Technical Report entitled "Technical Report on an Update to the Fenix Project

Saprolitic Fe₂O₃ < 30%

-0.0009304

Saprolitic Fe₂O₃ > 30% (includes transition)

EX-99.106 107 dex99106.htm TECHNICAL REPORT ENTITLED "TECHNICAL REPORT ON AN UPDATE TO THE FENIX PROJECT Technical Report entitled "Technical Report on an Update to the Fenix Project

Exhibit 99.106

HMI Nickel Inc. - Fenix Project

Technical Report

HMI Nickel Inc. (formerly Skye Resources Inc., “HMI” or “Skye”): Technical Report

TECHNICAL REPORT ON AN UPDATE TO THE FENIX PROJECT, IZABAL GUATEMALA

NOVEMBER 19, 2008

(Effective date September 15, 2007)


Prepared by		John B. Scott Manager of Engineering Fenix Project HMI Nickel Inc.

		Colin B. McKenzie Consultant to HMI Nickel Inc.

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Table of Contents


IMPORTANT NOTICE		7

1.	Summary	9

	1.1 General	9
	1.2 Geology, Deposit Evaluation and Mineral Resource Estimates	10
	1.3 Mineral Reserve Estimates	14
	1.4 Mining and Metallurgical Plant Operations	15
	1.5 Economic Analysis	16
	1.6 Hydrometallurgical Processing of Limonite	17

2.	Introduction	17

	2.1 General	17
	2.2 Geology, Deposit Evaluation and Mineral Resource Estimates	17
	2.3 Metallurgical Testwork	18
	2.4 Mineral Reserve Estimates	18
	2.5 Mine and Process Plant Feasibility Study	19
	2.6 Qualified Persons and Site Visits	20

3.	Reliance on Other Experts	22

4.	Property Description and Location	22

5.	Accessibility, Climate, Local Resources, Infrastructure and Physiography	23

6.	History	23

7.	Geological Setting	23

8.	Deposit Types	23

9.	Mineralization	24

10.	Exploration	24

	10.1 Introduction	24
	10.2 General	24
	10.3 Exploration Conducted by Exmibal from 1960 to 1981	25
	10.4 Exploration Conducted by CGN from April 2005	25
	10.5 Service Providers 2005	26
	10.6 Interpretation of Exploration Information	26

11.	Drilling	29

	11.1 Introduction	29
	11.2 Exmibal Drilling and Pitting Prior to 1981	29
	11.3 CGN Drilling from 2005	30
	11.4 Extent of Drilling	33
	11.5 Results	33

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12.	Sampling Method and Approach	33

	12.1 Introduction	33
	12.2 Sampling Methods Prior to 1981	33
	12.3 Sampling Methods from 2005	38

13.	Sample Preparation, Analyses, and Security	39

	13.1 Introduction	39
	13.2 Sample Preparation prior to Dispatch of Samples	39
	13.3 Statement Regarding Sample Preparation	40
	13.4 Dispatch of Samples, Sample Preparation, Assaying and Analytical Procedures	40
	13.5 Quality Control Measures	42
	13.6 Security	44
	13.7 Statement on the Adequacy of Sample Preparation, Security and Analytical Procedures	45

14.	Data Verification	45

	14.1 Introduction	45
	14.2 General	45
	14.3 Historic Database	46
	14.4 Calculated Grade Data	47
	14.5 Bulk Density	48
	14.6 Twin Drilling	50
	14.7 Verification by Snowden	50
	14.8 Statements Regarding Verification	54
	14.9 Reliance on Owner Supplied Information	54

15.	Adjacent Properties	56

16.	Mineral Processing and Metallurgical Testing	56

	16.1 Introduction	56
	16.2 Pyrometallurgical Testwork	59

17.	Mineral Resource and Mineral Reserve Estimates	61

	17.1 Mineral Resource Estimates	61
	17.2 Mineral Reserve Estimates	75

18.	Other Relevant Data and Information	81

	18.1 Mining	81
	18.2 Existing Facilities and Equipment	83
	18.3 Process Plant Description	86
	18.4 Power Supply	94
	18.5 Process and Power Plant Infrastructure, Utilities and Services	105
	18.6 Operations Logistics and Transportation	111
	18.7 Metallurgical Process and Recoverability	118
	18.8 Markets and Sales Agency Agreement	120
	18.9 Environmental Considerations	124
	18.10 Taxes	128
	18.11 Capital and Operating Cost Estimates	130
	18.12 Economic Analysis	154
	18.13 Project Payback	161
	18.14 Project Risks and Opportunities	164
	18.15 Mine Life and Exploration Potential	169
	18.16 Hyrdometallurgical Expansion Preliminary Assessment	170

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19.	Interpretation and Conclusions	170

	19.1 Geology and Mineral Resources	170
	19.2 Mineral Reserve Estimates and Mining	171

20.	Recommendations	172

	20.1 Mineral Resource Evaluation	172
	20.2 Mining	173
	20.3 Process and Power Plant	174

21.	References	174

22.	Dates and Signatures	176

23.	Certificates	177

Tables

Table 1-1: Fenix Mineral Resource Estimates – Effective Date March 31, 2007		13
Table 1-2: Mineral Reserve Estimates – 25 September 2006		14
Table 2-1: Author’s Responsibilities		20
Table 10-1: Summary of CGN Drilling to June 21, 2007		25
Table 10-2: Summary of Material Type Classification		26
Table 10-3: Chemical Classification of Laterite Materials		27
Table 10-4: Chemical Classification of Laterite Materials- Exmibal Samples		27
Table 10-5: Chemical Classification of Laterite Materials- CGN Samples		27
Table 10-6: Chemical Classification of Bedrock Materials		28
Table 11-1: Historic Drilling Methods and Extent		30
Table 12-1: Comparison of Historic Exmibal and CGN Diamond Drilling		36
Table 13-1: Formulae to Calculate SiO₂ from Ni and Fe₂O₃		41
Table 13-2: Formulae to Calculate MgO from Ni and Fe₂O₃		42
Table 14-1: Verification Stages		45
Table 14-2: Large-scale Validation of SiO₂and MgO Estimates		48
Table 14-3: Sources of Density Data		49
Table 14-4: Regression Formulae to Calculate DBD for Drill Samples		49
Table 14-5: Overburden Volume from 2006 Pre-mine Resource Model Compared with Historical Records from Blocks 212-1 and 212-5		53
Table 14-6: Comparison of 2006 Resource Model with Historical Production from Blocks 212-1 and 212-5		55
Table 16-1: Description of Pilot Plant Samples		58
Table 16-2: Metallurgical Results During Production of Ferronickel from Guatemalan Ore		60
Table 17-1: Mineral Resource Estimates of Transition and Saprolite at 0.8% Nickel Cut-off for Selected Areas		62
Table 17-2: Mineral Resource Estimates of Limonite at 1.0 % Ni Equivalent Cut-off (NiEq = Ni + 3Co) for Selected Areas		62
Table 17-3: Mineral Resource Estimates – Other Saprolite Deposits		63
Table 17-4: Mineral Resource Estimates – Other Limonite Deposits		64
Table 17-5: Topcuts Applied to MgO Data Composites		70
Table 17-6: Topcuts Applied to SiO₂ Data Composites		70
Table 17-7: Mineral Reserve Design Criteria		76
Table 17-8: Fenix Project Optimization Parameters		77
Table 17-9: Fenix Mineral Reserve Estimates – 25 September 2006		80

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Table 18-1: Overview of Generating Capacity on the Guatemalan Electrical Grid		95
Table 18-2: Process Plant Electrical Power Demand		95
Table 18-3: Largest Generating Stations on the Guatemalan Electrical Grid		96
Table 18-4: Contracted Capacity Covering the Term of the PPA		97
Table 18-5: Process Plant Turndown Power Demand Requirements		101
Table 18-6: Smelting Furnace Upset Conditions		101
Table 18-7: Fuel Cost as a Function of Cycle on a kW-h Basis		102
Table 18-8: Summary of Boiler/Turbine (B/T) Configuration Options		103
Table 18-9: Major Equipment Gross Ratings of Selected Power Plant Configuration		105
Table 18-10: Non-Process Buildings		106
Table 18-11: Information Systems Software Packages		108
Table 18-12: Telecommunication Systems		109
Table 18-13: Summary of Water Users (excludes Backwash Flows and Other Losses)		110
Table 18-14: Fenix Consumable and Product Transport Requirements		113
Table 18-15: Petroleum coke and Coal Freighter Requirements		115
Table 18-16: Daily Trucking Requirements (Santo Tomas - Fenix)		115
Table 18-17: Trucking Requirements for Backhauling Ferro-nickel		115
Table 18-18: Daily Trucking Requirements for Secondary Consumables		115
Table 18-19: Average Ore Assay, dry wt%		118
Table 18-20: Sources of Nickel Losses		119
Table 18-21: Ramp-up for Each Kiln		119
Table 18-22: Historical Nickel Prices		123
Table 18-23: Forecast Real 2007$ Nickel Prices		123
Table 18-24: Summary of Social and Environmental Management Plans		126
Table 18-25: Additional Royalty Schedule (Post-Payback)		130
Table 18-26: Comparison of Feasibility Study and Updated Capital Costs		130
Table 18-27: Phase 1 Capital Cost Summary by WBS		133
Table 18-28: Phase 2 Capital Cost Summary		134
Table 18-29: Summary of Basis of Estimate		136
Table 18-30: Exchange Rates		138
Table 18-31 Summary of Operating Costs for Feasibility Study and Update		140
Table 18-32: Scope of Operating Cost Estimate by Area		141
Table 18-33: Project Operating Costs – Years 1-30		143
Table 18-34: Project Operating Cost Breakdown by Area – Year 4		144
Table 18-35: Project Operating Costs by Element – Year 4		145
Table 18-36: Project Operating Cost Breakdown by Area – Year 6		147
Table 18-37: Project Operating Costs by Element – Year 6		148
Table 18-38: Power Unit Cost (per MWh)		150
Table 18-39: Power Cost Breakdown for Year 6		151
Table 18-40: Coal Cost Breakdown		152
Table 18-41: Labour Cost for Year 4		152
Table 18-42: Shipping Cost Breakdown (US$/tonne)		153
Table 18-43: Fenix Project Base Case Economic Analysis Summary		155
Table 18-44: Key Cash Flow Model Assumptions		155
Table 18-45: Summary of Production Costs after Brook Hunt		156
Table 18-46: NPV of Fenix Project Net Cash Flows (at US$6.50/lb Ni)		161
Table 18-47: Impact of Ni Price on Payback Period and IRR		161
Table 18-48: Fenix Project Sensitivity to Ni Metal Price		162
Table 18-49: Fenix Project Sensitivity to Site Operating Costs		162
Table 18-50: Fenix Project Sensitivity to Project Capital Cost		162
Table 18-51: Fenix Project Sensitivity to Petroleum coke Price		163
Table 20-1: Recommended Budget		173

