analytical precision (repeatability) and analytical accuracy (comparability to know reference material), but was affected by natural variability of the sample material.


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12.2 PQ Twin Drilling

Rio Tinto drilled four PQ sized core holes (CR-265, CR 266, CR 267, and CR268) as twins of original core holes in order to show that earlier holes with poor core recovery preferentially lost chalcocite, resulting in lower copper grades than was realistic. The twin PQ sized core holes were very carefully drilled (triple tube with steel inserts) to maximized core recovery and relatively few sample intervals fell below an 80% core recovery. Comparison of the original hole with the twin was affected by short-range lateral variability in copper grade that made for an irregular comparison on a sample-by-sample basis. The weighted averages of the 116 original samples to the new twin samples showed that on average the recovery increased by 21% (71% to 92% recovery) and the average copper grade increased by 8.5% (5.07% to 5.50% Cu). PAH [believed] that this provides further evidence of the tendency to have better copper grades with better core recovery (preferential loss of chalcocite). The four PQ twin holes were included in the drillhole database for resource estimation.

12.3 Other Verification

The project has been sampled and tested over several years by various geologists, operating under the ownership of two different mining companies. The mineralization can be observed visually, so there is no doubt of its existence. As such, PAH did not conduct any additional verification of the sample grades.

12.4 Drillhole Database Verification

All data for the Las Cruces project is stored in a database. The raw lengths for the geotechnical sections of the log were stored and all calculations and outputs were carried out on reports or forms backed by queries. The different lithological units were coded for input in the digital database. Sample intervals and sample type as marked for cutting by the geologist was entered manually from the sample control sheets filled out by the sampling technicians. Assay data was returned from the laboratory on spreadsheets. The data loading and merging with the assay table was carried out directly from the imported spreadsheets. Duplication, overwriting and missing data was all flagged as part of the import control routine. All collar information is entered manually from the survey report sheets. Downhole survey data was read from the film produced from the survey equipment and entered manually.

A wide range of checks were made to ensure parity. A number of spot checks were also made on the hangingwall and footwall surfaces and also on the data processing to produce the final estimation figures allocated to the block model. Random checking of the database was also done and only minor errors and missing values were found mainly within stockwork and CZ zones. A considerable amount of core recovery and core diameter entries were missing and were subsequently entered.

In addition to the work conducted by others at the various stages of Feasibility studies, PAH conducted independent checks for integrity and statistical bias as well.

SRK has not reviewed any of the project data directly and cannot comment on the result of data verifications.


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Pursuant to Part 9.2(1) of NI 43-101, the royalty holder is not required to perform an onsite visit of the Project site, nor is it required to complete those items under Form 43-101F1 that require data verification, inspection of documents, or personal inspection of the property. The royalty holder is relying on the exemption available under Part 9 of NI 43-101, as it has requested but was denied access to the necessary data from Inmet and is not able to obtain the necessary information from the public domain.


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13 Adjacent Properties (Item 17)

The Las Cruces deposit is a blind deposit that does not outcrop, due to the 100 to 150m of sandstone and calcareous mudstone (marl) that was deposited on top of the deposit. No other deposits have been found in the immediate area, but exploration is difficult given the thickness of the overburden. The nearest deposits are Aznalcollar and Los Frailes, both occurring approximately 10km to the west in the area where the host rock assemblage outcrops at the surface. The Aznalcollar and Los Frailes deposits consist of a zinc-lead massive sulfides that were in production over the last 10 to 20 years.

SRK is unable to verify adjacent property information due to lack of access to specific Project data.


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14 Mineral Processing and Metallurgical Testing (Item 18)

14.1 Metallurgical Testing Programs

Metallurgical testwork began in 1996 by the previous owner, Rio Tinto. Rio Tinto found the ore difficult to treat by conventional flotation methods. The ore was amenable to treatment by a variety of hydrometallurgical techniques, including pressure oxidation, bacterial oxidation, ferric sulfate, ammonium sulfate and ammonium carbonate leaching. Rio Tinto decided that processes utilizing ferric sulphate chemistry to oxidize the copper minerals were likely to be the most successful technologies to recover the copper. Rio Tinto investigated both atmospheric leaching with bacterial oxidation and pressure leaching technology using continuous miniplant tests with fully integrated circuits and determined that either option would be economically viable approaches.

CLC purchased the property in 1999, reviewed all previous testwork and determined that a combination of atmospheric and pressure leaching technology was a more attractive alternative. In January 2000, CLC contracted Bechtel to prepare a bankable Feasibility Study based on the pressure leaching technology developed by Dynatec. The Feasibility Study was completed in early 2001.

In January 2003, CLC contracted Lurgi Metallurgie GmbH and Outokumpu Technology Group (OTG) to prepare a revised technical and economic evaluation of the Las Cruces process plant to optimize costs for the facility, improve copper recovery and simplify the process flowsheet. OTG reviewed the project information and then recommended a revised flowsheet that incorporated atmospheric leaching, Outokumpu Compact SX technology and other innovations anticipated to improve copper recovery and lower capital and operating costs. To confirm their predictions, OTG performed further metallurgical testwork (batch and mini-continuous pilot-scale) at their Outokumpu Research Center (ORC) in Pori, Finland to demonstrate the viability of their proposed flowsheet and supply design criteria for the envisioned process facility. A Feasibility Study was completed in November 2003. Since the Feasibility Study was completed some additional testing has been performed by OTG to provide more design criteria for the [basic] engineering phase.

14.2 Metallurgical Samples

The mineralization of primary interest has been designated by CLC as High Copper (HC) with two sub-classifications, HCL and HCH. The copper in the ore is primarily chalcocite with minor amounts of covellite, bornite, chalcopyrite, tennantite-tetrahedrite complex and enargite. The main differences between HCL and HCH ore types are that the HCL has more silica and less pyrite than the HCH. Though there are wide variations in the copper content of the ore types, typical metal contents are 6 to 8% copper, zinc from 0.1 to 0.4%, lead from 0.3 to 1.4%, iron from 28 to 36% and sulfur ranges from 33 to 46%. Precious metals content of the ore types averages about 0.5gpt gold and 26gpt silver and do not dissolve to any significant extent in the selected process. Because copper accounts for over 90% of the metal value found in the ore types, metallurgical testing tended to focus mainly on recovery techniques for the copper.


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In 1999 a Master Composite (MC) was prepared from Riomin’s drill core collected during 1996 to 1998. One half of MC was used for the 1999 Dynatec testwork and the remaining material was kept in a freezer in Seville for future testing. The Master Composite (MC) weighed about 5.6t and was prepared by CLC from over 1,000 drill core intervals to represent the entire ore body. The MC consists mainly of the HCH and HCL ore types with minor amounts of the C4 ore type, a small high-copper area located in the very western portion of the deposit. The C4 ore represents only about 3% of the total ore planned for mining; however, the copper content in the C4 ore is high at about 9.5% . C4 material contains mainly covellite and responds to leaching in similar manner to the HC ores. The MC was made up in the weight ratios of HCH:HCL:C4 of 7:7:1, nearly the same as the planned ore mining tonnage ratios. In addition, numerous smaller composite and individual samples were prepared and tested from drill core that represented ore types and envisioned annual mine production. About 1.8t of the MC was sent to OTG for the 2003 testwork. Approximately 1t of MC remains in cold storage in Seville. The MC contains about 6.3% copper, nearly the same as the average ore grade planned for processing of 6.6% copper.

Figures 14-1 and 14-2 present the Metallurgical Sample Location Maps for the Master Composite sample drillholes and intervals. PAH [believed] that the MC fairly represents the ore body and the results from the testwork should be sufficient for projection of the metallurgical response of the various ore types planned for processing.

The mineralogical composition analysis of the MC indicated that over 60% of the material is pyrite, less than 1% is chalcopyrite, about 23% is chalcocite/covellite/bornite, 13% is quartz, and less than 2% as miscellaneous sulfide minerals (galena, sphalerite, tetrahedrite-tennantite). Sequential Copper Phase analysis also indicated that less than 5% of the MC was made up of primary sulfide copper minerals, therefore, the recovery of copper by ferric sulfate leaching had the potential to be above 90%.

The ore grinding requirements were initially determined by Bond Ball Mill Work Index measurements on 15 different samples of HCH and HCL ore types during 1998 testing. The HCL sample averaged 11.88kWh/t while the HCH samples averaged 13.9kWh/t with test values ranging from 9.0 to a high 16.5 meaning that the samples have a high variability. Additional work index measurements were performed in 2000 on 17 different samples and the Master Composite. The MC measurement was 10.8kWh/t and the average of all tests being 10.2kWh/t with values ranging from 7.9 to 13.3. The current grinding circuit design assumes that the work index will vary from 10 to 15kWh/t. PAH [believed] that the grinding circuit design criteria utilized work index values that are acceptable and based upon the results of the testing. PAH [also believed] that OTG’s selection of three stages of crushing followed by a single variable speed ball mill is an acceptable comminution circuit for this project, and has the flexibility to process ores of varying hardness and grades.

14.3 Batch Metallurgical Testing

In OTG’s review of the previous Dynatec pilot plant results, it was determined that atmospheric leaching was not completed after the 50 minutes of retention time thus, the need for atmospheric leaching to be followed by pressure leaching in the Dynatec process scheme. Also, based upon OTG’s experience in commercial leaching of zinc, cobalt, copper and nickel, they believed that further investigations using atmospheric leaching for extended periods of time to complete the reactions were warranted.


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Eleven batch leaching tests were performed using the MC in lab-scale Outokumpu TOP (OKTOP) reactors under atmospheric pressure with oxygen. Copper recoveries and iron extractions in the batch tests were above 92% and about 3%, respectively. Optimum results were obtained with ferric iron concentrations at 20 to 25g/L, a temperature of about 90ºC and leach times of about 8 hours.

14.4 Pilot Leaching and Solvent Extraction Testing

Based on results of the batch leach testing, two separate ten day continuous pilot plant leach tests were performed at ORC to demonstrate the technical operability of a combined atmospheric leaching and solvent extraction circuit on the MC, confirm key control parameters for the process, determine the amount of bleed solution, confirm iron extraction without lowering copper recoveries, confirm optimum leaching time, and determine solid/liquid separation parameters. The test results would also be used to generate design criteria for the process facility.

The first leach test run found that the ground MC material was too coarse (-300µ) for proper suspension in the leach reactors and for optimization of power requirements for agitation. The MC material used in the second test run was therefore ground finer (-150µ) prior to initiation of the second test run. Daily copper recoveries during the first run, however, were generally above 94% despite the suspension and high power consumption issues. Daily copper recoveries during both tests were above 94%, as long as the oxygen and ferrous to ferric iron ratios were optimized.

Overall, the continuous leach tests resulted in copper recoveries of over 94%, established a leach retention time of about seven hours, confirmed desirable iron extractions at about 3%, showed zinc extractions of about 45% that leveled off when the bleed rate to control iron levels was established, indicated solution acid concentrations of 20 to 60g/L, resulted in oxygen consumptions ranging from 36 to 54kgpt of ore and showed arsenic extractions of about 25%. Settling and filtering characteristics were also obtained during the testing. The calculated unit area for thickening of the residues ranged from 0.169 to 0.192m ²/tpd. The calculated filterability of the residue ranging from 400 to 1,200kg/m²/hr with the resultant cake moisture being about 9%. These results were used in establishing the plant design criteria for the Feasibility Study. ORC completed slurry-mix tests in larger pilot-scale vessels to better determine the requirements for agitation, wear materials and corrosion. Results of these tests indicate that 0.514kW/m ³will be required for leach agitation and that special attention should be paid to the design of the leach agitator blades to minimize erosion. Corrosion tests on three types of corrosion-resistive metal were also performed with no corrosion observed. Additional tests were performed (see Section 14.7) to confirm the agitation requirements and reactor and agitator design parameters. Continuous SX pilot tests, without electrowinning, were operated using PLS from the leach test runs and simulated the primary solvent extraction circuit. Raffinate from the SX pilot tests was returned directly to the leach tests for use as leach solution. A brief secondary SX test was performed using lime-neutralized solution obtained from primary SX raffinate solution. The primary SX circuit consisted of two extraction stages, one wash stage and two stripping stages, similar to the current plant design. The secondary SX circuit consisted of one extraction stage, one wash stage and two stripping stages, similar to the current plant design.


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The SX pilot operation demonstrated that the combination of leaching and SX was successful and resulted in predicted copper extractions from PLS of 70% and raffinate containing 10g/L copper and 70g/L acid and iron concentrations of about 15g/L. Improved removal of suspended solids in the plant operation should result in less “crud” formation. The secondary SX test indicated that copper extraction would be about 90%.

14.5 Pilot Effluent Testing

Raffinate bleed from the SX testing was neutralized with milk of lime to precipitate heavy metals. Resultant clarified solutions can then be discharged to the environment after pH adjustment. Selenium, arsenic, cadmium, copper and lead concentrations in the treated effluent were well below CLC discharge limits.

Thickening and filtering tests were performed on precipitates to generate design criteria for the process plant. Cake moistures experienced in the filtering tests ranged from 52 to 68% by weight.

14.6 Ore Variability

Thirty samples were tested to determine the variability of the mineralogy within the ore types. All samples were subjected to sequential copper phase analysis while nine samples, representing different years of the Project, were subjected to batch leaching tests.

Most of the thirty sequential analyses indicated that primary copper sulfide minerals (low anticipated copper recovery) make up less than 10% of the material; therefore, the ore types should be amenable to the atmospheric leaching technology. The nine samples representing different years of the project had an average copper extraction of about 92% with values ranging from 88 to 95% and iron extractions averaging about 5%. The C4 sample had an abnormally high iron extraction at about 14%.

Because only one leach test was performed on C4 material, additional batch leach tests were performed on C4 material prior to final design to analyze the behavior of the iron as well as obtain more information on copper extractions from this material. Additional batch leach tests with samples representative of the different project years, based upon the current mining plans, were performed to expand the database on copper and iron extractions of the primary ore types.

14.7 OTG Additional Testing

Since the November 2003 Feasibility Study, OTG has completed additional metallurgical testing at ORC to expand and/or confirm process design criteria for the project in anticipation that Basic Engineering would commence.

According to Inmet’s March 2007 AIF, development financing was arranged in December 2005 and Basic Engineering was completed in April 2006.

The additional testing was completed in November 2004 and included the following test types:

Batch atmospheric leach tests on various ore samples;
Batch settling, filtration and washing tests on the leach residues;


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Slurry agitation tests to better determine agitator type and power requirements; and
Agitator wear tests to determine optimum agitator material.

Sixteen batch-leaching tests were conducted on composite samples for the now envisioned operating years 1, 2, 3, and 8, two samples of C4 type material and a new sample of the Master Composite ore. In all tests, the samples were first ball-mill ground to about 80% passing 105µ and then leached in a 5L OKTOP reactor for 8 hours with oxygen at 90ºC and with synthetic SX raffinate to simulate the current process design. The raffinate contained 10g/L copper, 50g/L iron (ferric to ferrous ratio of 0.57:1) and 60g/L sulfuric acid.

Batch leach results indicated that the Years 1, 2 and 8 samples achieved copper recoveries that averaged 92% compared to the predicted values (based upon the earlier test results) that averaged 91%. The copper recovery for the Year 3 sample was 88% compared to the predicted recovery of 91%. The C4 sample copper recovery was 94% compared to the predicted value of 96%. The MC sample copper recovery was 90% compared to the earlier batch leach tests that averaged 92%.

Batch leach test calculated iron extractions were determined to be unreliable and compromised by the downstream thickening/filtration/washing testing of leach residue samples. The leach residue wash solution samples were set aside for several days prior to performing iron analyses which may have resulted in precipitation of iron from solution. Iron extractions from earlier tests were calculated from solutions immediately after leaching. Therefore, OTG recommends that iron extractions from all the earlier tests of 3% be used in the plant design.