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Figures

Figure 4-1: Location of Skye Mineral Licences		23
Figure 10-1: Location of Feasibility Study Deposit Areas		24
Figure 11-1: Location of CGN Feasibility Study Drilling		31
Figure 11-2: Location of Recent and Current CGN Drilling		31
Figure 12-1: Comparison of Nickel Grades Between Co-located Boreholes		35
Figure 12-2: Comparison of Iron Grades Between Co-located Boreholes		35
Figure 12-3: Comparison of Mineralized Thicknesses between Co-located Boreholes		36
Figure 12-4: Comparison of Historic Auger and Twin Core Assays		37
Figure 12-5: Nickel Grade Profiles from Twinned Exmibal and CGN Holes		37
Figure 12-6: Iron Grade Profiles from Twinned Exmibal and CGN Holes		38
Figure 17-1 :Mine Cost Sensitivity Using Updated Feasibility Estimates		78
Figure 17-2: Process Cost Sensitivity Using Updated Feasibility Estimates		78
Figure 17-3: Metal Price Sensitivity Using Updated Feasibility Estimates		79
Figure 17-4: Sensitivity Using Updated Feasibility Estimates On Selected Pit Shell		80
Figure 18-1: Mill Feed Tonnes and Grade		82
Figure 18-2: Process Block Flow Diagram		88
Figure 18-3: Furnace Power Levels vs. Time		91
Figure 18-4: Nickel Laterite Furnace Sizing Parameters		92
Figure 18-5: Falcondo Furnace Power With and Without Compensation		107
Figure 18-6: Bulk Material Transport Route		114
Figure 18-7: Nickel Production		120
Figure 18-8: Forecast Nickel Cost Curve		122
Figure 18-9: Project Operating Costs and Ore Grade – Years 1-30		142
Figure 18-10: Operating Cost Distribution by Area – Year 4		144
Figure 18-11: Operating Cost Distribution by Element – Year 4		146
Figure 18-12: Operating Cost Distribution by Area – Year 6		147
Figure 18-13: Operating Cost Distribution by Element - Year 6		149
Figure 18-14: Fenix Project Capital Cost Expenditure Distribution Profile		157
Figure 18-15: CGN Annual Net Sales Revenue (at $6.50/lb Ni and $0.20/lb Ni Fe Credit)		158
Figure 18-16: Total Income Tax Payable by CGN (at US$6.50/lb Ni)		160
Figure 18-17: Net Cash Flows for CGN (at US$6.50/lb Ni)		160
Figure 18-18: Fenix Project Relative Sensitivity to Variable Changes		163
Figure 18-19: Fenix Project Relative Sensitivity to Fuel Supplies		164

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IMPORTANT NOTICE

General

This report was prepared by the qualified persons listed in Table 2-1. Each qualified person (QP) assumes responsibility for those sections or areas of this report that are referenced opposite their name in Table 2-1. None of the QPs, however, accepts any responsibility or liability for the sections or areas of this report that were prepared by other QPs.

This report was prepared to allow Skye to reach informed decisions respecting the development of the Fenix Project. Except for the purposes legislated under provincial securities law, (a) any use of this report by any third party is at that party’s sole risk, and none of the QPs (nor any of the companies for whom they work) shall have any liability to any third party for any such use for any reason whatsoever, including negligence, and (b) each of the QPs hereby disclaims responsibility for any indirect or consequential loss arising from any use of this report or the information contained herein.

This report is intended to be read as a whole, and sections should not be read or relied upon out of context. This report contains the expression of the professional opinions of the QPs, based upon information available at the time of preparation. The quality of the information, conclusions and estimates contained herein is consistent with the intended level of accuracy as set out in this report, as well as the circumstances and constraints under which the report was prepared which are also set out herein.

As permitted by Item 5 of Form 43-101F1, the QPs have, in the preparation of this report, relied upon certain reports, opinions and statements of certain experts. These reports, opinions and statements, the makers of each such report, opinion or statement and the extent of reliance is described in Section 3 of this report. Each of the QPs hereby disclaims liability for such reports, opinions and statement to the extent that they have been relied upon in the preparation of this report, as described in Section 3.

As permitted by Item 16 of Form 43-101F1, the QPs have, in the preparation of this Report, relied upon certain data provided to the QPs by Skye and certain other parties. The relevant data and the extent of reliance upon such data are described in Section 14 of this Report.

This Technical Report is a revised version of a report that was initially filed on September 15, 2007 and is being filed at the request of securities regulators. The only change to the Technical report that was filed on September 15, 2007 is the deletion of sections relating to the possible hydrometallurgical expansion of the Fenix Project. The reasons for this deletion are explained in Section 1.6 of the report.

Readers should be aware that, since the effective date of the report (September 15^th, 2007), there have been changes to both the Project and to the pricing and availability of Project inputs and outputs that would require changes to the Report if an update was prepared. However at the instruction of the issuer an update of the report has not been prepared.

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Hydrometallurgical Expansion Preliminary Assessment

The Technical Report filed by HMI dated September 15, 2007, together with an Addendum Report dated October 25, 2007, referred to and summarized a Hydrometallurgical Expansion Preliminary Assessment issued in October 2006 (the “Preliminary Assessment”). Since October 2007, HMI has conducted additional test work including pilot plant testing of a High Pressure Acid Leaching process, as reported in HMI’s press release dated February 19, 2008. Additional testing and evaluation of this process may be conducted, however an update to the capital costs and economic assumptions in the Preliminary Assessment has not been undertaken and HMI does not intend to update this information as HMI no longer considers the Hydrometallurgical Expansion material to HMI. Accordingly, the information and conclusions in the Preliminary Assessment should no longer be relied upon with respect to the development of the Fenix Project.

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1.	Summary

1.1

General

The Fenix nickel laterite deposits are located near El Estor in the Department of Izabal, Guatemala. The mineral rights to most of the deposits are held under a 25 year renewable exploitation licence and the remainder are held under an exploration licence, both by Compañia Guatemalteca de Níquel S.A. (CGN). CGN is a Guatemalan company, the shares of which are owned approximately 92.4% by Skye Resources Guatemala (B.V.I.) Inc. (a wholly-owned indirect subsidiary of Skye Resources Inc.), and 7.6% by the Guatemalan Government.

CGN was formerly known as Exmibal. From 1960 to 1981 the deposits were evaluated by Exmibal and a mine and process plant for treating saprolite was established in 1977. This operation continued until high fuel costs and low nickel prices caused the process to become uneconomic.

Exmibal’s name changed to CGN in 2005 and a feasibility study for a 50 million pound per annum ferro-nickel project at Fenix using conventional smelting technology commenced.

The Fenix Feasibility Study was completed on October 16, 2006 (the “October 2006 Feasibility Study”), and the related Technical Report was issued in September 2006 and filed at www.sedar.com on November 17, 2006 (the “November 2006 Technical Report”). The Report has been updated to incorporate the following revisions:

Exploration up to June, 2007 and Revised Mineral Resource Estimates

This report describes the results of exploration work up to June 2007 and presents up-dated mineral resource estimates, which now include new areas not previously estimated. According to the Qualified Person the changes to the mineral resource estimates that have resulted are not material to the Fenix Project.

Revised Power Supply Strategy

The October 2006 Feasibility Study contemplated the construction of a new dedicated power plant concurrently with the refurbishment and upgrade of the process plant. Since that time expected lead times for key components of the new power plant have lengthened significantly. CGN is in the process of finalizing an arrangement for the interim supply of power under a power purchase agreement (PPA). The availability of interim power supply removes potential constraints on the ramp-up of the process plant, and reduces the project’s dependency on on-time completion of the new power plant. However, in order to provide Fenix with a long-term power supply source that is not dependent on Heavy Fuel Oil (HFO), a new power plant is intended to be constructed once the process plant is in service. This deferral of the new power plant is made possible by the acquisition of interim power supply under the PPA, and reduces the initial capital requirements of Fenix.

Trucked Transport of Bulk Materials

The method for transporting bulk materials (coal and petroleum coke), proposed in the October 2006 Feasibility Study, consisted of trucking coal/petroleum coke from a trans-shipment terminal at the port of Santo Tomas to a ferry terminal near Mariscos, a town located on the south shore of Lake Izabal. Barges would ferry the truck trailers from Mariscos to the Fenix plant site.

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The truck-barge transport system has been replaced with an all trucking system, due to perceived environmental and socio-economic risks associated with lake transport. Coal/petroleum coke will be trucked from the Santo Tomas trans-shipment terminal to the plant site, a distance of 130 km.

Hatch Mott MacDonald (HMM) completed a site audit of the existing public highways (CA9, CA13 and 7E) between Santo Tomas and the plant site. HMM developed a preliminary analysis (scoping study level) of upgrade requirements to the pavement structures, road alignment and bridges. CGN is negotiating an agreement with the Guatemalan government whereby, in return for the government upgrading the highways as per the HMM recommendations, CGN will pay for the maintenance of Highway 7E from Panzos (west of the Fenix plant) to Fronteras (junction 7E and CA13) and the maintenance of a gravel highway from 7E to Cahabon, during the operational life of the Fenix facility. In addition, the government will divert the funds it would normally have spent on maintaining the sections of 7E that CGN will be maintaining, to improved maintenance on those parts of highways CA13 and CA9 that the Fenix operation will be using.

Updated Economic Analysis

The economic analysis has been updated to include

•

Changes in the project scope and cash flow as indicated in items 1 and 2 above

•

Updated capital and operating cost data (July 2007 base date)

•

Updated forecast nickel pricing

Basic Engineering

Hatch has completed Basic Engineering for the project and has issued revised capital and operating cost estimates, incorporating the above noted changes, in addition to the on-going changes made during design development. Certain other minor changes, not considered material by the authors of the report, have not been described.

1.2

Geology, Deposit Evaluation and Mineral Resource Estimates

The nickel deposits are lateritic weathering profiles that have formed on peridotites thrust into place during the early Tertiary. The most important deposits are largely confined to NE or SE trending terraces and spurs on the south flanks of the Sierra de Santa Cruz at elevations between 370 and 800 metres. Total topographic relief in the area is 1,000 metres.

Previous exploration involved traditional methods of mapping and sampling by way of manual and power drills, test pits and channel sampling. In 2005 CGN undertook a diamond core drilling program to evaluate known laterite mineral resources in the 212, 213, 215, 217 and 251 areas; and to validate the historic Exmibal sample database.

Dr. Paul Golightly of Golightly Geoscience verified the historic data and compiled a located sample database of nickel, cobalt, iron, magnesia, silica and dry density. CGN’s core drilling assays were used to establish factors to correct an iron bias in the historic sample database, and to derive calculated values for magnesia, silica and dry density as these variables were not present for the entire historic database.

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CGN’s core drilling was conducted in two phases:

•

A program of twin drilling in areas of historic drilling at 212, 213, 217 where historic drill spacing is generally 25 m or less.

•

A program of in-fill drilling in 215, 217 and 251 to achieve a general drill spacing of 50 m or less.

Dr. Golightly’s opinion, the core sampling, sample preparation, analytical procedures and security for the CGN programs are industry standard. The procedures for sampling, compiling and security of the historic data are acceptable and there has been sufficient checking from original sources to confirm that the historic database is acceptable for Mineral Resource estimation.

Snowden compiled located sample databases from the historic data and recent CGN core drilling programs; and reviewed the material classifications of Golightly Geoscience that were derived from nickel and iron assays. In Snowden’s opinion the compiled data are acceptable for Mineral Resource estimation.

Snowden prepared limonite, transition and saprolite Mineral Resource estimates for Ni, Co, Fe₂O₃, MgO, SiO₂ and dry density for areas 212, 213, 215, 217 and 251 by first constructing 3-D wireframes of the laterite layers and then interpolating the grade variables by ordinary kriging into conventional 3-D block models in accordance with CIM guidelines. The Mineral Resource estimates were validated by alternative interpolation methods and classified as Measured, Indicated and Inferred categories consistent with the requirements of CIM and NI 43-101. The classification scheme took borehole spacing, mineralization continuity and relative error into account.

The Snowden estimates for two blocks in area 212 were compared with mine production and reconciled well on a global basis. The within-pit portion of the resource model reconciled in terms of overburden volume and mineralized material grade and tonnes to within 8% of the historical production data on a global basis.

Information from historic mining and the latest core drilling indicates that the full depth extent of the laterite profile and short scale variation can be misinterpreted from drilling data alone. This uncertainty is addressed in the Mineral Resource classification scheme, whereby measured category in saprolite is achieved by drilling at spacings of 35 m or less, and spacings of 35 m to 75 m are required for Indicated saprolite mineral resources.

Golightly Geoscience reported Mineral Resource estimates for the other deposits that occur in CGN’s licences. The estimates were originally prepared for all areas using Inco’s 2-D LES method however the LES estimates in five of the areas are superseded by the Snowden 3-D estimates. The LES resource estimates are realistic in their stated purpose of defining a resource for a Preliminary Assessment and these deposits will require re-estimation in 3-D format to enable mine planning to proceed.

Mineral resource estimates at the effective date of 25 September 2006 are provided in Table 1-1 for saprolite and limonite. These estimates have been modified from a disclosure of 4 July 2006 to reflect lower cut-off grades for resource reporting.

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Programs of additional drilling, both twin and in-fill, followed by updated mineral resource estimates are recommended in the following areas:

•

Area 218 deposits.

•

Montúfar property, the Cristina area 221 & 222 deposits.

•

The lower terraces of area 215, south of terrace 215-1. In-fill drilling at 50 m spacing is recommended, 260 (Amate) and 219.

A budget of approximately $US16M is estimated for these programs.

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Table 1-1: Fenix Mineral Resource Estimates – Effective Date March 31, 2007

(Mineral Resources are inclusive of Mineral Reserves and they must not be added together)


	Feasibility Study Estimates¹ - Areas 212, 213, 215, 217, 251			Other Area Estimates²
	Tonnes x 10⁶	Nickel (%)	Cobalt (%)	Tonnes x 10⁶	Nickel (%)	Cobalt (%)
Saprolite†
Measured	24.67	1.54	*	8.7	1.79	*
Indicated	36.78	1.45	*	27.1	1.82	*
Measured & Indicated	61.45	1.49	*	35.8	1.81	*
Inferred	43.01	1.25	*	48.2	1.64	*
Limonite
Measured	44.32	1.12	0.110	2.30	1.10	0.093
Indicated	2.21	1.02	0.098	10.30	1.20	0.101
Measured & Indicated	46.53	1.11	0.110	12.60	1.18	0.100
Inferred†	14.70	1.01	0.083	32.80	1.15	0.095

Notes:

¹	0.80% Ni cut-off saprolite; 1.00% Ni equivalent cut-off limonite where Ni equivalent = Ni + 3 Co. Mineral Resource Estimates also reported at 1.50% Ni cut-off for saprolite; 1.25% Ni equivalent cut-off for limonite in Snowden (2006).

²	1.60% Ni cut-off saprolite; 1.00% Ni equivalent cut-off limonite where Ni equivalent = Ni + 3 Co.

*	Not reported

†	Saprolite includes transition material

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1.3

Mineral Reserve Estimates

Mineral reserve estimates were based on the mineral resource estimates at the effective date of July 4, 2006 for five deposits and have been appropriately modified to account for:

•

Dilution, contamination, and ore losses.

•

Metallurgical process recovery.

•

Economic evaluation by means of resource optimization and the scheduling of phased pits in order to meet processing requirements and ore chemistry constraints.

A pit shell containing approximately 44 million tonnes of saprolite and transitional saprolite (transition) material grading 1.60% nickel was selected using discounted cash flow analysis on diluted resource models optimized in Whittle Four-X (Whittle) software.

The variables in the original optimization were based on metal prices, process costs, metallurgical recoveries and other criteria established in the October 2006 Feasibility Study. Both nickel and iron were considered as revenue generating in the optimization. A nickel price of $US5.00/lb was assumed and for every pound of nickel in ferro-nickel produced an additional $US0.20/lb was added for iron, for a combined metal price of $US5.20/lb. Average mining costs of $US3.10/tonne used in the optimizations were benchmarked from existing producing nickel laterite operations of similar scale.

A new optimization was completed using an updated mining cost estimate of $3.15/tonne, a processing cost estimate of $68.04/tonne, and a combined ferro-nickel metal price estimate of $6.70/lb. The resultant shell selected was within 2% of the October 2006 Feasibility Study as such Snowden determined that no changes to the life of mine schedule and the Mineral reserve estimate would be required.

Sensitivity analyses on the optimization were, run on metal price, milling costs, and mining costs, indicating that the pit size is insensitive to these variables largely due to the shallowness of the deposit. The project’s value, however, is most sensitive to metal prices and least sensitive to mining costs.

Mineral reserve estimates at the effective date of 25 September 2006 are summarized in Table 1-2 for saprolite and transition.

Table 1-2: Mineral Reserve Estimates – 25 September 2006


Mineral Reserve	Tonnes (millions)	%Ni	Contained Ni (Tonnes)
Proven	8.7	1.81	157,000
Probable	32.7	1.58	516,000
Proven and Probable	41.4	1.63	673,000

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1.4

Mining and Metallurgical Plant Operations

At steady state operation the mine will provide 1,464,000 tonnes of saprolite and transition ore annually to the process plant. Total average annual mine production is approximately 3,100,000 tonnes at an average stripping ratio of 1.1:1. The minimum selective mining unit is 2.5 m high. High grade and low grade stockpiles will be created in order to optimize nickel grade delivered to the plant. The general mining sequence will begin with removal and storage of soil and organic cover for future rehabilitation of the mine areas. Waste cover and limonite will be stripped by excavator or track dozer and placed in either temporary external or in-pit storage areas. Limonite with potential for future processing will be stripped and stored separately from the waste. Mining will begin at the top of the hill and proceed to lower elevations. Ore is mined and loaded into off-road haul trucks for transportation to the plant.

The metallurgical plant, which uses conventional smelting technology, involves refurbishing and expanding the existing plant which produced nickel matte from 1977 to 1980. Modifications to the plant include:

•

Replacing the existing oil-fired dryer with a larger coal-fired unit.

•

Converting the existing oil-fired kiln to coal-firing.

•

Installing a second, coal-fired, calcination and reduction kiln parallel to the existing kiln.

•

Installation of a new ladle refinery to produce a 35% Ni ferro-nickel product.

•

Upgrading the existing electric furnace to operate at 90 MW in shielded-arc mode.

•

Installation of a 230 kV transmission line from the nearest 230 kV Guatemalan grid connection (located at Tactic) to the Fenix plant site, a distance of approximately 160 km. The transmission line will be constructed, owned and operated by a Guatemalan energy supplier who is currently negotiating a Power Purchase Agreement (PPA) and Transmission Tolling Agreement (TTA) with CGN to provide power for the first five (5) years of operation of the Fenix project, commencing in October, 2009.

•

Installation of a 230kV/34.5kV step-down sub-station at Fenix, including the necessary compensation equipment to ensure that the Fenix load does not de-stabilize the Guatemalan grid.

•

In sufficient time for operation in October, 2014, construction of a new 150 MW petroleum coke-fired boiler and a new 90 MW steam turbine generator which, in combination with the existing generator, will furnish the plant with 150 MW of electrical power.

When the new 150MW power plant is operating nearly all fuel requirements will be supplied by coal and pet-coke, thereby essentially eliminating the prior sensitivity of operating costs to the oil price.

The processing line of the plant, shown as a process block diagram in Figure 18-2, consists of one dryer, two reduction kilns, one electric furnace and a ferro-nickel refinery. The capital project implementation schedule consists of two phases.

Phase 1 (July 2007 – October 2009):

•

Refurbishment of the existing facility including, the process plant, upgrading the electric furnace and converting the existing kiln to coal-firing.

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•

Installation of the new process equipment including a new dryer, ladle refinery, coal processing plant and the second kiln.

•

Installing the 230 kV transmission line (from the Tactic substation to the Fenix plant site) and 230 kV/34.5 substation.

•

Upgrading public roads and bridges between the port of Santo Tomas and the Fenix site.

The facility is scheduled to begin operation in October 2009 with electric power provided via the 230 kV transmission line.

Phase 2 (post October 2009)

Construction of a new on-site power plant will begin after the process plant begins production. The new power plant is expected to be ready for operation by October 2014 and will include:

•

A new 150 MW circulating fluid bed (CFB) boiler. The boiler will be designed to operate using petroleum coke (petcoke).

•

A new 90 MW steam turbine generator (STG).

•

Refurbishment of the existing 60 MW STG.

•

New ancillary facilities including cooling water system, petroleum coke handling system, limestone handling system, etc.

Production over the 30 year project life is estimated to be 1.3 billion pounds of nickel. The average production for the first 20 years after ramp-up (years 3 to 22) is 48.5 million pounds per year of nickel contained in ferro-nickel.

1.5

Economic Analysis

The Phase 1 capital cost is estimated to be $640 million, with an intended level of accuracy of -5%, +12%. This total includes direct costs of $355 million, indirect costs of $157 million, a contingency of $65 million and owner’s costs of $63 million. The Phase 2 capital cost of the power plant, including direct, indirect, contingency and Owner’s costs, is $344 million with an intended level of accuracy of -10%, +20%. (The capital cost estimates have a base date of July 2007, and no allowance has been included for price escalation or currency fluctuations.)

Based on a nickel price of $6.50 per lb, an iron credit of $0.20 per lb Ni and royalties and other costs of $0.49 per lb, the ferro-nickel project’s IRR is estimated at 14.3% and it’s NPV at $490.6 million with an 8% discount rate, and $275.0 million with a 10% discount rate.

Cash operating costs during the first five years of the project have been estimated to be $3.47 per lb Ni, (with an intended level of accuracy of +/-10%). After the CFB power plant is installed at the end of year 5, cash operating costs for the remaining 15 years (up to Year 20) of the project are expected to decline to $2.34 per lb Ni. Operating costs have been calculated using estimates of long term pet-coke, coal and oil prices.

The study shows, that of all the key parameters, the project’s economics are most sensitive to nickel prices. At an average nickel price of $8.00 per lb, the ferro-nickel project’s IRR climbs to 19.8% and its NPV rises to $977.2 million at 8% or $666.7 million at 10%.

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1.6

Hydrometallurgical Processing of Limonite

2.	Introduction

2.1

General

The Fenix nickel laterite deposits are located near El Estor in the Department of Izabal, Guatemala. The mineral rights to most of the deposits are held under a 25 year renewable exploitation licence by Compañia Guatemalteca de Níquel S.A. (CGN). CGN is a Guatemalan company, the shares of which are owned approximately 92.4% by Skye Resources Guatemala (B.V.I.) Inc. (a wholly-owned subsidiary of Skye Resources Inc.), and 7.6% by the Guatemalan Government.

Skye Resources Inc. (Skye) is based in Vancouver, British Columbia and listed on the Toronto Stock Exchange (TSX).

CGN is the new name of Exploraciones y Explotaciones Mineras Izabal, S.A. (Exmibal), which mined and processed the La Gloria deposits (Areas 212 and 213) near El Estor from 1977 until 1980.

2.2

Geology, Deposit Evaluation and Mineral Resource Estimates

The purpose of these sections of the report is to disclose, on behalf of Skye, the results of the exploration drilling program undertaken by Skye/CGN and the subsequent mineral resource and mineral reserve estimates for several deposits that were undertaken by Snowden.

The cut-off date for the drilling used in the Snowden mineral reserve estimates is 31 January, 2006.

Mineral resource estimates were disclosed by Skye in a press release dated 4 July 2006, which disclosure gave rise to the requirement to file a Technical Report Mineral Resource Estimate, Fenix Project, Izabal, Guatemala (Snowden 2006). Subsequently Snowden has completed a mining study on several of the Fenix deposits as part of Hatch’s feasibility study and, as a consequence, mineral reserves are estimated for several deposits. Incremental increases in mineral resources reflecting new drilling carried out in Area 217 in 2006 and a re-evaluation of limonite mineral resources outside of the Fenix project area are also incorporated the mineral resource inventory.

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2.2.1

Background

Historic Exmibal and new sampling data obtained by CGN were used in the July 2006 resource estimates. CGN conducted drilling and sampling activities in 2005-06 under the direction of Colin McKenzie P.Geo., Vice-President of Exploration for Skye. The drilling program followed the general recommendations of Golightly Geoscience as described in Hatch 2005.

Dr. Paul Golightly, P.Geo., President of Golightly Geoscience is the independent Qualified Person, as defined under National Instrument 43-101 (NI 43-101) for the quality control and verification of the Fenix Project geological data. During 2005-06 Golightly Geoscience compiled and verified the new data from CGN’s drilling program as well as the historic Exmibal sampling data.