The theoretical oxygen consumption was also calculated from the batch leach testwork results assuming 3% iron extraction and analyses of the leach residue for elemental sulfur and sulfate. Based upon the copper content of the ore, the theoretical oxygen consumption is estimated to range from 31 to 64kgpt of ore, or 0.5 to 0.67kg/kg of contained copper.

The sixteen batch test leach residues were settled using flocculent then subjected to vacuum filtration and two-stages of water washing. Filtration rates were good and ranged from 70 to 1,200kg/m²/hr and final moisture contents ranging from 6% to 12%, depending on ore type. It was also determined that increasing the wash water temperatures to 80ºC decreases the residue moisture levels by 1 to 2%.

The earlier testwork indicated that leach reactor power demands would be high in order to properly suspend the ore in the leach slurry. In the latest testwork, seven agitation tests were performed to determine optimum agitator design and to obtain a better design value for power consumption. Agitation tests were performed in a 50L plexi-glass reactor vessel using ground ore at a slurry density of 500g/L, the maximum for design purposes. Results indicate that the optimum agitator has 8-bladed upper and lower blades, at different pitches, gave the best mixing at the lowest power consumption.

Four agitator wear tests were also performed. Four different agitators made of different materials were subjected to 15 days of agitation in 10L vessels operating at 500g/L slurry density. The optimum (good wear characteristic at the most advantageous price) material, LDX2101 (OTG confidential) was selected for the larger scale test.

The large-scale agitator test was operated in a 5m³ vessel and equipped with the recommended dual 8-blade design agitator. The material, however, was not ground to design fineness but was


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only finely crushed to minus 6 mesh due to time constraints. Regardless, the test was run for 22 days and the power consumption and wear characteristics of the agitator was determined. The power consumption was obtained from the test and was factored to estimate the power required for the envisioned 500m³tanks. The current agitator power estimate will be about twice that of the initial estimates and at 0.9kW/m ³. The wear rates were also determined and will be provided in the plant design criteria for estimation of agitator replacement.

Testing also included further evaluation of the reactor oxygen consumption, wear tests of agitators, power estimation for industrial-scale agitators, settling tests of residues, and filtering and washing tests of residues. The results of the metallurgical testing will be used to further refine the copper recovery estimates, provide additional process design criteria for the leach reactor oxygen requirement and agitators and better determine the filtration requirements for the residues.

Minimal process flowsheet changes will be made based upon the additional metallurgical testing results. The power requirement increase for the leach reactor agitators, if confirmed during basic engineering, will increase process plant operating costs by several percentage points while the filtering and washing tests will result in changes to the numbers and types of filters used for residue and sludges.

14.8 Copper Recovery Predictions

The overall life-of-mine copper recovery is predicted to be 91.4%, based on the current mine production plan. Batch leach test results of nine variability samples (4 HCH, 4 HCL, 1 C4) were used to develop relationships between copper recoveries and copper grades by ore type.

The following copper recovery relationships were developed from the test results of the earlier OTG metallurgical testing for each ore type and were used to estimate the copper recoveries for the project:

HCH: Cu recovery = (0.0027 * Cu Grade) + 0.8873
HCL: Cu recovery = (0.0027 * Cu Grade) + 0.9101
C4: Cu recovery = (0.0027 * Cu Grade) + 0.9241

Copper losses in the plant operation were also estimated and considered solution losses contained in leach residue and precipitate filter cakes. An overall copper loss adjustment factor due to these losses, for all ore types, is estimated at 0.70% .

Due to expected start up problems, a copper loss adjustment factor of 1.7% is also estimated to occur during the first year of operation.

Figure 14-3 presents the metallurgical recovery graphs by ore type and shows the predicted recoveries versus ore grades based on the results of the metallurgical tests performed for the 2003 Feasibility Study.

A combination of the earlier results with the more recent test results has resulted in the following copper recovery relationships:

HCH: Cu recovery = (0.0017 * Cu Grade) + 0.893
HCL: Cu recovery = (0.0027 * Cu Grade) + 0.9101


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C4: Cu recovery = (0.0037 * Cu Grade) + 0.8988

The latest predictions indicate no difference in copper recovery for the main HC ore types while that for the C4 ore is predicted to decrease by 2%. The impact on the overall predicted copper recovery for the project is negligible, thus the 91.4 recovery estimate can be used for the project.

14.9 Copper Production Estimate

A first-year production ramp up factor of 71.7% cathode production capacity has been estimated for the Project based upon predicted losses in plant availability and lower than design plant efficiencies during that time period. Cathode quality is also anticipated to ramp up to the predicted 95% of LME Grade A quality by the fourth month of operation.

The first-year ramp up factor, the loss factors and the recovery prediction relationships discussed above were all used to develop the copper production schedule for the life of the Project. Approximately 16Mt at a grade of 6.6% copper will be processed over the 15 year project life. About 91.4% of the copper will be recovered to produce approximately 966,000t of cathode copper, or an average of 66,000tpy The estimated life-of-mine copper production schedule is presented in Table 14.9.1.

Table 14.9.1: Las Cruces Copper Production Schedule

Item	Units	Year 1	Year 2	Year 3	Year 4	Year 5	Year 6	Year 7	Year 8
Ore Processed	kt	618	1,043	1,065	758	1,105	1,075	953	1,138
Mill Feed Assay	% Cu	8.597	6.967	6.819	9.467	6.552	6.683	7.477	6.34
Cathode Copper Produced	t	47,354	65,934	66,053	65,952	66,055	66,129	66,087	65,996
	Units	Year 9	Year 10	Year 11	Year 12	Year 13	Year 15	Year 15	Totals
Ore Processed	kt	1,277	933	1,094	1,296	1,168	1,138	1,311	15,972
Mill Feed Assay	% Cu	5.694	7.718	6.608	5.544	6.137	6.403	4.997	6.616
Cathode Copper Produced	t	66,060	66,060	66,043	66,037	66,053	66,072	60,138	966,055

The current estimated production is 72,000tpy with a 15 year LoM total of 996,000t. (www.inmet-mining.com, 2008)

14.10 PAH Comments

The OTG metallurgical testwork was performed in a thorough and professional manner. PAH [believed] that the atmospheric-leaching technology selected for the project, though unique in the copper industry, will be successful;
PAH [believed] that the 5t of drill core intervals that made up the Master Composite sample, along with numerous variability samples, were sufficient to adequately project the metallurgical response of the Las Cruces ore deposit to the selected process technology without the collection of a “bulk sample”;
Additional batch metallurgical testing was preformed after the 2003 Feasibility Study that better determined the metallurgical response of C4 ore and confirms the annual copper recoveries from the main ore types based on the current mine plan;


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Additional OTG testing also finalized the leach reactor agitator designs and determined reactor oxygen requirements;
The copper recovery predictions are acceptable and consider the test results, and adjustment factors for plant losses during ramp up and equipment operation;
The copper production schedule is acceptable and considers the recovery predictions by ore type;
Design criteria for the process facility has been properly extracted from laboratory test results; and
In PAH’s opinion, OTG’s operation of 20 “batch-scale” tests and two “continuous mini- pilotscale” tests I sufficient to design the process facilities and predict copper recoveries. Copper recoveries for batch and continuous tests were comparable.

Data regarding mineral processing and metallurgical testing are unavailable to SRK, therefore a statement of adequacy and recommendations are not made.


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Figure 14-1: Metallurgical Sampling Composites Planview

Source: PAH Report, May 27, 2007


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Figure 14-2: Metallurgical Sampling Composites Section

Source: PAH Report, May 27, 2007


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Figure 14-3: Copper Extraction vs. Copper Grade

Source: PAH Report, May 27, 2007


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15 Mineral Resources and Ore Reserve Estimates (Item 19)

This section for the mineral resources and reserves for the Las Cruces project are separated into the resource discussion, followed by the mine design and reserve discussion. PAH [noted] that subsequent to feasibility resource modeling and reserve estimation work, additional infill drilling was conducted by CLC from late 2004 to early 2005, with the objective of providing more detailed information for short-term mine planning. At the time of this report, however, the results were not complete and a comprehensive review will be conducted in the future when all the results are available. Otherwise the sample data analysis and resource modeling results presented in this report do not reflect this new information.

DMT-MC created a revised resource model for the Las Cruces deposit in 2003, largely to incorporate changes in what data was included or excluded as a result of sample recovery issues. The modeling approach otherwise was very similar to the previous 2000 model created by IMC. The resource model was set up to reasonably cover the deposit area using Maptek’s Vulcan Software package. Block size was set up to be 10 x 10m in plan and a 5m bench height, which locally was subcelled to 5 x 5m in plan and a 2.5 bench height. Figure 15-1 shows the resource model limits along with the project area topography and drillholes.

15.1.1 Drillhole Database

The drillhole data for the project is stored in an Access database that contains all CR series holes used for resource evaluation. Assay files include grade values for copper, zinc, lead, gold, and silver, as well as other trace metals. The majority of the grade data is from PQ sized core for which the sample consisted of a quarter of the core (3466 samples or 79% of all the secondary sulfide core). Analytical data from the laboratory was provided in electronic format and merged directly into the drillhole database. Where analytical values were missing, either due to non-assay or sample loss, they were assigned a 0 value in the databases. For compositing and other processing of data, the 0 values are treated as non-existing. Assays below detection limits were assigned a value of one half of that limit in the database. Geological and geotechnical logging data were entered into the database through a data entry form.

The lithology code in the database is a two-part identifier that indicates the group of rocks (i.e. sulfide, gossan, volcanics, etc.) and the principle rock type (massive sulfide, semi-massive sulfide, disseminated sulfide, intense hematite, etc.). The lens code indicates the mineralized grouping (HCH - high copper/high density with secondary sulfides, HCL – high copper/low density with secondary sulfide, HC4 – high copper western zone, AU – gold enriched gossan, C1 – chalcopyrite dominant primary sulfide). The drillhole data file provided contains 18,093 samples.

A total of 4,407 samples (24% of the sample data) are from the secondary sulfide mineralization. Of the secondary sulfide samples, 57% are from the HCL zone, 36% are from the HCH zone, 5% are from the HC4 zone, and 2% are from the HCF zone.


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Although not considered as an economic zone in the current feasibility, the AU zone consists of oxidized gossan with secondary enriched gold and silver grades. A total of 704 samples (4% of the sample data) are from the precious metal enriched gossan.

The original massive sulfide deposit consists of primary sulfides, dominated by pyrite and chalcopyrite, and is considered waste material in the current feasibility. A total of 625 samples (3% of the sample data) are from the primary sulfide zone (C1 and CB lenses).

PAH [noted] that no core recovery was recorded in the drillhole database for holes CR001 to CR023, and eleven sample intervals in CR258. This includes six holes (108 samples) in the secondary sulfide zone significant for resource estimation (2.5% of the secondary sulfide sample data contained in holes CR001, CR002, CR019, CR020, and CR021, and CR258). In order not to lose these holes for compositing and subsequent resource estimation after applying a core recovery cut-off, the core recovery for these holes was set to 100%. PAH [recommended] that an actual core recovery value be estimated from core photographs or remaining core reference material for future work.

DMT-MC noted in their 2003 feasibility report that densities were not recorded for 58 secondary sulfide sample intervals. PAH found only 12 secondary sulfide intervals for which no density was recorded, all in drillhole CR233.

Database Checking

IMC conducted a validation of drillhole database integrity by randomly selecting 15 holes (about 5% of the data) by comparing original assay certificates with the database grade entries for copper, gold, lead, zinc, silver, and sulfur. IMC found that the error rate was extremely low. The few discrepancies found were minor in nature and none were related to copper. Physical examination of 10 of these holes also found that the geologic logging and analytical values appeared consistent with the core. Original films for the downhole surveys of these holes also found no problems. As a result, IMC concluded that the database integrity was good and was adequate for feasibility study work.

IMC also conducted a validation of drillhole database density values. The error rate for the initial check of select holes was found to be high enough to warrant a complete check of all the available density data. They verified the density calculations from the basic weight data and found that 9% of the density data could no longer be verified and that 1% of the density data was incorrect. The erroneous data was discarded to produce clean density data. No independent density measurements were taken as part of this work as the remaining core was found to be too degraded to produce meaningful results.

PAH previously (October 2001) searched the database with an automated checker that seeks out data inconsistencies and/or anomalous values. Based on this examination, PAH found no significant errors that would adversely impact the quality of the resource estimate.

DMT-MC for the current feasibility (2003) checked 20 holes from the original down-hole photographs and found 9 errors out of 145 readings (6% error rate). Previous review of down-hole surveys for 15 holes by IMC found no discrepancies with the database. DMT-MC further identified that the magnetic declination for true north was incorrect. The original 6º declination was corrected to 4º.


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Drillhole Core Recovery Cut-off

Samples from drilling of the Las Cruces deposit was subject to considerable variability in core recovery. This was discussed in the 2001 Bechtel feasibility study and has been evaluated by CLC (July 2003) and IMC – UK (August 2003) and well summarized in the current 2003 DMT-MC feasibility study. Geologic observation during the logging of the Las Cruces core provides good evidence of the variable loss of core material due to the abrasion and erosion of friable and sooty, chalcocite-bearing mineralization. The apparent preferential loss of this material during drilling and sampling has tended to decrease the relative content of copper minerals, with a proportionally less decrease in the relative content of the other associated minerals, including pyrite, quartz, sphalerite, and/or galena. As a result of this, the copper grade in the actual analyzed sample of the recovered core tends to be less than in the copper grade in the original in-place location.

PAH [found] that the majority of the geologic and statistical data provide evidence for the preferential loss of chalcocite, although reduction in copper grade is not a uniform systematic statistical relationship, but is somewhat irregular due to chalcocite distribution both as more durable masses and elsewhere as friable aggregates. It appears that both high core recovery and low core recovery zones have been subject to variable chalcocite loss. As such, core recovery itself may provide a good general indication that original in-place copper grades are higher than reported, but are not an absolute indication on a sample-by-sample basis that the original in-place copper grade was higher. The effect of the preferential loss of chalcocite appears to have tended to increase the difference between original in-place copper grade and the actual assayed copper grade (after core recovery affects) at progressively lower core recoveries.

This is significant in terms of generating a reasonable resource model, resulting in the question of what sample core recoveries result in reliable enough copper grades. If too many of the lower core recovery samples that potentially contain the lower copper grades are used in the resource modeling, the resulting effect is to potentially understate the copper grade of the deposit. If too few of the lower core recovery samples that potentially contain the lower copper grades are used in the resource modeling, the resulting effect is to have a reduced confidence in the grade estimation because too few samples were used in the estimation. This is shown graphically in Figure 15-2.

For resource modeling purposes, a limit on the core recovery percentage of each sample is required, to determine which samples should be included (relatively reliable copper grades) and which samples should be excluded (relatively unreliable copper grades) for the subsequent resource modeling. This core recovery cut-off was been applied in the modeling process. In original Rio Tinto feasibility work a 50% core recovery cut-off was applied, while in the subsequent Bechtel Feasibility work (IMC modeling), no core recovery cut-off was applied. For the current work by CLC, the reliability of the sample grades at those recovery levels were questioned and investigated further. As a result, the percent core recovery cut-off grade was increased to an 80% core recovery. This clearly gives more reliability for the copper grades to those samples that meet this higher core recovery cut-off grade criteria.

As a result, however, there are fewer actual copper grade values that are available for use in the modeling process, which then implicitly relies on whatever adjacent, non-excluded copper grades are available for the copper grade modeling process. In other words, the more sample copper grades are excluded by the core recovery cut-off grade, the less the actual data remaining


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for copper grade modeling. For secondary sulfide lenses, at core recovery cut-off grades above 70%, the number of available samples drops off quickly (at 50% cut-off about 90% of the samples are available, at 80% cut-off about 75% of the samples are available, at 90% cut-off about 50% of the samples are available, at 100% cut-off about 15% of the samples are available). The 80% core recovery cut-off for secondary sulfide samples includes 3,275 samples out of a total of 4,407 samples (74%).