Skye/CGN’s feasibility study drilling program was conducted in several phases:

•

Drilling of diamond core holes in close proximity to earlier Exmibal auger holes to confirm the historic results (twin hole drilling).

•

Drilling of close-spaced diamond core holes in specific areas to understand the local grade and thickness variability of the laterite profile (geostatistical crosses).

•

Pattern drilling at spacings of 50 m to in-fill and extend the exploration conducted by Exmibal (in-fill drilling).

In May of 2005, Skye engaged Snowden to undertake an independent resource estimate on several of the deposits on completion of the feasibility study drilling program. The deposits studied by Snowden are referenced as Areas 212 and 213 (the former La Gloria Mine), 215, 217 and 251. The estimates for Area 212 include a small amount of data from the adjoining Area 216.

At the completion of the phased drilling programs, the field data was compiled and validated by Golightly Geoscience and Snowden personnel. Geological classification of the historic and recent CGN samples was completed by Golightly Geoscience and reviewed by Snowden personnel, who subsequently generated geological interpretations and 3-D layer models of the deposits. The July 2006 resource estimates were undertaken by Snowden under the technical supervision of Andrew Ross.

2.3

Metallurgical Testwork

Section 16 provides a discussion of the history of the pyrometallurgical operation at CGN and its relevance to the processing plant proposed in the feasibility study.

2.4

Mineral Reserve Estimates

Snowden evaluated the mineral resource block models by means of pit optimization software which takes into account a set of financial and metallurgical criteria to produce a series of pit shells. The most favourable pit shell in terms of financial return was scheduled using MineMAX software to provide 24 years of plant feed as well as a stockpile and backfilling strategy. A second larger stage pit was scheduled to extend plant feed to 30 years. The mineral reserve estimates were based on the measured and indicated mineral resources for five deposit areas. The estimates are tabulated in the proven and probable categories by Snowden under the technical direction of Dick Matthews.

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2.5

Mine and Process Plant Feasibility Study

2.5.1

Mining Study

Skye appointed Snowden to provide resource evaluation, mining engineering and mine planning services to support the engineering phase of the feasibility report for the Fenix nickel project.

In addition to mineral resource and reserve estimates, the key components of Snowden’s tasks were:

•

Review mining methods and dilution as laid out in the preliminary assessment (Hatch, 2005).

•

Create a detailed life of mine schedule.

•

Estimate mine capital and operating costs to feasibility levels of accuracy.

•

Undertake alternative mine transport studies.

•

Design a limestone quarry operation and estimate capital and operating costs.

Other studies under Snowden’s supervision but completed by BGC Engineering (BGC) and Klohn Crippen Berger (Klohn) concern the mine water and sediment control, reclamation and rehabilitation (BGC) and emissions reports (Klohn); respectively.

2.5.2

Process Plant, Power Plant and Infrastructure

Skye contracted Hatch to engineer and cost an upgraded and expanded process facility at Fenix, together with the power, infrastructure and transportation and environmental control facilities and logistics required to sustain a minimum 30 year operation while meeting Guatemalan and World Bank environmental guidelines. Hatch’s role included coordination of the overall Feasibility Study document. Subsequently Hatch completed basic engineering for the process plant and site services. Specific tasks included:

•

Based on resource and mining plan data provided by the resource and mining consultants and on the environmental specifications provided by the environmental consultant, generate process flowsheets, heat and mass balances for the process and power plants, infrastructure and transportation facilities to a standard sufficient to be able to identify, specify and solicit quotations for the equipment required.

•

Manage the evaluation of the existing equipment and facilities to determine their fitness for use in the renovated and expanded plant.

•

Identify the additional equipment required, solicit and evaluate vendor quotations with regards to major equipment and supplies.

•

Generate such engineering drawings and specifications as are necessary to allow project costing, such drawings to include general layout, process, instrumentation, electrical, structural and civil.

•

Develop project implementation, staffing, and ramp up plans.

•

Develop, with input from equipment vendors and Skye’s other consultants, capital cost estimates for the Phase 1 and Phase 2 projects.

•

Develop, with input from Skye’s other consultants, a +/-10% operating cost estimate.

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•

Develop a project financial model incorporating the capital and operating costs so developed, tax regime and contractual information provided by Skye, market assumptions provided by the marketing consultant and working together with Skye’s VP Finance generate project financial performance data.

•

Complete an audit of the public roads CA9, CA13 and 7E between Santo Tomas and the Fenix plant site and the development of a preliminary cost estimate to upgrade the roads/bridges to meet Fenix operating and capital project requirements.

•

Review electrical grid analysis work completed by an Independent Power Producer and develop a preliminary design and cost estimate for a 230/34.5 kV Fenix substation.

2.6

Qualified Persons and Site Visits

The authors of this Technical Report are independent Qualified Persons as defined by NI 43-101.

•

John B. Scott has had an association with the Fenix Project since early 2007 and visited the site on several occasions. Since that time he has managed the engineering work carried out by Hatch and others.

•

Colin B. McKenzie has had an association with the Fenix Project since early 2002 and visited the site on several occasions. He has managed the exploration programs and the mining studies since that time.

Table 2-1: Author’s Responsibilities


Report Section	Subsection	Q.P.	Input from Others
1: Summary	1.1	J.B. Scott
	1.2	C.B.McKenzie	Golightly, Snowden
	1.3	C.B.McKenzie	Snowden
	1.4 & 1.5	J.B. Scott	Hatch
	1.6
2: Introduction	2.1	J.B. Scott	Hatch
	2.2	J.B. Scott	Hatch
	2.3	J.B. Scott	Hatch
	2.4	J.B. Scott	Hatch
	2.5.1	J.B. Scott	Hatch
	2.5.2	J.B. Scott	Hatch
3: Reliance on Other Experts	All	J.B. Scott	Input from Skye
4: Property Description and Location	All	C.B.McKenzie
5: Accessibility, Climate, Local Resources, Infrastructure and Physiography	All	C.B.McKenzie
6: History	All	C.B.McKenzie
7: Geological Setting	All	C.B.McKenzie
8: Deposit Types	All	C.B.McKenzie
9: Mineralization	All	C.B.McKenzie
10: Exploration	All	C.B.McKenzie	Golightly
11: Drilling	All	C.B.McKenzie	Golightly
12: Sampling Method and Approach	All	C.B.McKenzie	Golightly
13: Sample Preparation, Analyses and Security	All	C.B.McKenzie	Golightly

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Report Section	Subsection	Q.P.	Input from Others
14: Data Verification	14.1 – 14.6	C.B.McKenzie	Golightly
	14.7	C.B.McKenzie	Snowden
	14.8	C.B.McKenzie	Golightly, Snowden
	14.9	C.B.McKenzie	Input from Skye
15: Adjacent Properties	All	C.B.McKenzie	Hatch
16: Mineral Processing and Metallurgical Testing	16.1	C.B.McKenzie	Snowden
	16.2	C.B.McKenzie	Hatch
17: Mineral Resource and Mineral Reserve Estimates	17.1	C.B.McKenzie	Snowden
	17.2	C.B.McKenzie	Snowden
18: Other Relevant Data and Information	18.1	C.B.McKenzie	Snowden
	18.2 – 18.7	J.B. Scott	Hatch
	(except 18.3.1)		18.4: Duke Energy
			18.6: Hatch Mott MacDonald
	18.3.1	J.B. Scott	Trow
	18.8	N/A	Gander and CRU Strategies
	18.9.1	N/A	Sosa and Soto
	18.9.2	C.B.McKenzie	BGC Engineering Inc.
	18.9.3	N/A	KCBL
	18.10	N/A	Deloitte, Sosa and Soto
	18.11.1 Capital Costs	J.B. Scott	• Snowden (mine costs) • Hatch • Skye (owner’s costs)
	18.11.2 Operating Costs	J.B. Scott	• Snowden (mine costs) • Hatch • Pace Global (coal/ petroleum coke and HFO prices) • Skye (heavy fuel oil/staff salary levels) • Trans Mar (transport costs)
	18.12 & 18.13 Economic Analysis and Project Payback	J.B. Scott	• Deloitte (taxation) • Hatch • Input from others as per other sections (i.e. Owner’s costs – Skye, Metal prices – CRU Strategies , Mine costs - Snowden
	18.14	J.B. Scott	Hatch
			Skye (opportunities)
	18.15	C.B.McKenzie	Golightly
19: Interpretations and Conclusions	19.1 & 19.2	C.B.McKenzie	Snowden
20: Recommendations	20.1	C.B.McKenzie	Snowden
	20.2	C.B.McKenzie	Snowden
	20.3	J.B. Scott	Hatch
21: References		N/A
22: Dates and Signatures		N/A
23: Certificates		N/A

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3.	Reliance on Other Experts

In preparation of this Report, the following reports and opinions of third party experts have been relied upon without independent verification (and neither Hatch nor any of the other contributors takes any responsibility for the accuracy or completeness of the information set out in such reports or opinions):

(a)

Hatch has provided engineering services for the Fenix project under a series of contracts with Skye, including scoping studies, feasibility study, Basic Engineering, Detailed Engineering and EP (Engineering and Procurment) and CM (Construction Management).

(b)

Gander Consulting LLC (sub-contracted to Skye) provided long term estimates of the nickel price relevant to the ferro-nickel product for the September 2006 Technical Report. CRU Strategies Limited (sub-contracted to Skye) provided updated long term estimates of the nickel price for the September 2007 Technical Report (Section 18.8.1).

(c)

Klohn Crippen Berger Ltd. carried out extensive environmental baseline programs and other environmental studies and analyses to evaluate project environmental and socioeconomic impacts. This work was used as the basis for designing remediation and reclamation of the project facilities (Section 18.9.3). In addition, BGC Engineering Inc. (“BGC” - sub-contracted to Snowden) provided design and costing of surface water run-off control systems for the mine (Section 18.9.2).

(d)

Trans Mar Pacific, Inc. (“Trans Mar” - sub-contracted to Hatch) provided the costing of maritime consumables and product shipping systems (Section 18.6).

(e)

Pace Global Energy Services LLC (“Pace Global” - sub-contracted to Skye) provided market and price outlooks for coal, petroleum coke and heavy fuel oil and developed procurement strategies for these supplies (Section 18.11.2).

(f)

Deloitte Touche, Guatemala provided advice on Guatemalan taxes affecting the project (Section 18.10).

(g)

Marsh Ltd. provided quotations for insurance coverage incorporated with estimates of owner’s costs in the initial project capital (Section 18.11.1.1.4).

(h)

Information respecting Guatemalan laws and regulations, title to the Fenix property, local licenses and royalties, agreements with Inco respecting the project and Skye’s ownership of CGN have been provided by Skye’s Guatemalan counsel, A.D. Sosa & Soto, S.C.

(i)

An Independent Power Producer (IPP) operating in Guatemala completed an interconnection study of the Guatemalan electric distribution system to assess whether the grid is capability of meeting the Fenix load requirements. This IPP also provided construction execution information concerning the installation of a 230 kV transmission line from the Tactic substation to the Fenix plant site.

4.	Property Description and Location

Information in this section has been previously disclosed in a Technical Report entitled Mineral Resource Estimate, Fenix Project, Izabal, Guatemala dated 4 July 2006 (Snowden 2006).

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LOGO

Figure 4-1: Location of Skye Mineral Licences

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

Information in this section has been previously disclosed in a Technical Report entitled Mineral Resource Estimate, Fenix Project, Izabal, Guatemala dated 4 July 2006 (Snowden 2006).

6.	History

Information in this section has been previously disclosed in a Technical Report entitled Mineral Resource Estimate, Fenix Project, Izabal, Guatemala dated 4 July 2006 (Snowden 2006).

7.	Geological Setting

Information in this section has been previously disclosed in a Technical Report entitled Mineral Resource Estimate, Fenix Project, Izabal, Guatemala dated 4 July 2006 (Snowden 2006).

8.	Deposit Types

Information in this section has been previously disclosed in a Technical Report entitled Mineral Resource Estimate, Fenix Project, Izabal, Guatemala dated 4 July 2006 (Snowden 2006).