PAH [found] that the selection of an 80% core recovery cut-off grade appears to be a reasonable compromise, whereby the core recovery cut-off is high enough that the remaining non-excluded samples have reliable grades, but that the core recovery is low enough that there are a sufficient number of the remaining non-excluded samples to effectively create a copper resource block model (discussed further in database subsection). Dr. Stephen Henley of IMC –UK (August 2003) evaluated this issue during the preparation of the current feasibility study and provided an independent opinion, concluding that the 80% core recovery cut-off grade is reasonable and still likely to lead to an underestimation of the resource grade. This reduced confidence was reflected in the DMT-MC resource estimate by disallowing any of the resource that has been heavily influenced by neighboring samples substituting for missing samples from being classified as measured, as discussed in the subsequent report subsection.

Sample Statistics

Sample copper statistics are summarized in Table 15.1.1.1. The table compares sample copper statistics for all secondary sulfide samples (0% core recovery cut-off) and for those secondary sulfide samples at an 80% core recovery cut-off. A volume weighted mean has been calculated to compensate for different sample lengths and densities. DMT-MC noted some 58 secondary sulfide samples did not have density determinations so these were calculated by regression relationships based on the iron and sulfur content of these samples. PAH [noted] within the data provided for checking, hole number CR233 has 12 secondary sulfide samples for which a density value is not present in the database. It can be seen that when the 80% core recovery cut-off is applied to the secondary sulfide samples the mean copper grade is approximately 10% higher grade than when no core recovery cut-off is used. PAH [noted]that secondary sulfide samples below the 80% core recovery cut-off have a mean copper grade of 5.10% copper, or 35% lower than the mean copper grade for the samples above the 80% core recovery cut-off.

Table 15.1.1.1: Las Cruces Sample Copper Data Statistics*

Zone (Lens)	Parameter	Samples (0% RECP cut-off)	Samples (80% RECP cut-off)
All Secondary Copper	Number	4,407	3,275
(HCH, HCL, HC4)	Minimum (%Cu)	0.01	0.01
	Maximum (%Cu)	39.35	39.35
	Vol. Wt. Mean (%Cu)	7.01	7.85
	Std. Dev. (%Cu)	5.71	6.00
	Coef. Of Var.	0.83	0.79
* Volume weighted mean = gradelengthdensity

Sample density statistics are summarized in Table 15.1.1.2. Density values were determined for every sample interval. PAH [noted] that CR233 had 12 sample intervals for which no density value was recorded in the drillhole database. The table compares sample density statistics for all secondary sulfide samples (0% core recovery cut-off) and for those non-excluded secondary sulfide samples at an 80% core recovery cut-off. As with the copper data, a volume weighted


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mean has been calculated to compensate for different sample lengths and densities. It can be seen that for the density data there is little difference when the 80% core recovery cut-off is applied to the secondary sulfide samples.

Table 15.1.1.2: Las Cruces Sample Density Data Statistics*

Zone (Lens)	Parameter	Samples (0% RECP cut-off)	Samples (80% RECP cut-off)
All Secondary Copper	Number	4,407	3,275
(HCH, HCL, HC4)	Minimum (g/cm³)	1.86	2.00
	Maximum (g/cm³)	5.04	5.04
	Len. Wt. Mean (g/cm³)	3.66	3.71
	Std. Dev. (g/cm³)	0.65	0.63
	Coef. Of Var.	0.18	0.17
Length weighted mean = length density.

15.1.2 Compositing

Composite Calculation

Drillhole sample data was composited into regular geological composites at a target 5m length. Basically, within a given geologic interval the composites were split into regular intervals as near as possible to the target length. PAH [noted] that this forces composites not to extend across a geologic interval, but results in some variability in composite lengths, especially for narrower geologic intervals. Additional dilution was subsequently incorporated into the reserve estimate to compensate for mining effects at the edges of the secondary sulfide mineralization (see mining section). The composite grades were generated by weighting the individual samples by assay length and by density. This provides a volume weighted mean that is more appropriate with variable density materials. All secondary sulfide composites were considered together, as any separation by lens for geostatistical purposes was considered to be impractical.

If any sample data contained within a composite interval was excluded because it was below the 80% core recovery cut-off, then the 5m composite value was calculated using only the remaining sample data with above 80% core recovery. There was no minimum sample length required for the composite to be calculated. However, only composites for which more than 50% of the length was “supported” by non-excluded sample data were used for resource estimation.

Composite Statistics

As a result of the application of an 80% core recovery cut-off to the samples and the subsequent use of a support cut-off of 50%, a total of 650 non-excluded composites were calculated, 71% of the total number without applying cut-offs. Acceptable composites were available for 134 holes out of the 153 holes that intercepted the secondary sulfide zone. Nineteen holes were entirely excluded because they contained no acceptable secondary sulfide composites. In addition, PAH [found] that there are another 23 holes for which the application of an 80% core recovery cut-off and a 50% support cut-off, results in less than half of the secondary sulfide composites for those holes being accepted. In addition, there were five drillholes (plus part of a sixth) for which a 100% recovery was assigned because no core recovery value was available, otherwise these holes may have experienced a high exclusion rate as well as they were among some of the first holes drilled. PAH evaluated the distribution of these holes and found them to be scattered throughout the deposit area. This is a better situation for resource modeling than if they were


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clustered together. A comparison of sample and composite copper statistics are shown in Table 15.1.2.1. PAH [found] this to be a reasonable comparison.

PAH evaluated statistical outliers in the 5m composites based on data population distributions and found some statistical evidence for special treatment of copper grades above 25%. DMT-MC did not apply any special treatment. PAH [noted], however, that this would have affected only four composites and as such would have a minimal effect on the global resource.

Table 15.1.2.1: Las Cruces Composite Copper Data Statistics*

			5m Composites
Zone (Lens)	Parameter	Samples (80% RECP cut-off)	(50% Support cut-off)
All Secondary Copper	Number	3,275	650
(HCH, HCL, HC4)	Minimum (%Cu)	0.01	0.18
	Maximum (%Cu)	39.35	27.88
	Mean (%Cu)	7.85	7.67
	Std. Dev. (%Cu)	6.00	5.07
	Coef. Of Var.	0.79	0.66

*For samples - volume weighted mean = grade * length * density. For composites – length weighted mean = grade * length.

A comparison of sample and composite density statistics are shown in Table 15.1.2.2 and PAH [found] a reasonable comparison. PAH [did] not find any evidence for special treatment of any of the higher density values.

Table 15.1.2.2: Las Cruces Composite Density Data Statistics*

			5m Composites
Zone (Lens)	Parameter	Samples (80% RECP cut-off)	(50% Support cut-off)
All Secondary Copper	Number	3,275	650
(HCH, HCL, HC4)	Minimum (g/cm³)	2.00	2.25
	Maximum (g/cm³)	5.04	4.80
	Len. Wt. Mean (g/cm³)	3.71	3.71
	Std. Dev. (g/cm³)	0.63	0.56
	Coef. Of Var.	0.17	0.15
* Note: Length weighted mean = length * density.

Composite Variography

DMT-MC conducted a geostatistical evaluation, with variography based on composites generated in holes at an approximate 50m spacing. Two “statistical crosses” of more closely spaced (12.5m) inclined (60º) were drilled to support the geostatistical evaluation, by providing sample pairs at a closer spacing to better define the shorter range sample variability. One cross was placed in the western part of the deposit, while the other was in the central part of the deposit and was not completed.

DMT-MC found that for the copper grade the longest effective range was 220m at a 20º azimuth. The shortest effective range was 110m at a perpendicular azimuth of 110º. These directions were consistent with the geologic observations for the deposit. The independent variance (nugget) was found to be approximately 40% of the total variance, which PAH noted is indicative of irregular mineralization.


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It is PAH’s experience that primary massive sulfide mineralization typically is deposited with a good degree of continuity such that relatively high distances or ranges of influence could be expected (up to 100m or more) for primary chalcopyrite massive sulfide mineralization. Subsequent secondary enrichment with chalcocite mineralization, however, is a very irregular process, dependent on leach solution grades and primary sulfide porosity and permeability and disrupts the original statistical continuity. Twin drilling corroborates this, in that intervals even at very short distances do not compare well.

PAH conducted a check of variography and found that variography for the mineralization using a 0% copper cut-off (all composites) found results similar to those of CLC, with maximum ranges of 200m and a moderate nugget variance. Using copper grade indicators, PAH found that copper grades at a 5% cut-off increasingly showed the influence of the more irregular secondary chalcocite mineralization, with the maximum variogram range dropping to around 100m and an increase in the nugget variance. At a 7.5% copper grade indicator the maximum variogram range is around 60m, with a further increase in the nugget variance. The nugget effect is relatively high for a copper deposit and provides a good statistical measure of the irregularity of the copper in higher-grade parts of the deposit, one that can be corroborated by geologic observation of the drillhole cross sections.

15.1.3 Rock Model

For the current feasibility study the Las Cruces deposit was completely reinterpreted, focusing on the secondary sulfide zones. Digital surfaces were created to delineate the three main secondary sulfide sub-lenses, including the HCH, which included the minor HCI lens, the HCL which included the minor MCL and HCF lenses, and the HC4. The primary sulfide and gossan zones were also delineated, as well as divisions of the Paleozoic host rocks and overlying Miocene sedimentary cover.

15.1.4 Grade Models

A copper model was created by DMT-MC using an ordinary kriging (OK) interpolation method. As the boundaries between the secondary sulfide lenses are complex and gradational, the HCH/HCL sub-lenses were grouped together for resource estimation. Interpolation was conducted using a search ellipsoid with the primary axis having a 200m radius oriented at a 20º azimuth as indicated by variography and geologic observation. The secondary axis had a 90m radius oriented at a 110º azimuth. The tertiary axis had a 10m axis oriented vertically. A minimum of two composites and maximum of 30 composites was specified for the interpolation. PAH [believed] that 30 composites probably result in more smoothing of grades during the interpolation process, however, this would be expected to have only a small affect on the global resource average grade.

A density model was created for all geologic units and mineralized lenses, with the exception of the Miocene marl and sandstone units, using an inverse distance squared interpolation method. Interpolation was conducted using a search ellipsoid with the primary axis having a 400m radius oriented at a 20º azimuth as indicated by variography and geologic observation. The secondary axis had a 300m radius oriented at a 110º azimuth. The tertiary axis had a 20m axis oriented vertically. A minimum of two composites and maximum of 12 composites was specified for the interpolation.


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Primary sulfides have been considered as a single domain for the purposes of resource modeling, but are not part of the economic evaluation of the current feasibility.

A comparison of composite and block model copper statistics are shown in Table 15.1.4.1 and PAH [found] a reasonable comparison. The slightly lower average copper grade for the block model is the result of declustering affects during the interpolation process and is to be expected by kriging.

Table 15.1.4.1: Las Cruces Block Model Copper Data Statistics*

Zone (Lens)	Parameter	5m Composites (50% Support cut-off)	Block Model
All Secondary Copper	Number	650	14,246
(HCH, HCL, HC4)	Minimum (%Cu)	0.18	1.41
	Maximum (%Cu)	27.88	18.16
	Mean (%Cu)	7.67	6.62
	Std. Dev. (%Cu)	5.07	2.95
	Coef. Of Var.	0.66	0.45
For composites – length weighted mean = grade length. For block model – arithmetic mean.

A comparison of composite and block model density statistics are shown in Table 15.1.4.2 and PAH [found] a reasonable comparison.

Table 15.1.4.2: Las Cruces Block Model Density Data Statistics*

Zone (Lens)	Parameter	Samples (80% RECP cut-off)	Block Model
All Secondary Copper	Number	3,275	14,246
(HCH, HCL, HC4)	Minimum (g/cm³)	2.00	2.77
	Maximum (g/cm³)	5.04	4.69
	Vol. Wt. Mean (g/cm³)	3.71	3.67
	Std. Dev. (g/cm³)	0.63	0.45
	Coef. Of Var.	0.17	0.12
For composites – length weighted mean = grade length. For block model – arithmetic mean.

DMT-MC conducted a number of checks of the reasonableness of the block model. DMT-MC concluded from statistical integrity checks that the block copper grade and core density had been adequately accounted for, with no inexplicable changes. DMT-MC noted that Rio Tinto had previously investigated the effects of non-Gausian data transformations by comparing their ordinary kriged model with a multiple indicator kriged model and found that they gave an almost identical mean copper grade and concluded that the ordinary kriging reasonably interpolated the copper grade. DMT-MC conducted a point validation check and found a slight bias that may tend to slightly overestimate block copper grades on a local basis; however, they concluded this was negligible on a global basis in the model. A point validation for core density showed no evidence of bias. Declustering investigations by DMT-MC found that the overall effects on the grade interpolation from the drillholes clustered in the statistical crosses was very small. Investigation of stationary effects (mean and variance constant over distance) by DMT-MC concluded that nonstationarity conditions do not have a major impact on the validity of the block model estimate. DMT-MC also visually compared composite data with block model data for sections and plans through the resource model and found a reasonable comparison.


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PAH [noted] that a reasonable amount of checking appeared to have been conducted by DMT-MC. PAH conducted some further independent checks of the DMT-MC resource model. PAH evaluated the statistics for each step of the modeling process and visually compared composite and block model data, checked the local estimation integrity of the copper grade model for grade bias by comparing individual composite values against the block value in which the composite occurred (backmarking) and by conducting a cross validation. PAH then checked the global integrity of the copper grade model for grade bias by creating a nearest neighbor model (computer polygonal model) and then compared the results against the PAH inverse distance model. As a result, PAH [believed] that the resource model, due to natural variability in the deposit and due to the exclusion of data with core recoveries of less than 80%, may be a modest estimator of grade locally, but believes that on a global basis the estimation is reasonable. In addition, CLC has noted that some faulting may be locally present in the deposit, but is not well defined at this stage of exploration, and may locally influence copper grade distribution.

15.1.5 Resource Statement

The Las Cruces mineral resource was estimated from the grade and rock models, developed as discussed above. The resource includes all secondary sulfide material in the model at the given cut-off grade. The resource tonnage uses modeled core densities for each block, which are then adjusted in the resource tabulation to account for permeable pore space. Table 15.1.5.1 shows that at a 1.0% copper cut-off, the measured + indicated resource is 15.6Mt of HCH, HCL and HC4 ore types, averaging 6.89% copper. There is an additional inferred resource of 0.4Mt averaging 8.66% copper. Incidental primary sulfide (CZ) and gossan (AU) that occurs in the planned pit will be stockpiled separately and is not considered part of this resource estimate.

Table 15.1.5.1: Las Cruces Mineral Resource Summary*
	Measured		Indicated		Measured + Indicated		Inferred
Ore Type	t (000s)	Grade (% Cu)	t (000s)	Grade (% Cu)	t (000s)	Grade (% Cu)	t (000s)	Grade (% Cu)
HCH	3,860	7.25	2,670	8.00	6,530	7.56	274	9.48
HCL	5,020	6.56	3,370	5.77	8,390	6.24	37	4.86
HC4	440	8.61	230	8.55	670	8.59	49	6.93
Total	9,320	6.94	6,270	6.82	15,590	6.89	360	8.66
*1.00% Cu Cut-off, 0.97 density adjustment.

PAH checked the resource estimation from the DMT-MC model by independently tabulating that and grade from PAH’s copy of the DMT-MC model and found a good comparison. PAH [noted] that within the secondary sulfide resource there are very few blocks with less than a 2% copper grade, so the resource tabulation at cut-off grades up to 2% copper is very similar.

15.1.6 Resource Classification

DMT-MC has classified the resource in accordance with the Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (JORC Code). Under the JORC Code, only measured and indicated resources may be converted to mineable reserves with the application of appropriate economic factors. PAH [believed] that the classification is in compliance with the Canadian Institute of Mining (CIM) standards required for National Instrument 43-101 reporting. PAH [believed] that the classification is consistent with the


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development of United States Securities Exchange Commission (SEC) acceptable mineable reserves.