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9.	Mineralization

Information in this section has been previously disclosed in a Technical Report entitled Mineral Resource Estimate, Fenix Project, Izabal, Guatemala dated 4 July 2006 (Snowden 2006).

10.

Exploration

10.1

Introduction

Much of the information presented in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included for completeness. The ongoing exploration program since 31 January 2006 is also summarized.

10.2

General

Exploration of the Fenix nickel laterite deposits has involved traditional methods of mapping and sampling by way of manual and power drills, test pits and channel sampling.

The work conducted by, or on behalf of Skye, is identified in Sections 10.4 and 10.5 below and areas explored for the feasibility study are shown in Figure 10-1.

LOGO

Figure 10-1: Location of Feasibility Study Deposit Areas

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10.3

Exploration Conducted by Exmibal from 1960 to 1981

Reconnaissance evaluation involved auger drilling on a 200 to 500 metre grid followed by successive stages of refinement by 100 to 50 metre grid drilling and pitting where encouragement warranted. During the process the control of the deposits by topographic terraces and spurs was recognized and used to guide exploration.

Areas included in Exmibal’s original 20 year plan in the La Gloria area of the Fenix Exploitation Licence and two areas in the Niquegua Montúfar II licence are largely drilled at 25 to 50 metre spacing. Development drilling on a 25 metre grid began in the mine-site Area 212 at La Gloria and on the Cristina deposit, Area 221, in the Montúfar licence in 1967.

A large area over the present pit at 212 was drilled at 12.5 metre spacing to provide more detail for mine planning and overburden stripping control. The mine-site in Area 212 has two small patches and a 100 by 100 metre patch with boreholes at six metre intervals as well as a cross pattern, two lines 150 metres in length, drilled at five metre spacing or locally closer.

Sampling of pits is discussed in Sections 11 and 12.

10.4

Exploration Conducted by CGN from April 2005

From April 2005 to January 2006, CGN’s program involved 1,463 HQ diamond core boreholes totalling 34,146 metres and providing approximately 27,700 assay samples. Additional drilling has continued to June 2007 and is compiled with the feasibility study drilling (Table 10-1). The drilling was done under contract. The purpose of the exploration was:

•

To provide a confirmation of historic data and multi-element chemistry in areas of historic 25 x 25 m grid drilling. These 116 “twin” holes totalling 4,111 metres were placed within 5 metres of the original borehole collar and averaged about 2 metres distance from representative historic borehole collars.

•

To in-fill and upgrade the accuracy of the resource estimates to at least a 50 m grid spacing in deposits 215-1, 217-1 & 2 and 251. Five cross patterns of 12.5 -18 m spaced holes were drilled in three of these deposits.

The holes include 10 re-drilled due to poor recovery.

In addition four holes were drilled at Chichipate to investigate the quality of limestone, intended for use in the new on-site power plant.

Table 10-1: Summary of CGN Drilling to June 21, 2007


Area	Holes	Length	Purpose	Date
212	65	2,466	Twin	2005
213	13	300	Twin	2005
217	38	1345	Twin	2005
215-1	310	6,688	In-fill to 50 m	2005
217	332	9,612	In-fill to 50 m	2005
215-2	122	3218	In-fill to 50 m	2006
215-3	97	2782	In-fill to 50 m	2007
251	701	13,614	In-fill to 50 m	2005
217	507	14099	In-fill to 35m	2006
216	290	8736	In-fill to 75m	2007
Chichipate	4	121	Limestone	2006
Total	2,479	62,981

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10.5

Service Providers 2005

Generally, CGN’s exploration activities were carried out by CGN geological staff, under the direction of Colin McKenzie, P.Geo. Vice-President of Exploration, Skye.

From time to time, specialist survey and geological activities were performed by:

•

LAND INFO Worldwide Mapping, LLC.

•

T-Bear Contracting Ltd.

•

Golightly Geoscience Ltd.

•

J.K. Nieminen Consulting Inc.

•

Geografía y Forestería, S.A. (GYFSA)

•

Century Systems Technologies Inc.

Drilling was performed by the Guatemalan subsidiary of the Boart Longyear Group.

10.6

Interpretation of Exploration Information

10.6.1

Material Type Classification

Golightly Geoscience interpreted the analytical data from both the recent diamond core drilling by CGN and historic drilling by Exmibal, to provide a consistent set of lithological records across all deposit areas. In general, classification of the constituent materials, particularly identifying the contacts, within the laterite profile from visual inspection of drill core and auger cuttings is relatively unreliable unless supplemented by an interpretation of the chemical analyses.

The constituent laterite materials are described in Table 10-2.

Table 10-2: Summary of Material Type Classification


Description		Comment
Limonitic laterite	Cover	Mixed limonite and exotic sediments most notably felsic ash
	Limonite	Mostly in situ
Saprolitic laterite	Transition	Mixture between upper saprolite & limonite. High Co and MnO
	Saprolite	Usually in situ derived from serpentinized harzburgite or dunite
Bedrock	Peridotite	Usually serpentinized harzburgite or dunite in some cases brecciated
	Gabbro	Low Fe material with Ni lower than peridotite

10.6.1.1 Laterite Classification

The classification of weathered material is based on nickel and iron content as these are the only elements determined across the whole set of historic data. The classification of the feasibility study core samples is also largely based on nickel and iron and position in the profile with the exception of bedrock, which is based upon nickel and magnesium content.

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The chemical formulae used are provided in Table 10-3, Table 10-4 and Table 10-5.

Table 10-3: Chemical Classification of Laterite Materials


Major Unit	Layer	Formula
Limonitic		(Ni-0.1) 0.0005 × (Fe₂O₃-10)²
	Cover	Ni < 5.42 - 0.08 Fe₂O₃
	Limonite	Ni 5.42 - 0.08 Fe₂O₃
Saprolite		(Ni-0.1) > 0.0005 × (Fe₂O₃-10)²
	Transition	Ni -2.4 + 0.14 Fe₂O₃
	Saprolite	Ni > -2.4 + 0.14 Fe₂O₃
Bedrock (Exmibal samples)		(Ni < 0.5) and Fe₂O₃ < 14
Bedrock (CGN core samples)		MgO > 32% and Ni < 0.4%

With additional results from post-feasibility study drilling the classification formulae have been modified for use in future resource estimates. These formulae determine how different units within the laterite profile are classified.

Table 10-4: Chemical Classification of Laterite Materials- Exmibal Samples


Major Unit	Layer	Formula
Limonitic		(Ni-0.1) £ 0.0005 × (Fe₂O₃-10)²
	Cover	Ni < 5.42 - 0.08 Fe₂O₃
	Limonite	Ni ³ 5.42 - 0.08 Fe₂O₃
Saprolitic		(Ni-0.1) > 0.0005 × (Fe₂O₃-10)²
	Transition	Ni £ -2.4 + 0.14 Fe₂O₃
	Saprolite	Ni > -2.4 + 0.14 Fe₂O₃
Bedrock		(Ni < 0.5) and Fe₂O₃ < 14

Table 10-5: Chemical Classification of Laterite Materials- CGN Samples


Major Unit	Minor Unit	Formula
Saprolitic:		(-0.14 + Ni) > 0.00052*(Fe₂O₃ - 10)²
	Saprolite	Ni > (-2.4 + 0.2*Fe)
	Siliceous Saprolite	((Al₂O₃ ÷ Fe₂O₃)0.245) and Quartz > 10
	Transition	Ni £ (-2.4+0.2*Fe)
	Siliceous Transition	(Al₂O₃ < 0.245*Fe₂O₃) and Quartz > 10
	Bedrock	((Ni < .4) and SiO2 < (40 ÷ 35)* MgO
	Chromitic	(Cr₂O₃+0.31) > 1.5 + 0.0452* Fe₂O₃
	Aluminous	Al₂O₃ > 0.3* Fe₂O₃
	Gabbroic	(TiO₂ ÷ Fe₂O₃) > 0.045* (Al₂O₃ ÷ Fe₂O₃)
	Fresh Gabbro	CaO > 1
Limonitic:		(-0.14+Ni) £ 0.00052 * (Fe₂O₃-10)²
	Limonite	(Ni ³ 5.416666667+ -0.119149747*Fe
	Cover:	(Ni < 5.416666667+ -0.119149747*Fe
	Gabbroic Cover	(TiO2 ÷ Fe₂O₃) > 0.045 * (Al₂O₃ ÷ Fe₂O₃)
	Anomalous Cover:	MgO > 2
	Calcic Cover:	CaO > 1
	Sand Cover:	Cr₂O₃ > 0.01720121613 × Fe₂O₃^1.3075

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Quartz is calculated as the amount of SiO2 in excess of the empirical requirement for pure serpentine-limonite assemblage in saprolite. Such an MgO – SiO2 profile is a power law relationship.

Quartz = (SiO₂-A×MgOⁿ)×(SiO₂-100) ÷ ((SiO₂-100)-MgO × n × A × MgO^(n-1))

Where A = 2.044417877 and n = 0.8348996748

10.6.1.2

Bedrock Classification

Multi-constituent assays of the CGN core drilling were used by Golightly Geoscience to interpret and classify bedrock and proto-lithologies.

Fe₂O₃, Cr₂O₃, Al₂O₃ and TiO₂ are sufficiently insoluble and immobile that the mutual ratios of these constituents can be used to estimate the proportions of olivine, pyroxenes, chromite and plagioclase in the protolith. Definite population breaks in the chemical data were interpreted which support the classification criteria in Table 10-6.

Table 10-6: Chemical Classification of Bedrock Materials


Bedrock		Discriminant Criteria
Gabbro		(Al₂O₃ > 0.3Fe₂O₃) and (Cr₂O₃ < 0.047Fe₂O₃)
Anomalous		(Al₂O₃ > 0.3Fe₂O₃) and (Cr₂O₃ > 0.047Fe₂O₃)
Clinopyroxenite		(CaO > 2)
Harzburgite 1		(Al₂O₃ < 0.3Fe₂O₃) and Al₂O₃ 0.18Fe₂O₃
Harzburgite 2		(Al₂O₃< 0.18Fe₂O₃) and Al₂O₃ 0.12Fe₂O₃
Dunitic Harzburgite		(Al₂O₃ < 0.12Fe₂O₃) and Al₂O₃ 0.03Fe₂O₃
Dunite		Al₂O₃ < 0.03*Fe₂O₃

10.6.2

Interpreted Layers for Geological Model

10.6.2.1

Prediction of Bedrock Position from Iron Chemistry

Twin core holes on average penetrated nearly 10 m deeper than the Exmibal auger holes and in general penetrated bedrock. Golightly Geoscience found that the variation in iron assays in the overlying saprolite relative to bedrock was very regular. It was thus possible to predict the relative depth to bedrock by deriving a mathematical function for the decrease of Fe₂O₃ grades in the saprolite. This function was then applied as a bedrock depth predictor in those boreholes that failed to terminate in bedrock.

The predicted depth interval to bedrock was entered into the borehole records as a special field, to enable integration into the geological interpretation. Additional volumes of saprolite that resulted from this interpretation were regarded as ‘Inferred’.

10.6.2.2

Generalization Classification of Layers

The geological sequence of layers in the lateritic profile is from top to bottom: limonite, transition, and saprolite. Above the limonite is a layer of cover material and below the saprolite is bedrock.

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However while the sequence is generally consistent there are significant exceptions where, for example, limonite and transition material in joints and fractures penetrate deep into the saprolite or where blocks of saprolite are completely surrounded by saprolite or transition material. The cover usually lies on undulating contact with the underlying limonite, but locally the two appear to be inter-layered. In some areas, particularly 217-1 and 215-1 limonite, transition and saprolite have either slid or have been eroded and re-deposited on top of cover.