DMT-MC has assigned confidence categories to the resource based on a combination of distance and number of drillholes. The strategy used is shown in Table 15.1.6.1. For the HCH and HCL zones, the search ellipsoid was horizontally oriented with the major axis oriented at an azimuth of 20º. For the minor HC4 zone, the search ellipsoid was dipped into the dipping orientation of the zone, using the same classification criteria. In addition, the measured resource confidence category was locally adjusted to reflect the decreased estimation accuracy resulting from the exclusion of about 30% of the total drillhole data (from the application of sample and composite quality cut-offs). For each measured block, if the ratio of excluded:non-excluded composites, within a 25m search radius, was greater than 0.5, then it was reassigned to indicated. This requires that more than two thirds of the nearby composites be non-excluded composites in order to be considered as measured. This additional criterion reclassified 4.7Mt averaging 6.9% copper of measured resource into an indicated category.

Table 15.1.6.1: Las Cruces Resource Confidence Classification Criteria


	Measured	Indicated	Inferred
Search Distance	100 x 45 x 10	200 x 90 x 10	200 x 90 x 10
Min. No. Of Composites Min.	3	3	2
No. Of Drillholes Excluded	3	3	1
Sample Ratio	<0.5	-	-

It is PAH’s experience that primary massive sulfide mineralization typically is deposited with a good degree of continuity such that relatively high distances or ranges of influence could be expected (up to 100m or more) for primary chalcopyrite massive sulfide mineralization. Subsequent secondary enrichment with chalcocite mineralization, however, was a very irregular process, dependent on leach solution grades and primary sulfide porosity and permeability and has disrupted the original statistical continuity. Twin drilling corroborates this in that intervals even at very short distances do not compare well, as does observations from cross sections.

PAH conducted a check of variography to corroborate the search distances and found that using a 0% copper cut-off (all composites) the results were similar to those of CLC, with maximum ranges of 200m and a moderate nugget variance. Using copper grade indicators at progressively higher cut-off grades, PAH found that the variogram range decrease to about 60m at a 7.5% copper grade indicator. The nugget effect at this indicator is relatively high for a copper deposit and provides a good statistical measure of the irregularity of the higher grade copper values in the deposit, which can be corroborated by geologic observation of the drillhole cross sections.

As a result, PAH [believed] that shorter ranges of influence may be appropriate, but because of the nominal 50 x 50m drillhole spacing, actual distances from composite points do not reach the maximum limit of the search ellipsoid. For measured blocks the maximum distance to the nearest composite was 80m, with 99% of the measured blocks within 60m of a composite, and the majority of the measured blocks being within 10 to 40m of a composite. The distribution of indicated blocks shows a similar distance distribution, with the maximum distance to the nearest composite of 90m, with 95% of the indicated blocks with 60m of a composite, and the majority of the blocks being within 10 to 50m of a composite. CLC finds that much of the measured and indicated blocks with higher distances to the nearest composite is a result of deposit and drillhole


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geometries, combined with a narrow vertical search, that prohibit closer composites from being identified.

15.1.7 Previous Resource Estimates

Previous resource estimates were prepared for Las Cruces and were compared by PAH to the current estimate, as shown in Table 15.1.7.1. These historical estimates have not been reviewed in detail by PAH and are superseded by the current feasibility Study resource estimate. No additional resource drilling has occurred since 1999. The last two estimates, DMT-MC in 2003 and IMC (Tucson) 2000, were both conducted for CLC. The current DMT-MC estimate has incorporated fewer, but higher quality composites (>80% core recovery) for resource estimation, which is the main resource model change since the previous estimate. DMT-MC classifies the resource as measured (59%), indicated (39%), and inferred (2%). In 2000, IMC (Tucson) prepared a resource model for use in the Bechtel feasibility study (2001) that planned for open pit mining. IMC (Tucson) considered the entire resource to be at a combined measured + indicated (100%), but did not separate the resource into individual categories. In 1998, Rio Tinto prepared a resource estimate for their internal feasibility study that envisioned underground mining, prior to transferring the property to CLC. Rio Tinto classified the resource as measured (58%), indicated (36%), and inferred (6%).

Table 15.1.7.1: Las Cruces Comparison of Historical Resource Estimates*

Estimate	%Copper Cut-off	Density Adjustment	Mineral Zones Included	Confidence Categories	Tonnes	Grade (%Cu)	Contained Tonnes Cu

DMT-MC 2003 IMC 2000 Rio Tinto 1998	1.0 1.0 1.0	0.97 0.95 0.97	HCH,HCL,HC4 HCH,HCL,HC4 HCH,HCL,C4	Meas+Ind+Inf Meas+Ind Meas+Ind+Inf	15,954,000 15,511,000 15,510,000	6.93 6.07 6.10	1,105,600 941,1000 946,100

*IMC classified the entire resource as measured + indicated, with no inferred material.

It can be seen that the three feasibility study resource estimates are reasonably comparable in tonnes. The current DMT-MC estimate, however, is about 10% higher in copper grade than the previous estimates due to the application sample and composite quality cut-offs, that tend to include higher grade samples and exclude lower grade samples, as discussed previously in this section.

SRK did not generate the Mineral Resource estimates and was unable to conduct an audit as prescribed by NI 43-101. However, the estimations and conversions are reported as CIM compliant.

15.2 Additional 2004/2005 Drilling Results

PAH [concurred] with CLC plan to drill about 40 to 50 more drillholes over the life of the mine, in order to better define local grade variability prior to actual mining. All future drilling should be conducted using large diameter drillholes (PQ – 3.345in or 85mm in diameter), with specialized core recovery mechanisms (such as triple tube core barrel), and careful drilling practices.

From late 2004 and extending into April 2005, CLC drilled an additional 14 holes (CR285 – CR298) into the western high-grade area of the deposit. This area, around the western statistical cross, will be the first part of the deposit to be mined. Feasibility Study modeling entirely excluded eight holes and partially excluded nine holes in this area. At the time of this report, the results for five of the new holes were available for evaluation, with the remainder still


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outstanding. Interim reports on the incomplete data was reviewed by PAH, although the data itself was not made available. PAH [assumed] that the new holes were drilled in a similar manner as the previous holes.

The interim drilling reports indicated that the infill drillholes, which presumably had reasonably good core recovery, intercepted generally similar mineralization as the DMT-MC Feasibility Study model had estimated. A structural complication, however, was identified that consisted of a high angle-fault trending north-south through the western part of the deposit. The main impact of this fault was the introduction of a thin wedge of low-grade shale in place of originally modeled high-grade HCH secondary enriched mineralization. Preliminary evaluation by CLC in March 2005, indicated that although this is significant locally, on a global basis the resource was negatively impacted by only around 1% of total contained copper metal. PAH [concurred] that the global affect of this structure is very small and not a serious issue to the project.

Likely, other areas of the deposit may similarly be affected by local structural complications, however, PAH would expect that these would have local significance for short term mine planning, but would not have a serious impact on global resources. PAH [anticipated] that all of the results will be available over the next couple of months, but from the preliminary data to date has not identified any serious issues.

15.3 Additional Exploration Potential

Additional exploration potential exists for the Las Cruces deposit in the secondary sulfide zone (the focus of this review), the primary sulfide zone (at depth), and the gossan zone (enriched precious metals overlying the secondary mineralization). In the secondary sulfide zone, potential exists for some upgrade to the copper grade based on the conclusions that copper tended to be preferentially lost during the drilling and sampling. Gold, silver, and minor base metal content also occurs in the mineralization, but will not be recovered through the planned processing methods. In addition to the copper grade of 7.0% copper (up to 39.4% Cu), the secondary sulfide also contains grades averaging about 0.2% zinc (up to 36.9% Zn), 0.6% lead (up to 40.0% Pb), 0.5gpt gold (up to 62.9gpt Au), and 26gpt silver (up to 1,472gpt Ag), based on averages for the drillhole data.

Below the Miocene unconformity, the uppermost part of the sulfide mineralization has been completely oxidized to an iron-oxide gossan, with local enrichment of gold and silver. The gossan is also enriched in lead, as the mineral galena. PAH’s review found that the gossan is generally 10 to 20m in vertical thickness, with grades from all gossan drillhole samples averaging 5.9gpt gold (up to 353gpt Au) and 98gpt silver (up to 1,733gpt Ag). Average copper grade in the gossan is 0.20% . Apparently associated with the gossan formation was the strong silicification of the host rock in the hangingwall immediately above the massive sulfide deposit, also containing erratic gold and silver enrichment. Preliminary historic estimates by Rio Tinto at a 1.0gpt gold cut-off indicate a mineralized material of 1.7Mt averaging 4.25gpt gold and 118gpt silver, however IMC found about the same tonnage, but with 2.52gpt gold and 136gpt silver. There is some evidence that the gold content has been under reported. None of the gossan mineralization is included in the current resource estimate.

Original primary massive sulfide mineralization contain massive to semi-massive sulfide minerals (generally more than 80% sulfide). Pyrite is the predominate sulfide mineral, with lesser finely intergrown sphalerite, galena, and chalcopyrite, as well as minor enargite,


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tennantite, and tetrahedrite. PAH’s review found that grades from all primary sulfide drillhole samples averaged 3.2% copper (up to 18.9% Cu), 1.1% zinc (up to 11.9% Zn), 0.4% lead (up to 5.4% Pb), 0.5gpt gold (up to 1.63gpt Au), and 20gpt silver (up to 174gpt Ag). Preliminary historic estimates by IMC previously estimated the primary sulfide mineralized material to be 17.9Mt averaging 1.85% copper, 2.97% zinc, 0.61% lead, 0.27gpt gold, and 23.7gpt silver, at a 1.0% copper cutoff. Most of the primary massive sulfide mineralization occurs down-dip to the northwest of the secondary sulfides. In the footwall below the primary massive sulfide is a stockwork of interconnected pyrite veins and veinlets, with local higher grades of copper and zinc. None of the primary sulfide mineralization is included in the current resource estimate.

The western part of the Las Cruces deposit has been cut off by a fault. Potential exists for finding the missing offset piece and some exploration drilling has been conducted. The Las Cruces deposit is a blind deposit, initially located by geophysical techniques, and subsequently drilled. Other anomalies are known in the area, and other historical producers (such as Los Frailes) are located within a 10km radius; therefore, other viable exploration targets are available.

15.4 Reserve Development

The reserves for Las Cruces are a subset of the mineral resource described earlier in this section, that are contained within an engineered mine plan. Several development options were considered, including underground methods; however, surface methods resulted in the best overall recovery at the lowest cost/t.

Inmet has since modified the mine plan to include some underground mining.

15.4.1 Mine Design

The Las Cruces project is planned to be a moderate-sized open pit mining operation using conventional truck and shovel operations. The use of underground methods was considered, however, the open pit method provided a more cost effective and controllable mining approach. CLC plans to use a mining contractor to mine ore and waste, and will employ a technical group to supervise the mine contractor. After the first year of production ramp-up, the mine is expected to produce up to a relatively steady annual production of 66,000t of copper cathode, even during years of below-average feed grade, using on-site stockpiles from earlier, higher production years. Based on this processing capacity and the current reserves, the total life of the mine will be approximately 15 years.

Mine development will require a pre-production phase of 18 months for pre-stripping to expose sufficient ore to assure steady ore production. The open pit mining method will utilize drilling and blasting, loading with hydraulic excavators, and transport by trucks. The overall pit slope angle will be 28° in the upper and lower tertiary marl and sandstone, and 45° in the Paleozoic bedrock.

Within the upper parts of the overburden, the marl will be mined without drilling and blasting. Trucks will haul the ore to the primary crusher located near the processing plant and will haul the rock waste to a surface storage facility where it will be covered by the marl overburden. Approximately 43Mt of overburden will be mined during the development phase. In the later years of the project, partial backfilling of the pit with marl will occur.


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The waste material handled in the course of the mining operations will be: (1), tertiary marl and sandstone; and (2) mine rock consisting of Paleozoic shale and volcanics. There will be four main dumps for waste: North, West, South and an in-pit backfill dump.

While the Las Cruces Project offers no unusual challenges from the mining perspective, there are some attributes of the property that influence some of the design decisions:

The overburden layer, consisting primarily of marls, presents an unusually thick overburden for an open pit copper operation of this size;
A good portion of the overburden is considered to be free-digging (i.e. does not require drilling and blasting);
The average ore grades, at 6.6% copper, are unusually high; and
The total ore involved, at 16Mt, is relatively low.

The open pit mineable reserves, pit designs, and production schedule were developed by IMC of Tucson, Arizona. Equipment selection, labor staffing, and capital and operating costs were prepared for the open pit mine by DMT-Montana Consulting (DMT-MC) of Essen, Germany, based on the design work compiled by IMC. The mining plans include a 1.5 year pre-production plan and a 15-year life. The operation will commence backfilling the pit around the 9^thyear of production and will continue until mine-out, followed by two years of partial backfill. Present plans call for the remainder of the pit to be used as a dump for sterile fill.

15.4.2 Cut-off Grade

For the most part, the cut-off grade computations are derived from the original IMC estimates and assumptions. These design elements were not specified in the DMT-MC report; however, use of the IMC values produces a breakeven cut-off grade that agrees well with that used in the Feasibility Study.

Typically, a cut-off grade is estimated by the engineers based upon the best state of knowledge at the time. A “first pass” is then developed on reserves, with the development of plans and more detailed operating costs. In most cases, the resulting operating costs are reasonably comparable to the initial assumptions, and the underlying mine plans require no modification. In cases where the estimated operating costs are materially different, (10 to 15%), it is prudent to consider a redesign of the pit based upon the more accurate numbers. Note that operating cost estimates are currently based on Euros although the older design plan made use of dollar-based estimates.

Table 15.4.2.1 presents the basic economic assumptions used in the pit designs, along with a cutoff grade calculation. Note that specific refinements for haulage and pit slopes are added later.


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Table 15.4.2.1: Las Cruces Basic Cut-off Calculation


Description¹	Units	Design & Revenue Basis		Feasibility Costs		Feasibility Costs/$1.00 Cu
Description¹	Units	€	$	€	$	€	$
Mining Cost Processing Cost G&A Cost Breakeven Cost Internal Cost² Dollar/Euro Copper Price Recovery Revenue BE Cut-off³ Internal Cut-off	/t-ore /t-ore /t-ore /t-ore /t-ore /lb-Cu %Tot Cu %/t/%Cu % Cu % Cu	0.72	0.80	0.90	1.00	0.90	1.00
		22.25	24.70	23.00	25.53	23.00	25.53
				8.00	8.88	8.00	8.88
		22.97	25.50	31.90	35.41	31.90	35.41
		22.25	24.70	23.00	25.53	23.00	25.53
		1.11
		0.72	0.80	0.72	0.80	0.90	1.00
		91	91	91	91.00	91	91
		14.46	16.05	14.46	16.05	18.07	20.06
			1.59		2.21		1.76
			1.54		1.59		1.27
1. Mining costs for waste and overburden are variable as a function of depth.
2. The internal cost is the lowest cost per tonne needed to add to the contribution margin of the project (i.e. milling).
3. The value of the cut-off does not change as a function of the underlying currency

It is apparent that the Design Basis differs considerably from the final costs estimated in the Feasibility (€22.97 in the basis, versus €31.90 after detailed development, or 30% higher). While PAH [noted] that this level of difference would normally trigger a re-design of the mine plans, the difference is not material for the following reasons:

The design of the mine is relatively insensitive to cut-off grade, due to the very high average grade and structurally constrained nature of the deposit;
The mine operating costs used within the mine design are based upon first-principle cost estimates and contractor quotes, varying by rock-type and depth, rather than a simple average;
The original price assumption of $0.76/lb of copper is lower than the current market price of $1.45/lb;
The resulting design, while possibly sub-optimal as a result of the change in economics, is still feasible; and
The combined effect of higher operating costs and higher metal prices largely offset each other.