For the purposes of resource modeling it was necessary to simplify the classification of the laterite layers in a non-repeating sequence of 1 cover, 2 limonite, 3 transition, 4 saprolite and 5 bedrock. Samples were taken in 4 m composites. Where there was repetition of layers the sample was iteratively composited with 3 to 5 adjacent samples. If >10% of the material in any hole was still incorrectly classified with 5m compositing it was classified manually, with particular attention to mis-classification near layer boundaries.

Snowden geologists examined the finally classified layers for lateral consistency in 3D.

11.

Drilling

11.1

Introduction

Information in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included for completeness.

11.2

Exmibal Drilling and Pitting Prior to 1981

The database identifies eight different drill types, of which seven were augers. These included hand operated and power equipment. There were three types of pits, with channel sampling of the walls in the XTPC series and bulk sampling of the excavated material in the others.

The bulk of the data was obtained with a tractor mounted B-40 auger drill rig. A limited amount of core drilling was done with a McKinney rotary rig. This was not diamond core drilling. Penetration however was limited but was thought by Exmibal to be of higher quality recovery than the augers. McKinney drilling was largely confined to a 100 metre grid pattern in La Gloria areas 212, 216 and 217, to Manto 4 area 251 and Montúfar Cristina area 221.

All boreholes and pits were vertical. The pit openings generally were at least one metre square.

The amount of historic data relevant to the present resource estimates totals 3,530 holes and 58,058 assay samples.

It is the case that grid locations may be sampled by more than one borehole or pitting campaign.

11.2.1

Procedures

Most boreholes drilled by Exmibal were power auger holes. In these cases generally, because the sampling tool is open to the hole, there is a tendency for drill cuttings to drop from the screws back into the hole giving only a partial recovery of the sample and contamination of the next sample. Augers are also expected to be ineffective for cutting and sampling hard corestones in the saprolite zone.

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Sampling practice in laterite deposits with low degrees of partial serpentinization and relatively hard corestones shows that all non-core drilling techniques tend to selectively recover middle to fine size fractions of the broken material, typically one to three inches in grain size diameter. Therefore it is generally expected that hard material such as corestones will be under-sampled and grades of recovered material should be biased high in iron in that zone. This issue was addressed by Exmibal in the exploration stage by a special pitting program in which samples from pits were compared to co-located borehole samples and an iron correction factor established.

Generally all historic samples but the bottom sample and some research pit samples represent 1 metre. The pits generally were 1 m by 1.5 m in area and up to approximately 40 metres in depth. All pits were subsequently filled after sampling but settlement and compaction over the subsequent 30-45 years results in some pits still being effectively open to a few metres in depth.

Table 11-1: Historic Drilling Methods and Extent


					Hole Numbering
Area	Holes	Length	Samples	Hole Type	Prefix
212	1,465	25,694	25,069	Track mounted Auger (e.g. B-40)	ABFF, ABTF, ABRF
212	15	129	134	Hand Auger	AHDS
212	50	879	891	McKinney (Core )	RMKW
212	302	4,094	4,189	Pit (1 x 1.5m)	XTPC, XTPR
213	247	2,103	2,184	Track mounted Auger (e.g. B-40)	ABFF, ABTF
213	2	27	28	Pit (1 x 1.5m)	XTPC
215	104	2,357	2,360	Track mounted Auger (e.g. B-40)	ABFF, ABTF
215	29	361	371	Hand Auger	AHDS, AHWF
215	22	275	279	Portable Winkie Auger	AWKF
215	44	679	687	Pit (1 x 1.5m)	XTPC,
216	2	23	24	Track mounted Auger (e.g. B-40)	ABTF
216	4	42	43	Hand Auger	AHDS
216	19	303	309	McKinney (Core )	RMKW
216	38	481	495	Pit (1 x 1.5m)	XTPC
217	660	11,828	11,848	Track mounted Auger (e.g. B-40)	ABFF, ABTF, ABRF
217	20	235	235	Hand Auger	AHDS, AHTF
217	4	42	42	Portable Winkie Auger	AWKF
217	115	2,586	2,595	McKinney (Core )	RMKW
217	135	2,142	2,181	Pit (1 x 1.5m)	XTPC, XTPR
251	187	3,285	3,285	Track mounted Auger (e.g. B-40)	ABRF
251	20	217	226	Hand Auger	AHDS
251	46	572	583	Pit (1 x 1.5m)	XTPC, XTPE
Total	3,530	58,352	58,058

11.3

CGN Drilling from 2005

In April, 2005 CGN began a two part drilling program (twin hole drilling and in-fill drilling) on the Fenix Project property. The aims were to validate selected existing resource data as well as to conduct in-fill drilling to ensure there is a saprolite resource sufficient to support a feasibility study.

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Figure 11-1: Location of CGN Feasibility Study Drilling

LOGO

Figure 11-2: Location of Recent and Current CGN Drilling

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11.3.1

Twin Hole Drilling

The twin hole drilling program was designed to twin a statistically significant number of historic drill holes in the best known deposits, namely, the deposits in areas 212, 213, 217-1, 217-2, 217-4 and 217-5.

The twin holes were targeted within one metre of the original borehole in order to allow an unbiased geological representation of the deposit, the re-confirmation of the extensive historic resource database and improve knowledge of the multi-element chemistry of the deposits. A total of 116 holes for an aggregate of 4,111 metres were drilled.

The results of this part of the program confirmed the validity of the historic database and provided the basis for integrating the new and historic datasets, and are discussed further in Section 14 – Data Verification.

11.3.2

In-Fill Drilling

The in-fill drilling program was designed to detail more widely drilled deposits by carrying out drilling on 50 metre centres. The drilling took place on the 217-1, 217-2, 215 and 251 deposit areas. With the exception of the 251 deposit area, all drilling took place within five kilometres of the process plant at elevations ranging from 150 to 750 metres above sea level. The 251 deposit area underlies an upland plateau at 500 to 600 metres above sea level and is 20 kilometres west of the process plant.

The program was completed on 31 January, 2006.

Since then CGN has diamond drilled in four areas shown in Figure 11-2. In the north part of area 217 the existing 50 metre grid was filled in a centred pattern to 35 m in order to upgrade the classification of the resource. Currently, to identify new mineral resources, historic drilling on a lower terrace and an upper ridge in 215 is being filled into a 50 m grid. Historic drilling on area 216 is being filled in to a 75 grid. Drilling statistics are listed in Table 10-1. The mineral resources identified by this drilling will be reported at a later date

11.3.3

Procedures

In support of the drilling activities truck access roads, drill roads and drill sites were prepared by Guatemalan construction contractors. Drill sites were surveyed for final locations and elevations by a certified Guatemalan surveyor using radar theodolite (total station) relative to a number of benchmarks.

Drill layout, numbering, sequencing, collar preservation and other procedures were devised at the start of the drill program. These were progressively refined to suit local operating conditions. Golightly Geoscience and Snowden observed drilling operations and core retrieval in the field and found that the procedures conformed to industry practice. These included:

•

Identification of location in the field.

•

Photographing drill site.

•

Supervising clearing.

•

Verification of hole location on cleared site.

•

Checking drill rig set-up including vertical drill rod alignment.

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•

Monitoring drill advance, core collection, and recoveries.

•

Verification of core handling, depth markers, and core diameters.

•

Confirmation of hole bottoms in 2-3 m of bedrock.

•

Photographing completed drill site.

•

Checking concrete marker.

•

Second photographing of cleaned up site, if required.

Recovered lengths and core diameters were measured at the drill site and placed in covered plastic core boxes to prevent excessive drying.

11.4

Extent of Drilling

A full listing of all boreholes and pits used in the 2006 mineral resource estimates is provided in Appendix D of Snowden 2006.

11.5

Results

The results of all drilling and pitting are summarized in Appendix E of Snowden 2006.

Mineralized intercepts of limonite, transition and saprolite are reported at nickel grade cut-offs of 1.25 % and 1.50 % over a minimum vertical thickness of 2 metres, and including 2 metres of internal sub-grade material. The true thickness of the mineralization is approximately equivalent to the vertical thickness; however nickel laterite mineralization thicknesses are typically variable throughout a deposit. Individual thick intercepts may be unrepresentative due to the presence of localized, enriched weathered structures of unknown orientation. Similarly, wide-space drilling cannot predict short-scale irregularities in the weathered profile such as bedrock pinnacles and troughs.

12.

Sampling Method and Approach

12.1

Introduction

Information in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included for completeness.

12.2

Sampling Methods Prior to 1981

The majority of samples were obtained by track mounted auger rigs, with selected sites re-sampled by pits.

12.2.1

Relevant Details

Details of location, number, type, nature and spacing or density of samples collected, and the size of the area covered, are provided in Appendix D and F of Snowden 2006.

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12.2.2

Sample Risk Factors

The reliability of the historic auger samples was investigated by Exmibal’s special pitting program and subsequently re-assessed by CGN during its diamond core twin drilling program in 2005. A description of the assessment is provided below.

12.2.2.1

Mechanical Bias of Auger Assays

In general, it is expected that auger drills tend to reject or not cut hard corestones or are deflected into soft matrix rich zones. This should give a sample biased towards the composition of the local matrix, i.e. generally iron rich and either high or low in nickel depending on position in the profile.

In order to establish confidence in the historical data and to further quantify any biases, 116 sample sites in areas of historic 25 m grid auger drilling patterns were “twinned” by a new CGN diamond borehole located about 1 to 2 m away from the original, well within the geostatistical range of the data.

To demonstrate the overall effect of the twin holes on grade and thickness, simple graded zones were calculated for the new and old holes using 1.0, 1.5 and 2.0% Ni cut-offs using all data bracketed by the shallowest and deepest intersections above cut-off. Nickel, iron grades and thickness of the graded zones are summarized in Figure 12-1 to Figure 12-3.

At a nickel cut-off grade of 1.5% the twin hole nickel grade is lower by 0.15%; the graded zones are 2 metres thicker and the twin iron grade expressed as Fe₂O₃ is 3.43% lower, seemingly confirming the general magnitude of the mechanical bias of auger drilling but also due in part to the deeper penetration of the diamond boreholes. 49% of the old holes chosen had bottom nickel grades greater than 1.5%. However only 43% showed a gain in thickness in the twin. This is probably due to the unbiased sampling of low iron, low nickel corestones at the bottom of the historic zone by the twin diamond borehole. All twin holes penetrated beyond the final depth of the corresponding old hole.

A high nickel bias in the historic holes as a function of iron is a maximum at 30% Fe₂O₃ and decreases upwards and downwards from there. As seen in Figure 12-5 a high bias in the lower part of the saprolite is reversed in the upper part and in transition zone material.

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Figure 12-1: Comparison of Nickel Grades Between Co-located Boreholes

LOGO

Figure 12-2: Comparison of Iron Grades Between Co-located Boreholes

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Figure 12-3: Comparison of Mineralized Thicknesses between Co-located Boreholes

12.2.2.2

Correction for Sampling Bias

The bias observed in twinned CGN and Exmibal holes was investigated further by comparing nickel and iron assays from the in-fill drilling program in other areas, and throughout the laterite profile (Table 12-1).

Table 12-1: Comparison of Historic Exmibal and CGN Diamond Drilling


	Distance Relative to Top of Saprolite		CGN Fe₂O₃	CGN Ni	Exmibal Fe₂O₃	Exmibal Ni	D Fe₂O₃	pNi
Twin	-30	-20	63.70	0.82	29.74	0.21	-33.96	-0.61
Twin	-20	-10	65.44	1.06	61.44	0.92	-4.00	-0.14
Twin	-10	0	66.45	1.30	64.12	1.25	-2.33	-0.04
Twin	0	10	17.42	2.18	20.57	2.21	3.15	0.03
Twin	10	20	10.48	0.95	14.98	1.61	4.50	0.65
Twin	20	30	8.38	0.28
Twin	30	40	8.22	0.24
215-1	-20	-10	61.30	1.40
215-1	-10	0	55.78	1.49	51.84	1.50	-3.94	0.02
215-1	0	10	17.81	1.90	21.07	1.83	3.26	-0.08
215-1	10	20			15.24	1.71
251	-20	-10	65.31	1.39	64.51	1.39	-0.80	0.00
251	-10	0	59.34	1.53	57.40	1.53	-1.94	0.00
251	0	10	17.28	1.99	22.04	1.97	4.76	-0.02
251	10	20	14.91	1.86	21.30	1.85	6.39	-0.01
217-1&-2	-20	-10	65.04	1.36	62.81	1.38	-2.23	0.03
217-1&-2	-10	0	57.34	1.49	55.83	1.52	-1.50	0.03
217-1&-2	0	10	18.85	1.90	21.77	1.94	2.92	0.05
217-1&-2	10	20	18.90	1.69	22.18	1.69	3.28	0.01

Note: Twin holes used all available Ni data. A cut-off of 1.25% Ni in a four sample running average was used for in-fill borehole data.