15.4.3 Pit Design

The mine production rate was established in an optimization study using an economic deposit model rather than a cut-off grade-based model. The basic underlying economic assumptions are presented in the previous Table 15.4.2.1, with the exception of mining costs, which are calculated based on the formulae provided in Table 15.4.3.1. The ultimate pit design was based on a Lerchs-Grossman algorithm that produces an approximate pit design that is mathematically optimal for the design criteria.


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Table 15.4.3.1: Las Cruces Design Basis – Mine Operating Cost Assumptions*


	Elevation				Cost
	Top	Bottom	Density	At Top	At Bottom	Halfway
Rock Type	msl	msl	dmt / m3	€ / m3	€ / m3	€ / dmt	Formula
Marl	100	15	1.96	1.314	1.314	0.670	(0.1803 € / m3+1.06395 € / m3+((32.5m-15m)/10m+1)*0.02535 € / m3)
	15	-5	1.96	1.314	1.365	0.683	(0.1803 € / m3+1.06395 € / m3+((32.5m-Block Elev.m)/10m+1)*0.02535 € / m3)
	-5	-55	2.02	1.184	1.311	0.618	(1.06395 € / m3+((32.5m-Block Elev.m)/10m+1)*0.02535 € / m3)
	-55	-125	2.02	1.619	1.797	0.845	(0.308 € / m3+1.06395 € / m3+((32.5m-15m)/10m+1)*0.02535 € / m3)
Sandstone	-100	-115	2.27	by dmt	by dmt	1.242	(0.547 € / dmt * Density dmt / m3)
	-115	-185	2.27	by dmt	by dmt	1.267	(0.558 € / dmt * Density dmt / m3)
Gossan	-125	-215	2.34	by dmt	by dmt	3.230	(1.378 € / dmt * Density dmt / m3)
Paleozoic Waste	-125	-215	2.71	3.652	3.880	1.390	(3.52545 € / m3+(1+(-95.1m-Block Elev.m)/10m+1)*0.02715 € / m3)
Paleozoic Ore	-125	-215	3.54	3.652	3.880	1.064	same as waste
*Marl and sandstone densities based on wet tonnes.

The mining costs in Table 15.4.3.1 are developed from a combination of DMT-MC calculations and contract mining vendor quotes. PAH added approximate cost calculations to provide a reference for comparison to similar operations. All cost estimates were generated in Euros.

Pit slopes have been developed by a series of consultants, with the most recent developed by Geocontrol (2003). The most recent variation is a more conservative design over the previous variations, requiring an overall slope of 28º in the upper sediments and 45º in the Paleozoic rocks. All studies reviewed by PAH noted the need for extensive dewatering systems, which are incorporated into the design and cash flow.

Table 15.4.3.2 provides a summary of the design parameters used in the development of the ultimate pit.

The pit parameters are acceptable. The upper levels of the marl appear to be soft enough for free-digging, but may see somewhat lower productivities than would be achieved with blasted muck. PAH [noted] that 30 road widths have been incorporated into current to accommodate up to 135t trucks.

The Lerchs-Grossman algorithm, while a powerful tool for determination of approximate pit geometry, does not produce a final design. The engineer must manually incorporate the roads and faces into the design, removing unstable slope structures from the design. DMT engineers developed the ultimate pit plan using the design parameters in the previous Table 15.4.3.2. The DMT-MC plan was subsequently refined by IMC, the results of which are presented in Figure 15.1.2.1


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Table 15.4.3.2: Las Cruces Pit Design Parameters


Description	Value	Units	Comments
Slope Stability Parameters
Overall Slope in Marl	28	degrees	Geocontrol 2003
Overall Slope in Other Rock Units	45	degrees	Geocontrol 2003
Bench Height in Marl & Sandstone	10	meters	Design Decision
Bench Face Angle in Marl & SS	60-62	degrees	Geocontrol 2003
Bench Height in Ore Levels	5	meters	Design Decision
Bench Face Angle in Ore Levels	75	degrees	Geocontrol 2003
Ramp Parameters
Maximum Inclination in Marl & SS	8	percent	Better ramp angle for most trucks
Maximum Inclination in Paleozoics Maximum Inclination	10	percent	Good ramp angle for most trucks
for Temporary Ramps	15	percent	Acceptable for short distances only
Minimum Width for Two-Way Traffic	25.0	meters	3.0 x truck width + 5m (23.3 for Cat 777)
Minimum Width for One-Way Traffic	15.0	meters	1.5 x truck width + 5m (14.2 for Cat 777)
Blasting Requirements
Marl (assigned to lower 25%)	25	percent	Light blasting may improve production
Ore and Waste other than Marl	100	percent	Reasonable

Mining bench height will be 10m in the marl and 5m below the marl, to allow for improved grade control selectivity and minimization of dilution in the ore zone.

The haul road width, including berms, is 30m in the main haulage roads, narrowing to 25m in lower portions of the marl and primary rock, and narrowing further to 15m within small areas at the pit bottom. These widths are well within compliance with the industry accepted standard width of three to four times the haul truck width for the 90t capacity trucks selected for waste mining. Primary haul road grades range from 7%, increasing to 12% over short distances in particularly tight areas at the deepest reaches of the pit. As with the road widths, these grades are compliant with industry standards.

Crushed rock will be placed on the haul roads during their construction and on an annual basis during mine operations to provide a good road surface for the haul trucks. This practice will be particularly important in the marls, which are likely to become slippery when wet, or will work into impassable mud with the heavy traffic.

The ultimate pit, will be approximately 240m deep, 1.5km long (east-west), and 0.9km wide (north-south). The top of the HC mineralization is about 150m below the surface, accounting for the large amount of pre-production stripping required.

15.4.4 Mining Dilution and Losses

DMT-MC has applied a correction factor of 0.97 to the ore tonnages expected from the pit to compensate for “porosity.” The effect of this correction is comparable to mining losses, although the correction is not described as such.


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Dilution was incorporated into the mine model by adding 0.5m of hangingwall material and 1.0m of footwall material to the ore zone. The diluting grade was based upon the estimated grades of the diluting material. The hangingwall contact is visually obvious while the footwall contact will be determined by assay. No ore mining losses were included other than the porosity correction. The effective dilution for the in-pit material is about 5%. Given the high-grade and relatively well-defined ore-waste contacts, PAH [felt] that the dilution allowance is appropriate.

15.5 Ore Reserves Statement

Table 15.5.1 summarizes the resulting mining reserves, incorporating corrections and dilution. The assumed price for copper for reserve definition is $0.76/lb (is lower than the current market price of $1.45/lb) . This price was based on past long-term copper prices rounded to a conservative figure. Since the copper will be produced as cathode copper and sold as such, there will be no refining cost. The metal will be shipped directly to buyers and shipping costs are included, offset by cathode premiums. The waste tonnage in the table does not include topsoil removal of approximately 499,000m³. The cost of removing this material is considered within the cash flow. Waste of marl and sandstone was based on wet tonnes.

Table 15.5.1: Las Cruces Diluted Mineral Reserve* (PAH Report, May 2005)


	Ore			Waste	Stripping Ratio
Category	Tonnes (000s)	Cu Grade (%)	Cu (Mlb)	Tonnes (000s)	Waste:Ore
Proven	9,808	6.55	1,415
Probable	6,164	6.72	912
Proven+Probable	15,972	6.62	2,329	232,013	14.53
*Mineral reserve is included in previous mineral resource. Mineral reserve reported at a 1.00% copper cut-off grade. Includes allowances for dilution. Waste of marl and sandstone based on wet tonnes.

PAH [noted] that further review of the reported reserves found that the previous assignment of proven and probable categories in the July 8, 2004 Technical Report was inaccurate, due to the classification criteria. As a result, an adjustment to the relative proportions of proven and probable categories from that reported previously in the July 8, 2004 Technical Report is appropriate. A total of 4.13Mt of proven was reassigned to the probable category. PAH [noted] that this does not change the total mineral reserves nor does it have any impact on the overall economic viability of the project.

The above reserves are based upon accepted engineering practice and are contained within an economically feasible mining plan, and therefore meet the standards for proven and probable reserves under the definitions of reserves as specified by both Canada and the United States. The mineral reserve is included in the previous measured and indicated mineral resource estimate.

PAH [noted] that the ultimate mine plans recover 95% of the available copper in the resource, which is much higher than the 60 to 70% average recovery generally experienced at other deposits. This recovery is understandable; however, given the high degree of structural control of the deposit and the very high grades present (6.5% Cu). At the average grade of 6.62 Cu, an average ore block has a revenue potential of $146/t ($1.00 copper price, no mill recovery), or a net profit of $102/t. Given these levels of potential profit, the open pit mining programs will extract virtually the entire leachable copper deposit.

Note that the primary sulfides below the pit bottom are not included within the reserve, as there is no metallurgical process currently planned to recover this material. There is some


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consideration for future underground mining of the sulfides; however, there are no current definitive plans to do so.

Table 15.5.2 presents the Las Cruces mine mineral reserves as of December 31, 2006, as reported in Inmet’s March 2007 AIF.

Table 15.5.2: Las Cruces Mineral Reserves (as at December 31, 2006)*


Category	Tonnes (000s)	Cu Grade (%)	Contained Cu (t)
Proven	9,790	6.4	629
Probable	7,835	6.0	469
Proven+Probable	17,625	6.2	1,098

*Reserve estimates prepared under the supervision of Michael Doyle, FIMM, FAIMM (Technical Manager, Cobres Las Cruces and QP, as defined by NI 43-101). Reserve estimates are based on the following assumptions:

copper price: US $1.10/lb
exchange rate: €1.00 = US $1.20
open pit cut-off: 1% copper (95.3% of copper in reserve)
underground cut-off: 3% copper (4.7% of copper in reserves)

SRK did not generate the Mineral Reserve estimates and was unable to conduct an audit as prescribed by NI 43-101. However, Inmet reports the estimations and conversions as CIM compliant, as stated in Inmet’s March 21, 2007 Annual Information Form.

15.5.1 Effects on Reserves by Other Factors

A number of factors may affect the reserves of a given mining operation, but are not expected to unduly impact the plans for Las Cruces. PAH’s considerations [were]:

Environmental Restrictions – the property is bounded on two sides by streams, which require setbacks or diversions. While the property plans must respect these requirements, the mine pit limits are minimally affected;
Permitting Restrictions – No unusual permitting restrictions are in place that would change the reserves;
Legal or Title Issues – Legal and Title issues are not anticipated to be troublesome under the Spanish Mining Law;
Taxation – No unusual taxes are present;
Socio-economic Issues – The project is located well away from population centers and other than increasing employment opportunities, no socio-economic issues are anticipated;
Marketing – Copper is a commodity metal with numerous potential consumers within Europe; and
Other Issues – While dewatering is clearly a concern in the design of the mine, the engineering responses anticipated by CLC will not result in any material change to the mine plan or reserves.


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16 Other Relevant Data and Information (Item 20)

There is no other relevant data or information to be included in this report.


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17 Additional Requirements for Development Properties and Production (Item 25)

According to Inmet’s 2007 AIF, the following updates have been reported since the May 27, 2005 PAH Report.

Inmet adjusted the original mine plan from the May 2004 feasibility study for the project to include some underground mining, which will be accessible beginning in 2012 from a ramp in the pit. The underground mine should provide significant benefits because it will reduce waste material by approximately 21Mt, and gives Inmet access to approximately 40,000 additional tonnes of contained copper. The adjusted mine plan calls for 738,000t of ore to be extracted from underground, mainly by drift and fill methods, over a period of four years from 2013 to 2016;
Pre-stripping commenced in March 2006 and by the end of the year, one-third or 9Mm³of the overburden had been removed. Inmet will continue to remove overburden in 2007, and expect to reach ore in the fourth quarter. We expect Las Cruces to mine between 0.6 and 1.3Mt of ore and up to 30Mt of waste each year at an overall strip ratio of 12.7:1. While this ratio is unusually high, it is offset by the very high ore grade. During years 5 through 9 of production (2012 – 2016), we plan to supplement ore from a small underground operation which will mine approximately 740,000t of high grade ore not accessible from the open pit;
The mine will produce copper cathode on site, using technology developed by Outokumpu Technology Oy. The process involves crushing and grinding the ore and then leaching it in the presence of ferric sulphate, sulphuric acid and oxygen in agitated tanks at atmospheric pressure. Once in solution, the copper will be recovered by standard solvent extraction and electrowinning technology. Recoveries of copper are expected to exceed 90% and the plant is designed for a production rate of 72,000t of copper cathode per year. The production of cathode sheets rather than concentrate will diversify Inmet’s copper production, which has been historically in the form of concentrate, and reduce our exposure to smelter and refining treatment charges, which can be volatile. Because cathode copper can be transported by truck, train and ocean-going vessels to plants in Europe that produce wire rod and other copper products, it also lowers freight risk;
Inmet completed basic engineering in April 2006, and a project budget of €380 million was approved at that time to complete the engineering design, procurement, construction and commissioning of the mine and a hydrometallurgical plant capable of annually producing 72,000t of copper as cathodes. A subsidiary of SNC Lavalin Group Inc. has been appointed as lead engineer for the engineering, procurement and construction management of the plant and associated infrastructure. Outokumpu Technology Oy has been contracted to provide grinding, leaching, solvent extraction and electrowinning technology; and
By the end of December 2006, Inmet had completed 55% of the detailed engineering and made commitments to procure €124 million of equipment and services. Total expenditures in 2006 were €65 million. Physical construction of facilities commenced in November 2006 with the building of temporary offices. Civil works on the plant itself began in January 2007


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Sections 17.1 through 17.6 were presented in the PAH Report and based upon the original open pit mine plan.

CLC has a completed Feasibility Study that has been subjected to an independent diligence review. A map of the planned development as of Year 10 of the mining operation is shown in Figure 17-1.

17.1 Mining Operations and Method

The Las Cruces deposit possesses several attributes that dictate the mining method employed. The ore body lies about 150m below relatively flat-lying terrain. The upper 140m of the overburden is comprised of marls (relatively weak calcareous silts and clays), followed by a sandstone unit with substantial water. The ore deposit itself is comprised of high-grade secondary chalcocite copper mineralization.

The depth of the deposit, in conjunction with the grade, suggests that underground methods might be considered. CLC investigated the option, finding that the ground conditions and relatively poor mine recovery made underground mining unattractive.

The feasibility study is based upon open pit mining methods. The operation will be a typical truck and shovel operation, with drilling and blasting expected in the lower levels of the mine. The upper levels of the marl are expected to be free-digging (i.e. no blasting is required), an assumption that has been examined in detail both by equipment manufacturers and the independent engineer. The operating cost estimates assume that the upper 75% of the marl will be free-digging.

Successful development of the mine will require a substantial dewatering plan. Dewatering must be managed carefully at Las Cruces, as water is a precious commodity in southern Spain.

CLC plans to employ contract miners to develop the deposit. Spain supports a number of excellent contract mining companies, such as Cavosa, Peal and Sanchez Y Lago. The use of these companies is not unique to CLC; other mining companies, such as Rio Narcea Gold Mines have successfully employed a similar philosophy.

Operating cost estimates have assumed a mixed fleet of equipment, relying on 140t trucks and 16m³hydraulic shovels for the overburden removal, followed by a combination of 91t trucks and 8.5m ³shovels. The feasibility has suggested a backhoe configuration for the shovels; however, the bench heights of 5m and 10m indicate a front-shovel configuration. The feasibility study further suggests a 1.9m ³shovel coupled with 35t trucks for ore movement, but there is no indication that this level of selectivity will be needed in ore mining.

Drilling and blasting specifications are reasonable regarding drill productivities and spacing. Overburden will be mined in 10m lifts, while ore zones will be mined in 5m lifts.


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17.2 Ore Processing

This section provides a brief description of the ore processing techniques, process plant equipment operation, process plant support facilities, and presents PAH’s comments regarding the plant design. Process plant capital and operating costs are discussed in other sections.

In January 2003, CLC contracted Lurgi Metallurgie GmbH and OTG to prepare a revised technical and economic evaluation of the Las Cruces process plant to optimize costs for the facility, improve copper recovery and simplify the process flowsheet. OTG reviewed the project information and then recommended a revised flowsheet that incorporated atmospheric leaching, Outokumpu Compact SX technology and other innovations anticipated to improve copper recovery and lower capital and operating costs. To confirm their predictions, OTG performed further metallurgical testwork (batch and mini-continuous pilot-scale) at their ORC in Pori, Finland to demonstrate the viability of their proposed flowsheet and supply design criteria for the envisioned process facility. A Feasibility Study was completed in November 2003. Since the Feasibility Study was completed some additional testing has been performed by OTG to provide more design criteria for the basis engineering phase.

The ore processing facility is designed to operate 365days/yr, 24hr/day and process ore at rates ranging from 2,000 to 3,000tpd in years 1 through 3 and then up to 4,000tpd beginning in year 4. The ore will come from the open pit mine and the ore grades will range from 5 to 10% copper. The copper in the ore is primarily found in chalcocite with some minor amounts found in chalcopyrite, tennantite-tetrahedrite complex and enargite. Approximately 66,000tpy of copper cathode will be produced over the envisioned 15-year operation. The plant will be designed to have the ability to produce copper cathode at the instantaneous rate of up to 70,000tpy.

The plant will utilize an atmospheric ferric-leach technology followed by solvent extraction and electrowinning; a processing system that is unique in the copper industry. The technology for this project has been tested, developed and designed by OTG. A similar technology has been used over the past three years successfully in the zinc industry in Finland to directly leach zinc concentrates prior to electrowinning. The more common treatment method historically for, zinc concentrates has been oxidation by roasting prior to acid leaching and electrowinning.


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The process plant currently envisioned for the Las Cruces project consists of unit processes and equipment that are commonly found in the mineral processing industry including crushers and screens, a grinding mill, agitation leach tanks, an oxygen plant, heat exchangers, cooling towers, thickeners, vacuum belt filters, a solvent extraction train, and an electrowinning tankhouse.

Another ferric-leach processing facility has recently been constructed and started up in Laos at the Sepon Copper Project owned by Oxiana Minerals and Rio Tinto. The plant was designed by Bateman and Ausenco of Australia. OTG has signed a long-term copper “off-take” agreement with the owners and is also supplying the SX-EW equipment and technology package for that project. Oxiana will produce about 60,000tpy of cathode copper from a chalcocite ore utilizing a technology that is similar to that which was initially planned for the Las Cruces project; pressure-leach autoclaves with an atmospheric ferric leach system.

A simplified process flow diagram for the planned project is presented in Figure 17-2.

Two types of tailings streams will be produced and will require disposal: filtered leach residue and filtered solids from the neutralization circuit. The leach residue will be about 92% of the total tailings and shows a strong potential for producing acid and metal-bearing leachates based on testing results. The leach residues are made up of about 80% sulfide minerals, primarily as pyrite with minor amounts of chalcocite and chalcopyrite along with constituents like arsenic and mercury. The neutralization solids will be about 8% of the total tailings, are inert, and contain mainly gypsum and magnetite with some metal hydroxides.

The Tailings Storage Facility (TSF) is embedded within the North Dump with the tailings material being confined by a compacted marl embankment, bottom and upper seals. The bottom seal will consist of 1.5m of compacted marl covered by a 1m thick HDPE liner and a 0.5m thick layer of sand and a geotextile material to allow moisture to pass into the sand layer. Any seepage collected in the sand layer will be directed to a seepage collection pond and then pumped to the Contact Water Pond at the process facility. The tailings will be delivered to the TSF by 40t dump trucks, dumped, and then the material will be dozer-spread and roller-compacted in 0.5m thick lifts to an ultimate design density of 90% of standard Proctor maximum dry density and to a thickness of about 25m.

Once the tailings are stacked to their ultimate height, over 5m of compacted marl material will be placed on the top and then covered by geotextile material, 0.6m of sand, more geotextile, 5m of uncompacted marl and then topped with 0.5m of topsoil to grow vegetation.

The embankment or barrier berms will be built to a height of 25m and constructed of upper marl material from the mine pit development. The marl that will not come in contact with tailings will be compacted to a minimum of 95% of standard Proctor density while marl that will come into contact with tailings will be compacted to a minimum of 98% of standard Proctor density. Pore pressures within the berms are not expected to be an issue; however, piezometers will be installed to monitor pressures.

Process water will primarily be obtained from the San Jeronimo wastewater treatment plant (WWTP). Since the discharge from this WWTP is required to maintain the “ecological flow” in the Guadalquiver River downstream of the mine during dry periods, CLC will only be allowed to draw plant effluent, at the rate of about 120L/s, for approximately seven months per year. To provide water for the operation during the remainder of the year, a primary water supply pond will be constructed. Because of the volume of the pond, the retention dam is considered to be a


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large dam; therefore a spillway will be constructed. Spanish regulations require that a detailed operating and contingency plan be submitted for approval with the dam permit application. This document will need to be developed by CLC as part of the final dam design. The wastewater will be chlorinated prior to placement in the pond and will be pre-treated by a reverse-osmosis plant for specific process uses.

PAH reviewed the ore processing plans for the project and found no material deficiencies in the process plant design or process technologies selected. PAH [believed] that the plant layout and equipment selections are reasonable and adequate and that the plant should be able to perform and operate as planned. OTG has prepared process flow diagrams, mass and energy balances, plant layouts and drawings of sufficient detail and accuracy to enable the preparation of a capital cost estimate at the level of accuracy as stated in the Feasibility Study.

17.3 Production Schedule

Las Cruces mining will begin with an 18-month pre-stripping program that should produce a nominal amount of ore for startup in the final months. Ore production will commence at the beginning of year 1 of operations and will continue for almost 15 years. The mining plan is somewhat unique in that later-phase waste stripping will be dumped at the west side of the pit, thus shortening the haulage distance and simplifying the reclamation process at the same time. The mining production schedule is developed from a series of seven phases, designed such that ore will be exposed beneath the thick overburden without requiring excessive stripped inventories. The geometry of the deposit is sufficiently well known that ore exposure is not likely to be a problem.

The production schedule is somewhat unusual, in that the feed tonnage to the mill is not the driving factor in the schedule. For Las Cruces, the electro-winning section will be the primary constraint, limiting total production to 66,000t of cathode per year. A secondary limitation is the crushing and grinding section at 1.3Mt of ore per year. Within these limitations, the mine may produce more tonnage at lower grades, or fewer tonnes at a higher grade.

Table 17.3.1 presents the production schedule.

Inmet’s website lists current project production at 72,000t of cathode per year and a total production of 996,000t over a 15 year mine life.


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Table 17.3.1: Las Cruces Mine and Mill Production Schedule


				Mine Production				Mill Production
		Model						Model
	Total	Diluted		Total	Total	Strip	Feed to	Diluted
	Ore	Cu	Rec. Cu	Waste	Mined	Ratio	Plant	Cu	Rec. Cu	t-Cu
Prod. Year	k-dmt	%Cu	%Cu	kt	k-dmt	w/o	k-dmt	%Cu	%Cu	dmt
-2	-	-	-	17,000	17,000	-
-1	-	-	-	34,195	34,195	-
1	618	8.60	7.66	22,948	23,566	37.1	618	8.60	7.66	47,354
2	1,043	6.97	6.32	14,454	15,497	13.9	1,043	6.97	6.32	65,934
3	1,065	6.82	6.20	17,605	18,670	16.5	1,065	6.82	6.20	66,053
4	758	9.47	8.70	16,093	16,851	21.2	758	9.47	8.70	66,953
5	1,105	6.55	5.98	10,158	11,263	9.2	1,105	6.55	5.98	66,054
6	1,075	6.68	6.15	15,994	17,069	14.9	1,075	6.68	6.15	66,129
7	1,382	7.55	7.00	10,643	12,025	7.7	953	7.48	6.94	66,091
8	1,215	6.43	5.89	16,613	17,828	13.7	1,138	6.34	5.80	65,993
9	1,141	5.48	4.97	10,188	11,329	8.9	1,277	5.69	5.17	66,059
10	1,872	7.54	6.92	10,624	12,496	5.7	933	7.72	7.08	66,047
11	990	6.52	5.98	9,275	10,265	9.4	1,094	6.61	6.04	66,034
12	1,100	4.66	4.27	9,835	10,935	8.9	1,296	5.54	5.10	66,057
13	1,016	6.14	5.66	9,395	10,411	9.2	1,168	6.14	5.66	66,074
14	818	6.51	5.87	5,303	6,121	6.5	1,138	6.40	5.81	66,095
15	774	3.22	2.95	1,690	2,464	2.2	1,311	5.00	4.59	60,136
Total	15,972	6.62	6.05	232,013	247,985	14.5	15,972	6.62	6.05	966,061
Note: Ore tonnes are reported as dry, waste tonnes as wet. Total tonnes combined dry ore tonnes with wet waste tonnes.

17.3.1 Recoverability

The mineable recovery of the Las Cruces Deposit is unusually high for copper deposits as a result of the very high grades. Open pit copper mines typically average between 0.30% copper for deposits in the southwestern US to over 1% in the Chilean and Peruvian deposits. Las Cruces, by comparison, has an average diluted grade of 6.62% copper, recovering well over 95% of the resource.

Following review of the test-work and plant design, PAH [felt] that process recoveries in excess of 90% are achievable. Further discussion of the metallurgical recovery has been provided in Section 14 of this report. The copper production listed in the previous Table 17.3.1 accounts for plant losses.

17.4 Markets

The Las Cruces project would produce approximately [72,000t] of LME Grade A electrowon copper cathode annually over a projected mine life of 15 years. The cathode production would be suitable for all copper fabricators, including those that produce sheets, strip, tube, rod, wire, castings and forgings. Equally important, after a registration process, the copper cathode would be deliverable to terminal markets against copper futures contracts and be attractive to merchants.

Spain does not mine enough copper to be self sufficient in refined production and relies heavily on the import of copper concentrate. By smelting imported concentrates, Spain is able to achieve a balance between production and consumption of refined copper. However, Europe is in a deficit and offers a ready market for any cathode CLC would not sell in Spain.


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The largest end-use markets for refined copper are: heat exchange applications, plumbing, and electrical power generation and delivery. World demand for these products and services has increased, thus increasing demand for refined copper. World production and consumption of refined copper has shown an upward trend during the last five years. Table 17.4.1 shows historic production and consumption.

*Table 17.4.1: Historic Refined Copper (t000s)**

	2000	2001	2002	2003	2004
World Production	14,631	15,539	15,121	15,202	15,879
World Consumption	15,051	14,376	14,392	15,403	16,755
Difference	(420)	1,163	189	(201)	(876)
*Source: CRU International

Table 17.4.2 presents updated historic world production and consumption of refined coper.

*Table 17.4.2: Historic Refined Copper (t000s)**

	2005	2006	2007
World Production	16,568	17,635	18,385
World Consumption	16,894	17,617	18,323
Difference	(326)	18	62
*Sources: CRU International and 2007 Macquarie Research

Copper cathode is priced on the basis of official commodity futures exchange quotes. CLC will receive the LME quote. The monthly average of the spot LME closing price for the month of delivery would be used as the base price. As an alternative to the average pricing, customers would also be allowed to fix the price of all or part of their monthly shipment on any trading day during regular LME trading hours, using the spot LME price.

Premiums are an amount charged over the base price of copper, and are typically negotiated annually. Since much of the copper consumed in Europe is imported, premiums are quoted on the basis of delivery through major Western European ports. As an example, cathode premiums in 2002 ranged from $0.017 to $0.024/lb of copper. Freight from the port to the consuming location is absorbed by the customer. CLC would set its premium based on competitive conditions and freight costs.

Off specification cathodes (off-spec) are typical for the industry. CLC estimates that 5% of their product will be off-spec (either bent or underweight cathodes or containing less than 99.5% copper). Off-spec cathodes can carry deductions ranging from $0.01 to $0.03/lb of copper. According to Inmet’s March 2007 AIF, the estimated cash costs is €0.39/lb of copper. Market prices are currently above $3.00/lb.

The profitability of the project will be dependent in part upon the market price of copper. Factors influencing the market copper price include; international economic and political conditions, expectations of inflation, international currency exchange rates, interest rates, global or regional consumptive patterns, speculative activities, levels of supply and demand, increased production, availability and costs of metal substitutes, metal stock levels maintained by producers, and others and inventory carrying costs.


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Current copper market prices are at historic highs. Further, independent copper price forecast reports suggest that this high trend may continue into the near future. However, as with any forecast, there is no guarantee that high market prices will be maintained.

17.5 Contracts

CLC has not established firm arrangement for either mining or marketing.

CLC anticipates the use of contract miners for the stripping and extraction of ore. Budgetary-level cost estimates have been provided to CLC from three vendors, two of which were reviewed by PAH. While budgetary-level estimates have been provided to CLC, the contracts cannot be considered binding. The contract mining estimates are quite similar to the rates that are paid by other mining operations in Spain that have worked with PAH.

Concentrating costs [were felt] to be adequate by PAH, further discussion of operating costs are presented later within this section. As a direct producer of cathode copper, there will be no smelting and refining charges other than those incurred by CLC in the normal course of operations.

Transportation costs will be in metal cathode form. The cost of transportation and selling costs to the consumer is expected to be roughly equivalent to the cathode premium, thus offsetting each other. Explicit treatment of these costs has been included within the detailed cash flows, but the actual transportation costs can only be estimated until either contracts are signed or specific consumers can be identified. The present assumptions employed at the cash flows is an adequate treatment for this stage of analysis.

Per industry norms specific contracts for mining projects are considered sensitive in nature. Due to the limitations of the information and data contained within the public domain, SRK is unable to comment on the terms of the Las Cruces Project contracts.

17.6 Environmental Considerations

Environmental issues for Las Cruces are typical for the mining industry regarding noise, dust and ongoing environmental management. Of these, the groundwater management and the reclamation process will result in material impacts on cash flow. Costs for both have been adequately factored into the economic analysis.

17.6.1 Permitting

Of importance in the permitting process for Las Cruces was the issuance of a positive Declaration Environmental Impact (DEI). This issuance gave CLC the right to apply for a mining concession. A positive DEI for Las Cruces was issued by the Provincial Delegation of the Regional Ministry for the Environment on May 9, 2002. CLC was granted the Mining Concession on August 6, 2003.

As of April 2005, CLC [had] obtained the following key permits:

DEI;
Mining Concession;
Integrated Environmental Authorization (IPPC); and
Special Plan (LOUA).


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Table 17.6.1.1 presents a summary of the status of the 49 permits CLC has identified for operation of the project. As of April 2005, seven permits were being processed for approval, and 14 permits were pending application to the regulatory authority. Of the pending permits, three require ownership or authorization from the landowner for submittal. CLC expects to receive approval for all the permits being processed or pending by September 2006. There are a number of permits outstanding in order for the project to proceed. CLC has continued to maintain the current emphasis on permitting including land acquisition and supporting detailed engineering to avoid potential delays.

Permits for diversion of surface waters will first require registering and platting of affected watercourses, including a policing zone (100m beyond the design flood stage). This is a relatively new Spanish regulation that postdates most property boundaries. Setting of the stream boundaries affects the total land area owned by the individual property owners. As such, approval by the affected property owners is required prior to acceptance of the final boundaries. CLC plans to wait until the lands have been purchased before they submit the permit application. This approach should mitigate any potential delays associated with this permitting.