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Comparisons of nickel and iron grades in equivalent parts of the laterite profile are illustrated in Figure 12-4 to Figure 12-6. Average grades from the twin holes and corresponding historic holes are plotted by depth above and below the limonite boundary in Figure 12-5 and Figure 12-6.

LOGO

Figure 12-4: Comparison of Historic Auger and Twin Core Assays

Note: The large symbols are the twin and old hole Ni:Fe₂O₃ medians plotted in Figure 12-5 and Figure 12-6. The dots represent about 51,000 historic borehole assays and more than 10,000 core assays from CGN drilling available in mid-December 2005.

The variation in nickel was found to be localized and inconsistent and no corrections were applied to the historic nickel assays (Figure 12-5).

LOGO

Figure 12-5: Nickel Grade Profiles from Twinned Exmibal and CGN Holes

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The bias in iron was found to be significant and consistent as is illustrated in Figure 12-6, thus supporting a decision to adjust some of the historic iron assays.

LOGO

Figure 12-6: Iron Grade Profiles from Twinned Exmibal and CGN Holes

The iron bias in the transition and saprolite was corrected by Golightly Geoscience by applying the following formulae to the Exmibal auger assays:

•

Corrected Fe₂O₃ in the saprolite = Maximum[(Fe₂O ₃ -6),(8.54 + 0.25Fe₂O₃ + 0.39 Ni)]

•

Corrected Fe₂O₃ in the transition = Fe₂O₃ – 6

In many of the old borehole cases, the iron grades in the limonite were lower than the corresponding twin hole. In some cases where the limonite profile exceeded 15 metres, cover was encountered in the core holes, thus contributing to the inconsistent pattern. Exmibal’s iron assays from limonite were therefore left uncorrected.

12.3

Sampling Methods from 2005

All sampling in 2005-2006 for the resource estimates was conducted by diamond core drilling.

12.3.1

Relevant Details

Details of location, number, type, nature and spacing or density of samples collected, and the size of the area covered, are provided in Appendix D and F of Snowden 2006.

12.3.2

Sample Risk Factors

Diamond core drilling of laterite deposits is considered to provide representative samples provided steps are taken to minimize core loss through the use of drilling mud and triple-tube recovery systems.

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It is a well-known feature of laterite deposits that rapid changes in material hardness may have an impact on sample recovery. Nickel may also occur in smectite clays that are susceptible to loss by washing during the coring action.

CGN monitored the core recoveries and instructed the drilling contractor to re-drill sites where recoveries were unacceptable. The required core recovery was 85%.

Core drilling is considered to be better than auger drilling in penetrating the full laterite profile and substantiating the position of the bedrock, as augers may terminate prematurely due to hard corestones. The decision to terminate the core hole is based on a visual assessment of the degree of weathering in the bedrock; however, this can be subjective. There is a risk that core holes may be terminated prematurely due to mis-identification of the bedrock and thus the full extent of the laterite profile may remain untested.

12.3.3

Density Determinations

Density measurements on moist core were derived by the water displacement method and subsequently calculated on a dry basis after contained moisture was measured.

At the drill site, three to five representative sections of core with good integrity were wrapped in plastic, preparatory for wet bulk density measurements at the logging station. There, the core was described, classified, and 0.3 to 1.3 metre lengths defined for sampling. Sample lengths and diameters were recorded, and bagged, tagged and weighed. The volume of the plastic wrapped core was then measured by the water displacement method and bagged. Subsequently moisture contents were calculated by BSI Inspectorate in Guatemala City.

The density results are discussed further in Section 14.

13.

Sample Preparation, Analyses, and Security

13.1

Introduction

Information in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included for completeness.

13.2

Sample Preparation prior to Dispatch of Samples

13.2.1

Exmibal Sampling prior to 1981

There is no detailed written statement of the routine sampling method for the Exmibal programs. During these programs, standard Inco practice would have been to reduce routine borehole and pit samples to about 1 kilogram. This was achieved by coning and quartering in the field or by riffle splitting after drying and crushing in the sample preparation laboratory. An exception was the special pitting program to determine bulk density, moisture and size distribution of the laterite. Entire 0.25 or 0.5 metre intervals from the pits were sampled. In this case, all material was sized on 6 inch and 2 inch screens, weighed and 5 to 10 kg portions assayed.

All the assay and location information in the exploration and development stages described above is compiled in the historic database used in this study.

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13.2.2

CGN Sampling from 2005

The drill core from the drilling program was logged, photographed and marked for sampling by CGN’s geologists in a logging station at the plant site. The sample lengths were nominally one metre.

All cores in the twin hole program were cut in half either by knife or diamond saw depending on sample hardness. One of the halves was bagged, tagged and shipped to the sample preparation laboratory. The remaining half was returned to the core box for storage and future reference. Sample rejects were returned to the Fenix Project property after 15 days for long term storage. Sample pulps are presently stored at the sample preparation laboratory.

In the in-fill drilling program one in ten holes was split and one-half of the cores saved. The other cores were sampled in their entirety. One sample in 20 or once in each borehole a core sample was split and a sample of half core was weighed, bagged and tagged in an identical fashion as a designated duplicate.

13.3

Statement Regarding Sample Preparation

All aspects of sample preparation prior to dispatch of samples were conducted by employees of CGN, the partially owned subsidiary of the issuer, Skye.

13.4

Dispatch of Samples, Sample Preparation, Assaying and Analytical Procedures

13.4.1

Exmibal Programs prior to 1981

According to Harju (1979) until about 1961, the samples were prepared in Guatemala and pulps sent to Canada for assaying. The first site laboratory was in Quirigua from late 1961 to 1964. When exploration entered the development stage in 1967, a laboratory was built at Las Dantas near El Estor, which operated until 1976, at which date the plant was built and a site laboratory installed.

Assaying at the Quirigua laboratory during what Inco regarded as the exploration phase was by cyanide titration. The titration method effectively measured the sum of nickel and cobalt grades rather than the nickel grade alone. Assaying at Las Dantas during the development stage, from about 1967 was by atomic absorption spectrometry (AAS).

Harju (1979) states that, in the field lab, the samples were crushed to pass an 80 mesh per inch screen and assayed. This is coarser than the nominal 150 mesh used today but will normally give adequate nickel, cobalt and iron extraction from nickel laterites which are naturally mostly very fine grained and/or porous. The grain-size of limonite, for example, is typically <200 mesh.

13.4.2

Calculation of Nickel and Cobalt from Cyanidation Ni+Co assays

A correction, developed by Golightly while employed by Inco, based on the AAS analyses of samples from the development drilling program in the El Estor region, was eventually applied by Inco to estimate nickel and cobalt individually from the Ni + Co and iron for the exploration phase data.

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Based on a regression analysis iron, cobalt and nickel XRF data of over 2,000 development stage assay samples, Golightly (1977) recommended the following equation as a method for calculating cobalt values from the cyanide titration Ni + Co assays.

< Co > = (0.00123 + 0.00083 * Cyanide Ni) – 0.0069

Here Cyanide Nickel = Ni + 1.25*Co (not Ni + Co)

A regression analysis conducted in 2005 of the data shows that, a slightly different equation,

< Co > = (0.001171 + 0.000805* Cyanide Ni) – 0.008354

better describes the actual implementation of the above correction.

13.4.3

Calculation of MgO and SiO₂ from Historical Data

MgO and SiO₂ are important constraints on pyrometallurgical processing and the SiO₂ / MgO target ratio impacts mine planning in the current feasibility study. Although these oxides were routinely monitored during mining, MgO and SiO₂ are generally absent from the historical drilling data. To calculate them for use in the current resource estimates it was necessary to develop a regression of MgO or SiO₂ on Fe₂O₃ and Ni based on the CGN core drilling data.

Many regression algorithms were tested and those finally selected proved robust and adaptable to different areas. Empirically based parameters varied from area to area. It was noted in the recent CGN drilling as well as from Exmibal ‘ore-zone’ composites that Areas 213, 212, 217 and 251 have distinctly different SiO₂-MgO-Fe₂O₃ relationships and that 213 and 217 are somewhat more siliceous. Area 251 is more MgO-rich than Area 212.

Apart from local quartz-rich structures in 212, 217 & 251, there was no consistent difference in the SiO₂-Fe₂ O₃ relationship for the four areas. There was however a substantial difference in the MgO-Fe₂O₃ curves for all areas. Each of these areas has a different mix of harzburgite and dunite protoliths which may explain the differences.

In both cases, MgO and SiO₂ for bedrock, saprolite and transition material were calculated with a different equation than limonite and cover. The latter uses a simple linear regression but the former have curved non-linear relationships that were not easy to model robustly (Table 13-1 and Table 13-2).

Table 13-1: Formulae to Calculate SiO₂ from Ni and Fe₂O₃

Equations for SiO₂

Saprolitic: SiO₂ = a₀ + a₁Fe₂O₃ + a₂ Fe₂O₃² + a₃ Fe₂O₃³ + a₄ ln(Ni/Fe₂O₃)

Limonite & Cover: = a₀ + a₁ Fe₂O₃

Regression coefficients for SiO₂


	a₀	a₁	a₂	a₃	a₄
Saprolitic	39.835373	0.281861	-0.019326	0.000115	0.175097
Limonite	31.251559	-0.400762

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Table 13-2: Formulae to Calculate MgO from Ni and Fe₂O₃

Equations for MgO

Saprolitic: MgO = c₀ + c₁(41.5-A(Fe₂O₃-7) ^B)

Limonite & Cover: = c₀ + c₁ Fe₂O₃

c₀, c₁ , A and B in the power law relation all differ from area to area. Otherwise all the

coefficients are equal.

Regression coefficients for MgO


Area	c₀	c₁	A	B
Saprolitic
212	-2.935258507	1.057613373	4.786217247	0.503403027
213	-1.788831961	1.022270142	5.274969597	0.498327413
217	0.087303889	0.968686251	4.285394364	0.563139727
251	-0.928109626	1.024271819	3.741105064	0.584585525
215	0.087303889	0.968686251	4.278275801	0.548272155
216	-53.9890767	2.254866248	1.86531075	0.58610776
Limonite	0.1147496859	0.012293

13.4.4

CGN Programs from 2005

Samples were shipped from the Fenix Project site to BSI Inspectorate in Guatemala City for sample preparation. Samples were dried at 105º C for 12 hours, crushed to 90% passing 10 mesh, riffle split and pulverized to 95% passing 150 mesh. The density samples were treated identically with their wet and dry weights recorded separately but then the samples were recombined into the assay interval to which they belonged.

All were then crushed to 80 % passing -10 mesh, riffle split and a 250 gram sample pulverized to -150 mesh.

The prepared samples (150g to 200g of pulverized laterite) were air freighted to the Lakefield, Ontario facilities of SGS for assay by lithium borate fusion and x-ray fluorescence. SGS has ISO/IEC 17025 accreditation for its mineral analytical services.

Eleven major element oxides, loss-on-ignition (LOI) plus nickel and cobalt were analyzed. Detection limits are 0.01% for nickel and 0.01% for cobalt. The assayed constituents are Ni, Co, MgO, SiO₂, Al₂O₃, Cr₂O₃, Fe₂O₃, Na₂O, K ₂O, TiO₂, MnO, P₂O₅, V₂O₅ and LOI.