Authorization for the withdrawal and reinjection of groundwater was issued by the water use authority in November 2003. The proposed reinjection system was reviewed by the Spanish Geological Society. Previously, no similar system had been permitted in Spain. Pumping and reinjection systems for aquifer management have been successfully used on mining and petroleum projects and is an accepted industry practice. For example, in the Carlin Trend in the state of Nevada, USA, re-injection of mine dewatering water is done at Jerritt Canyon mine, Cortez Joint Venture, Newmont’s Lone Tree mine, and Barrick’s Goldstrike mine. Re-injection is done by both injection wells and infiltration basins and is regulated by both the State of Nevada and the U.S. EPA under the Underground Injection Control Permit program.


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Table 17.6.1.1: Las Cruces Summary of Permit Status

			Regulatory	Application	Status/Auth
Category		Permit	Authority	Date	Date	Notification	PAH Comment
Mining- Hydrometallurgy Project	1	Declaration of Environmental Impact	DPCMA	3/23/2003	5/9/2002	5/16/2002	Approved EIA serves as basis for project environmental status and underlies other Central government and regional agency approvals.
	2	Mining Concession	DPCEDT	3/23/2003	8/6/2003	8/6/2003	Granted by Central government with 30 year term and two possible 30-year extensions.
	3	Mining Project Optimization	DPCEDT	11/23/2004	1/29/2004	2/6/2004
	4	Integrated Environmental Authorization	DPCMA	22/12/03	3/10/2005	3/10/2005	Requirement for preparation of this document is a result of EU Directive 96/61 being incorporated into Law 16/2002. This will combine several permits/approvals including the DEI, DPMT, DPH and hazardous waste management into an integrated control of pollution approach.
	5	Special Plan	DPCOPT	4/16/2004	2/11/2005	3/9/2005	Submittal to Regional Ministry of Public Works and Transport (DPCOPT)
Dewatering-Reinjection System	6	Dewatering-Reinjection Authorization	CHG	7/15/2002	10/30/2003	11/10/2003	Addresses pit dewatering and re-injection. Has four conditions as part of approval including establishing contingency plan prior to startup. Analysis of water in each dewatering sector prior to startup, daily monitoring requirements, and monitoring and control plan
	7	Crossings over Regional Highways	DPCOPT	2/11/2003	2/19/2003	3/17/2003
	8	Crossings over National Highways	Devpt. Min.	2/11/2003	6/10/2003
	9	Crossings over Livestock Trails	DPCMA	5/10/2001	BP		Process has been suspended until just before construction starts.
Water Concessions, Consumptive Use (Incl supply-discharge system and supply pond) and related permits	10	Water Concessions, Consumptive Use	CHG	12/19/2000	6/18/2004	6/15/2004	Concessions required from Ministry of Environment for three sources of water including Sewage Treatment Plant (STP) effluent, groundwater flowing to pit, groundwater for domestic use. CHG has issued report indicating this is compatible with the Guadalquivir Hydrologic Plan (PHG) and Regional Ministry of Health issued a favorable report on suitability of using STP effluent for industrial purposes. Final approval is pending change in agency personnel following recent change in government.
	11	Utilization of Rainwater	CHG	12/14/2001	PA
	12	Occupation DPMT by supply/Discharge Pipes	Costas	5/27/2002	6/30/2003	7/8/2003
	13	Occupation DPMT easement zone by supply/discharge pipes	DPCMA	5/14/2002	9/20/2002	10/3/2003
	14	Crossing and occupation of regional highways by supply/discharge pipes	DPCOPT	2/11/2003	2/19/2003	3/17/2003
	15	Crossing of national highways by supply-discharge pipes	Devpt. Min.	2/11/2003	5/26/2003	6/10/2003
	16	Crossing under railways by supply- discharge pipes	Renfe	4/29/2003	8/26/2003	10/29/2003
	17	Technical Report for supply/discharge pipes to cross Viar Canal land	CHG		6/2/2003		Technical report to support approval of authorization to cross Via Canal Land.
	18	Authorization for supply/discharge pipes to cross Via Canal land	CHG		PA		Approval is pending final approval of construction details such as width of pipeline corridor and method of crossing canal. Expected 11/05.
	19	Crossing of Livestock Trails	CMA	5/10/2001	BP		Same as 9, above.
	20	Sundry Crossing (power lines, underground telephone line, 85 irrigation channels, Emasesa pipeline)	Various		PA		Expected 11/05.
Discharges and Related Permits	21	Authorization for discharge into DPMT	CAN	5/14/2002	10/21/2003	10/28/2003
	22	Occupation of DPMT easement zone by discharge works	CMA	5/14/2002	10/21/2003	10/28/2003
	23	Occupation of DPMT by discharge works	Costas	5/14/2002	9/23/2003	10/9/2003
	24	Authorization of discharge into DPH	CHG	4/26/2001	3/11/2003	3/20/2003
Demarcation of Public Water Domain (DHP) and Policing Zone (ZP)	25	Technical Feasibility Report	CHG	5/14/2001	11/6/2002
	26	Demarcation of DPH and ZP	CHG		PA		Demarcation of the Public Domain Waters (DPH) and the Policing Zone (ZP), which is a 100m zone on each side of the DPH, is required for final approval of discharge of unaffected runoff from the project to the Gamacha, Molinos and Almendrillos drainages. This is on hold pending land purchase negotiations to define the limit of land that CLC will own. Expected 4/06.
Demarcation of Public Inland-Maritime Domain (DPMT) and Easement Zone	27	Demarcation of DPMT of Rivera de Huelva	Costas	7/4/2001	12/11/2002	2/4/2003
Mining- Hydrometallurgy Project	28	DEI on power line	MMA	4/24/2002	10/9/2003	10/27/2003
	29	Authorization of Preliminary Project	Econ. Min.	4/25/2001	BP		Authorization for high voltage power line and substation. Expected 4/06.
	30	Authorization of Construction Project	Econ. Min.		PA		Authorization to construct power line and substations including right of way. Expected 5/06.
	31	Crossing over DPH	CHG	2/21/2003	4/28/2003	6/10/2003
	32	Crossings over Livestock Trails	DPCMA		PA		Same as 9, above.
	33	Crossings of SE-520 Regional Highway	CPCOPT		PA		No Information.
Diversion of Medium Voltage Power Line and Related Permits	34	Environmental Report on Diversion	DPCMA	4/7/2003	7/29/2003	8/11/003
	35	Authorization of Preliminary Project	DPCEDT	4/7/2003	8/28/2003	9/1/2003
	36	Authorization of Construction Project (Year 7+)	DPCEDT		PA		Expected 4/06.
	37	Crossings over DPH	CHG	4/7/2003	7/11/2003	11/4/2003
Works and Installations in DPH & ZP	38	Technical Feasibility Report	CHG	7/1/2001	5/28/2003	6/6/2003
Works and Installations in DPH & ZP	39	Authorization of works and installations in DPH and ZP	CHG		PA		Subject to demarcation of DPH and DPH-ZP as addressed under approval No. 26. expected 11/05.
Steam Diversions	40	Technical Feasibility Report	CHG	5/25/2001	12/17/2002
	41	Authorization of Stream Diversions	CHG		PAL		Expected 3/06.
	42	Disencumbrance/Encumbrance	Revenue Min.		PAL		No Information.
SE-520 Underpass	43	Authorization of Project	DPCOPT		PA		Expected 12/05.
SE-520 Roundabout	44	Authorization of Project	DPCOPT		PA		Expected 8/05.
Livestock Trails	45	Demarcation	DPCMA	4/2/2001	BP		Expected 11/05.
	46	Route Modification	DPCMA	5/4/2001	BP		Expected 6/06.
	47	Authorization of Occupation by supply-discharge pipes and dewatering-reinjection system	DPCMA	5/10/2001	BP		Expected 9/06.
Rural Tracks	48	Diversion/Abolition	Councils		PAL		No Information.
Municipal Licenses	49	Works and Activity Licenses from Gerena, Guillena, Salteras, La Algaba, La Rinconda and Seville Town Councils.	Councils		PA		Expected between 10/05 and 4/06.
Note: BP: Being Processed; PA: Application Pending; PAL: Application Pending, Subject to Land Acquisition.


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While much of the project site is under single ownership, there are approximately 90 landowners involved with acquisition of the entire property (including water and power transmission corridors). CLC has been in contact with several owners, and has indicated that they would offer a price greater than the local fair market real estate value to acquire the properties. Under Spanish law, the mining concession allows the mine owner to acquire land through expropriation if no other avenue is available.

17.6.2 Environmental Restoration

The environmental restoration plan for Las Cruces is based on the plan submitted as part of the Project EIS dated March 2001. The current plan incorporates the results of this study plus modifications for project optimizations and incorporation of all permit conditions specified in the DEI. Two principal modifications to the original restoration plan are:

	1.	The artificial lake to be created after the closure of the open pit has been replaced by a landfill for inert waste material, mainly from construction and demolition works. This alternative was analyzed and developed in a report prepared by INTECSA-INARSA entitled “Study of Post-Closure Inert Material Dry Seal, Las Cruces Project” dated December 2001. The new restoration plan for the pit will include partial backfill with marl, which will create a residual void completely lined with marl as a dry seal.

	2.	The surface areas to be restored have been modified in line with the new mining plan. Subsequent to the issuance of the November 2003 Feasibility Study, an optimized mining plan has been prepared by IMC of Tucson, Arizona. The surface areas in IMC’s mining plan differ slightly from those in the November 2003 Feasibility Study; in PAH’s opinion, it will not materially impact the restoration plan or costs.

PAH reviewed the other required modifications stated the DEI and found them to be achievable within the revised restoration plan. However, the 2003 Feasibility Study did not address in any detail the pit backfilling with inert materials, nor has CLC been required to submit a formal plan or restoration costs to the environmental authorities. It is estimated that under this scenario the pit backfilling would occur over more than 30 years. CLC recognizes that a specific Performance Project must be presented for authorization regarding the plan for backfilling with inert materials.

[As of March 2005], CLC [had] not made a decision as to whether the management of the backfilling operation (in effect, a controlled land fill) would be sold to an existing European landfill company, or under CLC management. PAH [considered] two options are being viable for the pit backfill: (1) with inert materials depending on availability, or (2) if sufficient inert materials are not available, as a worst case, backfilling with low-ARD mine waste materials (marl and sandstone) from the waste dumps.

The Las Cruces project will utilize on-going environmental restoration comprised of:

Temporary planting for erosion protection;
Final revegetation during various project phases; and
Removal of infrastructure at project closure.

Temporary planting will be implemented for erosion prevention and landscaping primarily to dump surfaces and stockpiled materials that will be covered subsequently with other materials or


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moved elsewhere. Permanent restoration will be applied to the final reclaimed surfaces in dumps, tailings storage, mine waste storage, or areas that were used for temporary stockpiles. CLC has identified nine main areas for environmental restoration. Restoration will be initiated at the outset of the two-year construction period, continue throughout the 15-year production period, and culminate during two years after production ceases. Approximately 29% of the permanent restoration will be completed during construction, and 32% during the two years after production ceases. The total restoration budget is €10.74 million per current cash flow.

Visual berms and waste piles will be graded at approximately 4 horizontal to 1 vertical slopes in all areas visible to the public. These slopes approximately match the rolling terrain in the project area. Slopes will be graded and seeded. The tailings disposal facility will be capped and seeded as placement of the tailings progresses across the facility. Similarly, the pit will be partially backfilled behind the mining operation as it proceeds across the proposed footprint of the pit. Backfilling of the pit will consume a portion of the marl produced over the life of the mine, limiting the size of surface waste dumps that will require reclamation.

Typically, the Spanish government requires posting of a bond based on the total disturbed area of the mine. Since CLC will be reclaiming the site throughout the life of the mine, it is anticipated that the largest bonding requirements will be early in the project, and will continuously decrease as mining and reclamation progresses. The current environmental bond is placed at €13.8 million, and an additional €5 million labor bond at start-up of the project.

Table 17.6.2.1 shows CLC’s annual budget for environmental restoration, unescalated, and as presented in the 2003 Feasibility Study. Almost 50% of the environmental restoration will be completed during the first five years of production. Budgets in the initial years (Years –2 and -1) include costs for development and seeding of visual berms, revegetation, and habitat development in the diverted stream channels. Operational years include costs for ongoing restoration of waste dumps and tailings disposal areas that have reached final proposed grades. Years 16-17 for post mining activities include final capping, grading, revegetation and other activities associated with reclamation of the waste dumps, livestock trails, power lines, tailings facility, mill site, water diversion and storage ponds, roads and other facilities. For the current estimate, the costs were escalated by 2.5% to reflect inflation.

The environment reclamation costs presented on a unit basis appear to be conservative. As an example, the unit cost associated with capping waste rock dumps and tailings facility with a marl cover is €15,400/ha. Restoration costs vary from €1,650/ha for temporary pasture on topsoil stockpiles to €25,900/ha for shrub – open Mediterranean forests.


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Table 17.6.2.1: Las Cruces Annual Restoration Areas and Budgets

			Cumulative %
	Restoration Area	% Permanently	Permanently	Annual Budget €
Project Year	(ha/yr)	Revegetated (%)	Revegetated (%)	(000’s)
-2	78	11.6	11.6	1,316
-1	114	17.0	28.7	1,814
1	82	12.2	40.9	1,417
2	33	4.9	45.8	835
3	9	1.3	47.2	362
4	9	1.3	48.5	362
5	9	1.3	49.9	197
6	9	1.3	51.2	197
7	9	1.3	52.5	163
8	27	4.0	56.6	495
9	27	4.0	60.6	495
10	20	3.0	63.6	389
11	20	3.0	66.6	389
12	3	0.4	67.0	43
13	3	0.4	67.5	43
14	3	0.4	67.9	43
15	3	0.4	68.4	43
16	62	9.3	77.6	477
17	150	22.4	100.0	1,661
Total	670	100.0	100.0	10,740
Note: Annual budget includes 10% contingency.

17.6.3 Groundwater Management

The groundwater pumping system is needed to depress the phreatic surface to elevations deeper than the mining zone. In the initial two years, this level must be depressed to an elevation below the sandstone aquifer (approximately 160m) to allow commencement of mining. The modeling reported in the Feasibility Study is based on drawing the groundwater down to the maximum mining depth as soon as possible. The model was used to evaluate the actual pumping required to depress the groundwater level at the rate that more closely approximates the proposed mining schedule.

CLC estimates that approximately 2.4Mm ³ of water will need to be pumped in the initial 2 years to account for the water volume stored in the portion of the sandstone aquifer removed in preparation for mining. The model assumes that the water will be extracted and then re-injected, as the local aquifer is already depleted from local overuse.

Once operational, CLC will withdraw groundwater from the sandstone and Paleozoic layers at a rate of 150L/s to 220L/s, which will be reinjected. The proposed reinjection system appears feasible to minimize impacts to the Niblea-Posadas aquifer for most groundwater users in the area. Because the pumping and reinjection is a closed system, no adverse impacts to water quality within the aquifer are anticipated with slight rises in the aquifer levels occurring in all areas except for the area around the Render Sur. The Render Sur, a commercial rendering facility located relatively close to the proposed mill site, will be most impacted as groundwater levels will likely be depressed to below the plant’s current well depth. To mitigate this impact, water will be supplied to the Render Sur facility by the Las Cruces project.


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17.6.4 Surface Water Management

The Las Cruces project will modify the actual surface water flows that currently exist. The Molinos, Garnacha and two other small streams will be diverted so that the mining and ore processing operations will have minimal impact upon the existing watershed. Runoff from the site will be maintained as either contact or non-contact water. Any water which enters the pit, waste dumps, tailings facility or haul roads, or may otherwise come in contact with potentially acid generating rock, will be maintained as contact water. This water will be collected and will report to either a holding pond and added to the process water stream or the contact water treatment pond, and treated prior to reuse or release. Non-contact water will report to sedimentation ponds prior to discharge. All runoff management systems have been designed based on a 100-year storm event. After mine closure, all potentially reactive surfaces will be covered and there will no longer be any contact water. The contact water treatment plant, however, will be kept operational as a contingency measure. PAH [believed] that the surface water management plan is reasonable and well conceived and should achieve the planned goals for diversion, collection and treatment, as required.

SRK is unable to verify the status of environmental and permitting data and information due to lack of access to specific Project data.

17.7 Economic Analysis

Inmet estimates that initial capital costs will total €380 million with interest, reclamation, bonding and other financing costs adding another €46 million. Sustaining capital over the life of the mine is estimated to be €24 million. The estimated cash operating costs of €0.39/lb of copper place Las Cruces in the bottom quartile of world producers. (Inmet AIF, March 2007).

The royalty terms and conditions presented in this section were provided to SRK by IRC.

Because IRC holds a royalty interest, its economics depend on the volume of copper produced and the price of copper. Royalty revenues were calculated based on annual production of 72,000t or 158.7Mlbs of cathode copper at copper prices of $1.50, $2.00, $2.50 and $3.00/lb. Deductions were assumed to total $0.10/lb for freight and sales expenses. Annual revenues range from $3.3 million at $1.50/lb to $6.9 million at $3.00/lb. Totals over the projected 15 year mine life range from $46 million at $1.50/lb to $96 million at $3.00/lb.

Changes in production levels will result in proportional changes in royalty revenues.

Table 17.7.1 presents the royalty revenue based on LoM production of 996,000t or 2.2Blbs of cathode copper. The average annual output is 72,000t or 158.7Mlbs and the deductions in the royalty calculation total $0.10/lb.


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Table 17.7.1: Royalty Revenue versus Copper Price (1.5% Royalty Rate)

			Revenue (US$millions)
Year	Ore Production (kt)	Cathode Production (Mlb)	$1.50/lb	$2.00/lb	$2.50/lb	$3.00/lb
2008	18	39.7	0.8	1.1	1.4	1.7
2009	72	158.7	3.3	4.5	5.7	6.9
2010	72	158.7	3.3	4.5	5.7	6.9
2011	72	158.7	3.3	4.5	5.7	6.9
2012	72	158.7	3.3	4.5	5.7	6.9
2013	72	158.7	3.3	4.5	5.7	6.9
2014	72	158.7	3.3	4.5	5.7	6.9
2015	72	158.7	3.3	4.5	5.7	6.9
2016	72	158.7	3.3	4.5	5.7	6.9
2017	72	158.7	3.3	4.5	5.7	6.9
2018	72	158.7	3.3	4.5	5.7	6.9
2019	72	158.7	3.3	4.5	5.7	6.9
2020	72	158.7	3.3	4.5	5.7	6.9
2021	72	158.7	3.3	4.5	5.7	6.9
2022	42	92.6	1.9	2.6	3.3	4.0
Total	996	2,196	46.1	62.6	79.0	95.5
NPV_8%			25.8	35.0	44.2	53.4


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Figure 17-1: Mine Development Plan, Mine Status Year +10 Main Infrastructure Items

Source: PAH Report, May 27, 2007


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Figure 17-2: Las Cruces Process Flow Diagram

Source: PAH Report, May 27, 2007


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18 Interpretation and Conclusions (Item 21)

18.1 Exploration Conclusions

PAH found that the drilling and sampling programs by Rio Tinto on the Las Cruces deposit were reasonably conducted using industry standard procedures and techniques. Drillhole coverage is adequate for Feasibility Study, with some infilling recommended. Abundant evidence exists of the tendency to preferentially loss of chalcocite during the drilling and sampling, particularly in lower core recovery intervals, resulting in lower copper grades than may actually be present. As a result, there is lower confidence in the copper grades reported for samples in the lower core recovery zones. The evidence indicates, however, that the copper grades should tend to be higher than reported. Analyses were conducted using standard procedures and were accompanied by reasonable check assay programs for quality control, which showed no significant grade bias or systematic analytical problems. More recently, 14 infill drillholes were drilled to check the initial mining area and the available information is summarized in the middle of Section 15, however, the information was not complete at the time of this report, and will be reviewed when the information is available.

18.2 Other Relevant Information

18.2.1 Geologic Evaluation Conclusions

PAH [found] that the degree of geologic evaluation of the Las Cruces deposit is reasonable for supporting a feasibility study. Geologic interpretation for the resource model was consistent with reasonable correlations between drillholes. The majority of the geologic data is from previous exploration conducted by Rio Tinto. The deposit geometry is reasonably defined from drillholes averaging around 50m spacing. The understanding of the copper grade distribution, however, has been somewhat impacted by poor core recovery, which has been evaluated further by 14 infill drillholes (results not yet complete at the time of this report). Previous evaluation of the secondary sulfide has found this mineralization to be subject to wide grade variations over short lateral distance, as would be expected given the nature of the mineralization.

18.2.2 Resource Estimation Conclusions

PAH [believed] that the DMT secondary sulfide resource model was carefully created using standard engineering methods appropriate for this deposit. The model provides a reasonable representation of the distribution of the secondary sulfide mineralogic zones. PAH [believed] that the resource model, due to natural deposit variability and due to the excluded sample and composite data (about 30% of composite data excluded), may be a modest estimator of grade locally, but believes that on a global basis the estimation is reasonable and may, in fact, be conservative. The resource estimate compares reasonably with those from previous resource models and grade differences are due to intentional changes in modeling methodologies. Based on the data available for the 2003 DMT Feasibility, PAH [believed] that the model provides an acceptable basis for which subsequent mine engineering work can be conducted in order to delineate mineable reserves acceptable to the SEC or any other financial institution.


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CLC plans for ongoing infill drilling over the life of the mine, in order to increase the resource confidence for short-term mine planning and reduce local grade variability. At the time of this report, 14 infill holes had been drilled in the initial mining area, with analytical results only available for 5 of these holes. PAH [considered] the review of the DMT Feasibility resource model with the new drilling to be an important check of the model sensitivity to this change in data support. Once all of the data for the 14 holes are available, PAH was to review the results.

18.2.3 Reserve Estimation Conclusions

PAH found that the mineral reserve and pit design procedures generally followed industry-accepted practices and provide a reasonable representation for the project, although there may be room for improvement on the design. The pit designs are fairly insensitive to changes in the price of copper, within a range of $0.75 to $0.90/lb, and cut-off grade, from zero up to 1.5% copper. At higher copper prices or lower cut-off grades, the mineral reserve nearly completely extracts the mineral resource of supergene enrichment.

18.2.4 Economic Analysis

PAH found the cash flow model to be complete, inputs were accurate and reflected project costs and development plans. The result indicates a positive project cash flow that justifies the material being categorized as reserves.

SRK is not aware of any issues that have not been otherwise disclosed in this report which would materially affect the Project.


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19 Recommendations (Item 22)

While a number of recommendations have been made throughout the report, none [were felt] to be fatal flaws within the components of the feasibility study. The following matters were identified in the course of the review:

PAH [noted] that CLC plans to drill an additional 40 to 50 core holes to improve definition and knowledge of the deposit. Of these, 14 holes have been drilled to date, with some of the results not yet complete at the time of this report;
The ultimate pit as designed is somewhat steeper than the recommended slope on the eastern side of the pit. While there is a good possibility that the slopes will prove to be acceptable after mining has provided actual data, the pit should be modified to conform with the 28º slopes recommended by Geocontrol, or, alternatively, have Geocontrol reassess the [recommended] slopes;
The mining plans envisioned by DMT require the use of 1.8m³loaders and 35t trucks for ore mining. There is no demonstrated need for mining selectivity of this order, therefore PAH recommended that the ore mining be accomplished with 8.5m³loader and 91t truck. This change should result in cost savings;
The contract mining estimates should be resubmitted for binding estimates. While material changes in the rates are unlikely, binding quotes are normally required prior to project commencement; and
There are some permits outstanding in order for the project to proceed. To avoid potential delays, PAH [recommended] that CLC maintain the current emphasis on permitting, including land acquisition and supporting detailed engineering.

PAH [felt] that the project Feasibility Study, as updated and described in this report, warrants further consideration, and should proceed with both financing and development.

SRK makes no further recommendations in regard to the Project or the royalty holder.


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20 References (Item 23)

Pincock Allen & Holt (May 27, 2005), Technical Report of the Las Cruces Copper Project, Southern Spain

Inmet Mining Corporation (2006), Inmet Mining Corporation Annual Report 06

Inmet Mining Corporation (March 21, 2007), Annual Information Form

International Royalty Corporation (December 21, 2007) International Royalty Enters Agreement to Acquire 16 Royalties from Rio Tinto

www.inmet-mining.com (January 2008)

www.metalseconomics.com (January 2008)


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21 Glossary

21.1 Mineral Resources and Reserves

21.1.1 Mineral Resources

The mineral resources and mineral reserves have been classified according to the “CIM Standards on Mineral Resources and Reserves: Definitions and Guidelines” (November 2005). Accordingly, the Resources have been classified as Measured, Indicated or Inferred, the Reserves have been classified as Proven, and Probable based on the Measured and Indicated Resources as defined below.

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.

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 drillholes.

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 drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

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 drillholes that are spaced closely enough to confirm both geological and grade continuity.

21.1.2 Mineral Reserves

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.


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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.

21.2 Glossary

Table 21.2.1: Glossary

Term	Definition
Assay:	The chemical analysis of mineral samples to determine the metal content.
Capital Expenditure:	All other expenditures not classified as operating costs.
Composite:	Combining more than one sample result to give an average result over a larger distance.
Concentrate:	A metal-rich product resulting from a mineral enrichment process such as gravity concentration or flotation, in which most of the desired mineral has been separated from the waste material in the ore.
Crushing:	Initial process of reducing ore particle size to render it more amenable for further processing.
Cut-off Grade:	The grade of mineralized rock, which determines as to whether or not it is economic to recover its gold content by further concentration.
Dilution:	Waste, which is unavoidably mined with ore.
Dip:	Angle of inclination of a geological feature/rock from the horizontal.
Fault:	The surface of a fracture along which movement has occurred.
Footwall:	The underlying side of an orebody or stope.
Grade:	The measure of concentration of gold within mineralized rock.
Hangingwall:	The overlying side of an orebody or slope.
Haulage:	A horizontal underground excavation which is used to transport mined ore.
Kriging:	An interpolation method of assigning values from samples to blocks that minimizes the estimation error.
Level:	Horizontal tunnel the primary purpose is the transportation of personnel and materials.
LoM Plans:	Life-of-Mine plans.
Milling:	A general term used to describe the process in which the ore is crushed and ground and subjected to physical or chemical treatment to extract the valuable metals to a concentrate or finished product.
Mineral/Mining Lease:	A lease area for which mineral rights are held.
Ore Reserve:	See Mineral Reserve.
Sedimentary:	Pertaining to rocks formed by the accumulation of sediments, formed by the erosion of other rocks.
Smelting:	A high temperature pyrometallurgical operation conducted in a furnace, in which the valuable metal is collected to a molten matte or doré phase and separated from the gangue components that accumulate in a less dense molten slag phase.
Stope:	Underground void created by mining.
Strike:	Direction of line formed by the intersection of strata surfaces with the horizontal plane, always perpendicular to the dip direction.
Sulfide:	A sulfur bearing mineral.
Tailings:	Finely ground waste rock from which valuable minerals or metals have been extracted.
Thickening:	The process of concentrating solid particles in suspension.
Total Expenditure:	All expenditures including those of an operating and capital nature.


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Abbreviations

The metric system has been used throughout this report unless otherwise stated. Currency is in Euros and U.S. dollars. Market prices are reported in US$ per pound of copper. Tonnes are metric of 1,000kg, or 2,204.6lbs. The following abbreviations are used in this report.

Table 21.2.2: Abbreviations

Abbreviation	Unit or Term
A	ampere
Ag	silver
Au	gold
°C	degrees Centigrade
CIM	Canadian Institute of Mining
CLC	Cobre Las Cruces S.A.
cm	centimeter
CSA	Canadian Securities Administrators
Cu	copper
°	degree (degrees)
DEI	Declaration Environmental Impact
dia.	diameter
DMT-MC	DMT-Montana Consulting
€	Euro
EIS	Environmental Impact Statement
ft	foot (feet)
g	gram
g/L	gram per liter
gpt	grams per tonne
g/cm³	grams per cubic centimeter
ha	hectares
HC	High Copper
HC4	High Copper lens Number 4
HCF	High Copper Footwall
HCH	High Copper-High Density
HCI	High Copper Intrusive
HCL	High Copper-Low Density
HDPE	Height Density Polyethylene
IMC	Independent Mining Consultants
IPPC	Integrated Environmental Authorization
IRC	International Royalty Corporation
JORC Code	Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves
kg	kilograms
km	kilometer
kt	thousand tonnes
kV	kilovolt
kW	kilowatt
kWh	kilowatt-hour
kWh/t	kilowatt-hour per metric tonne
L	liter
L/s	liters per second
lb	pound
LoM	Life-of-Mine
m	meter
m²	square meter
m³	cubic meter
masl	meters above sea level
MC	Master Composite
MCL	Medium Copper Low Density
mm	millimeter
Mlb	million pounds


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Abbreviation	Unit or Term
Mt	million tonnes
MW	million watts
NI 43-101	Canadian National Instrument 43-101
OCR	Outokumpu Research Center
OK	ordinary kriging
OKTOP	Outokumpu TOP
OTG	Lurgi Metalurgie GmbH and Outokumpu Technology Group
%	percent
PAH	Pincock Allen & Holt
PLS	Pregnant Leach Solution
QA/QC	Quality Assurance/Quality Control
QP	Qualified Person
s	second
SEC	U.S. Securities & Exchange Commission
SRK	SRK Consulting (US), Inc.
SX/EW	solvent extractioin electrowinning
t	tonne (metric ton) (2,204.6 pounds)
tph	tonnes per hour
tpd	tonnes per day
tpy	tonnes per year
TSF	tailings storage facility
µ	micron or microns
V	volts
W	watt
WWTP	San Jeronimo wastewater treatment plant
yr	year


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Appendix A
Certificate of Author

	SRK Consulting (U.S.), Inc.
	7175 West Jefferson Avenue, Suite 3000
	Lakewood, Colorado
	USA 80235
	e-mail: denver@srk.com
	web: www.srk.com
	Tel: 303.985.1333
	Fax: 303.985.9947

CERTIFICATE of AUTHOR

I, Neal Rigby, CEng do hereby certify that:

1.	I am a Principal of: SRK Consulting (US), Inc. 7175 W. Jefferson Ave, Suite 3000 Lakewood, CO, USA, 80235

2.	I graduated with a BSc degree in Mineral Exploitation with first class honors in 1974 and a PhD in Mining Engineering in 1977 both from the University of Wales, UK.

3.	I am a member of the Institute of Materials, Mining and Metallurgy.

4.	I have worked as a mining engineer for a total of 33 years since my graduation from university.

5.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

6.	I am responsible for the content, compilation, and editing of all sections of the technical report, titled, International Royalty Corporation NI 43-101 Technical Report, Las Cruces Project Royalty, Sevilla Province, Spain, and dated February 5, 2008 (the “Technical Report”) relating to the Las Cruces Project royalty. I have not personally visited the Las Cruces Project.

7.	I have not had prior involvement with the property that is the subject of the Technical Report.

8.	I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

9.	I am independent of the issuer applying all of the tests in Section 1.4 of National Instrument 43-101.

Group Offices in:	North American Offices:
Australia	Denver	303.985.1333
North America	Elko	775.753.4151
Southern Africa	Reno	775.828.6800
South America	Tucson	520-544-3688
United Kingdom	Toronto	416.601.1445
	Vancouver	604.681.4196
	Yellowknife	867-699-2430

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

Dated February 5, 2008.

/s/ Neal Rigby

Neal Rigby, CEng., MIMMM, PhD (signed)

International Royalty Corporation Las Cruces Project Royalty NI 43-101 Technical Report, Spain, February 5, 2008.

Dated 5 February 2008.

/s/ Neal Rigby

Dr. Neal Rigby, CEng, MIMMM, PhD (signed)