13.5

Quality Control Measures

13.5.1

Exmibal Programs Prior to 1981

According to Harju 1979, a routine quality control (QC) program using other Inco laboratories in Canada and some external laboratories was practiced. The details of the historic QC are no longer readily available. Original assay sheets have not been consistently preserved, so it is difficult to compare with present day practice. However, the concept was similar and Golightly Geoscience is confident, based on experience with Inco practice in the 1970’s that QC was thoroughly and diligently carried out.

13.5.2

CGN Programs from 2005

Field duplicates (1:20) and standards or blanks (1:20) are inserted at the Fenix Project site to monitor laboratory procedures and assay quality. All batches of analysis are subjected to statistical tests to ensure they meet established QC criteria. SGS also has its own QC procedures which include insertion of standards and duplicates.

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ALS-Chemex in Brisbane Australia was selected following a round robin assaying procedure to provide additional analytical checks on the SGS analyses. ALS-Chemex used a fusion ICP-AES (ME-ICP93) method for the same suite of elements measured at SGS Lakefield. Analyzed pulps were selected as 60 g replicates for re-assay at a frequency of 1:50.

13.5.2.1

Standards

Four standards, two saprolites and two limonites, were prepared from small bulk samples taken from the Area 212 mine originally for metallurgical test work in 2005. The first two were prepared at the beginning of CGN’s drill program and the second pair in December 2005. They were pulverized, homogenized and assayed by SGS. A dozen representative pulps of each were also assayed by three external labs 1) ALS-Chemex in Brisbane, Australia 2) Amdel in Adelaide, Australia and 3) BSI in Reno, Nevada USA.

The first pair of standards was characterized by XRF by SGS, Amdel and BSI and by fusion ICP AES by ALS-Chemex Brisbane. For the second pair of standards BSI used a combination of ICP and XRF.

In mid 2006, two additional standards, a limonite and a saprolite, were prepared and characterised by SGS using the same procedures and a dozen of each standard were also analysed using the same techniques as above by ALS-Chemex in Brisbane and Amdel in Adelaide. The two additional standards have been in use since the exhaustion of the previous two, in spring 2007.

13.5.2.2

Blanks

Limestone blanks were prepared from a stockpile at the limestone quarry near Chichipate and replenished on several occasions. There is no control on the composition of this blank beyond the >2000 assays of it to date. It clearly has below detection limit nickel and the very high calcium and Loss on Ignition (LOI) make it very easy to detect in cases of sample switching.

13.5.2.3

Monitoring of Assays

Assays were checked against standards by Golightly Geoscience using a program expressly designed for the purpose and re-checked afterwards when the assay files were imported into a DHLogger database. Golightly Geoscience’s program searched for and identified individual standards.

In general, a three standard deviation limit was used as a “failure” criterion. In cases where a failure was apparently part of a pattern of drift, batches were re-evaluated and usually re-assayed by SGS. Larger deviations unlikely to be due to analytical drift were usually found to be sample switching either at the logging station, the sample preparation facility, or the assay laboratory. Initially the average and standard deviations of the standard characterization analyses were used to screen new data. Eventually however, because in some cases this gave too narrow a window to accommodate normal inter-batch variability, the standard deviations used were revised to match those of the assayed standards.

Every 50^th assay pulp was re-analyzed by ALS Chemex, Brisbane using fusion ICP-AES.

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13.5.2.4

Results and Corrective Actions taken

Analytical problems detected were:

•

A period of high LOI in the first 3,000 assays which caused low biases notably in nickel and magnesia; according to external checks, this left a probable small overall low bias <0.05% Ni during that time period which was not completely rectified by routine re-assays.

•

A definite high alumina bias relative to the external laboratory checks for the last half of the in-fill program.

The dominant practical problem was incorrect standard identifier entered in the database, much of it relating to the change-over between the first set of standards and second set.

The majority of failures were caused by assay sample sequence mix-ups, usually a simple exchange between adjacent samples:

•

25 of 3,000 re-assays had major element changes > 5% in particular sample numbers indicating <25/3000 or 0.8% sequence mix-ups in the assay laboratory corrected in the final issue.

•

6 of 674 or 0.8% external check assays differed strongly in multiple elements indicating an assay sample sequence mix-up, confirmed in two cases by SGS.

•

9 of 1483 or 0.6% blank samples were recorded as assay samples.

•

27 of 1495 or 1.8% blank samples registered measurable 0.01-0.04% Ni.

Golightly Geoscience is confident that most significant errors have been found and corrected and that the data are of sufficient reliability on which to base resource estimates.

13.6

Security

13.6.1

Exmibal Programs prior to 1981

Golightly Geoscience is unaware of specific precautions taken against tampering with the samples or assays. The system of QC and the 15 year chain of verifications all the way from early exploration through to production suggest that security was not breached.

According to R. C. Osborne (pers. comm.) no electronic exploration database for Exmibal ever existed at site. Rather, it was compiled, administered and evaluations conducted out of the Inco corporate office in Toronto and later out of the Ontario Division in Copper Cliff. In both locations the access to the database was not available to anyone other than the computer administrators at Toronto or Copper Cliff. Additions, deletions and revisions were done in duplicate keypunch by computer operators following instructions from exploration staff. In the period 1995 to 2004, Inco Exploration & Technical Services and then Inco Technical Services Limited’s Laterite Group took over the management of the database. Since then only one person from exploration, Inco’s laterite specialist, had read and write access to the server directory on which the files were stored. Computer Services also had an administrator with all rights. The database was provided to Skye as part of a GIS project in Access & Arcview format on CDROM following revisions by PT Ingold Management of Jakarta under Inco’s laterite specialist’s supervision.

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Golightly Geoscience has complete confidence in Inco’s diligence in this regard and has no reason to doubt the security of the earlier procedures.

13.6.2

CGN Programs from 2005

The drilling data were assembled into a database using the DHLogger software program. The central repository is resident on the server in the exploration office at the plant site near El Estor. Assay data and QC were preliminarily assessed by Golightly Geoscience using proprietary software systems and then uploaded into the central database under supervision at the exploration office. The DHLogger system has restricted access and a detailed record of all data entry and changes for every sample in the system is available.

The historic Exmibal data is also resident in the DHLogger database and incorporates corrections to the verified historic data made by Golightly Geoscience and CGN.

13.7

Statement on the Adequacy of Sample Preparation, Security and Analytical Procedures

In the opinion of Golightly Geoscience the procedures undertaken by Exmibal and CGN in the exploration and evaluation of nickel laterite deposits were adequate.

There is no evidence to suggest that samples were subject to tampering.

Analytical procedures are consistent with industry practice.

14.

Data Verification

14.1

Introduction

Except for as set out in Section 14.9, information in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included for completeness.

14.2

General

This section provides a discussion of the quality control (QC) measures and verification procedures concerning the data used for the 2006 resource estimates (Table 14-1).

Table 14-1: Verification Stages


Period	Data	Primary Verification	Subsequent Verification
1960 - 1980	Exmibal sampling data	Inco	AMEC 2003 Golightly 2005-6 CGN 2005-6 Snowden 2006
2005 - 2006	CGN sampling data	Golightly	Snowden 2005-6

Golightly Geoscience verified the historic Exmibal sampling data by recompiling the source data and assessing the results of the twin drilling program undertaken by CGN.

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Data from CGN’s drilling program was checked by Golightly Geoscience and included monitoring of QC of the geological logging and analytical results.

Upon receipt of the compiled data provided by Golightly Geoscience, Snowden undertook further validation to ensure that the data was clean, correct and useful for the purposes of resource estimation.

The resource estimation for Area 212 included a new estimate for the mined volume based on historic Exmibal exploration data. Snowden reconciled the exploration model with production data to provide another check on the reliability of primary exploration data.

14.3

Historic Database

14.3.1

Background

Assay data were routinely digitized by key punch entry with verification in the late 1970’s and metal resource inventories (MRI’s), calculated in 1980 using the Inco Laterite Evaluation System, at that time a mainframe computer program first written in about 1968 at Inco’s J. Roy Gordon Research Laboratory (JRGRL) in Mississauga, Canada and revised, improved and adapted to different hardware over the following 30 years. However, the comprehensive database used in the present study was re-assembled in part from earlier digital compilations and from paper logs during the early 1990’s.

14.3.2

Verification by AMEC, 2003

Appendix G of Snowden 2006 is an extract from AMEC 2003 that discloses the data verification performed by AMEC.

14.3.3

Verification 2005-06, Golightly Geoscience

14.3.3.1 Initial Checking of Historic Location Information

The borehole logs existed in paper copy at site and on microfiche, but were only partially re-digitized during the early 1990’s. The re-digitization project was curtailed at some point in the 1990’s so that logged sample descriptions for only 6,827 of the 104,345 assay samples (6.5%) were transcribed into the database. For this study, this made it practically necessary to create a chemically based sample classification scheme described in Section 10. Similarly, start and finish date, ‘logged by’ and hole-ending comments also were only transcribed for about 400 of 7,069 holes (5.6%) although all archival sources were routinely accessed as the data were prepared for use in resource estimates.

The following observations were made of data outside of the areas in the current resource study:

•

23 holes mostly from Amate, area 260, have no collar coordinates. Similarly the small suite of holes in area 254, meant to be from the Chulac property, actually plotted at the west end of Amate. An error in coordinate conversion is believed to have occurred. Wide spaced reconnaissance collars in Mantos 5 -10 lack elevations. Terrain slopes calculated from the database collar coordinates showed a number of anomalous values suggestive of data entry errors.

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14.3.3.2

Detailed Location Checking

The 116 twin hole program attempted to place a new diamond borehole within 0.5 metres of the collar of an old hole. Several dozen borehole collars in deposits 212, 213 and 217 selected for re-drilling were precisely confirmed by re-surveying using surveying benchmarks in the mine site as reference. On the other hand, physical evidence of boreholes at the limonite/saprolite transition in the stripped areas within the mine workings of area 212, block 2 were rarely found; they are thought to have filled by collapse in sufficiently wet material in the saprolite and lower limonite zones.

In actual performance, an average distance of about 2 metres from the old survey co-ordinates was attained.

All new surveying was done by GYFSA using radar theodolite relative to a number of well documented mine benchmarks including at least one of which is a Guatemalan geodetic survey point.

A four-step procedure was used to verify historical collar locations in the areas selected for in-fill and twin diamond core drilling:

•

The differences between database collar elevations and the elevation of the nearest contour on one of three digitized contour maps were checked. The two maps digitized were custom photogrammetric maps of the La Gloria and Manto 4 (area 251) areas by Hunting Survey Corporation 1961, and a satellite image photogrammetric map produced by LAND INFO for the Fenix Project in spring 2006. Due to tree cover, collar surveys that agreed with contours to within 7 metres were accepted. Any differences greater than this were investigated.

•

Elevation or slope anomalies relative to neighbouring holes were investigated.

•

Inconsistencies in drill grid spacings were resolved.

•

Inconsistency of hole numbering systems and local grids were resolved.

The changes of borehole location had a significant impact on the interpreted extent of the mineralization in area 251. As a consequence, the area of the historic resource lying outside the 2005-2006, 50 m in-fill drilling program was adjusted and reclassified to Inferred in the current study.

14.4

Calculated Grade Data

14.4.1

Calculated MgO and SiO₂ Values

The calculated values were validated by comparison with actual MgO and SiO₂ grades from approximately 200 ore-zone composites (K-composites) that were prepared and analyzed by Exmibal first in 1970-1972 and reviewed in 1980. The K-composites were re-analyzed by Inco both at site in 1980 with concurrent checks at JRGRL and later by ALS-Chemex (Toronto) in 1993. Their contents were used to predict future ore composition in several feasibility studies. They were reanalyzed again by SGS for CGN using the same methods as the current drill program.

Calculated MgO and SiO₂ values were also validated by comparing average estimates of MgO and SiO₂ on deposit scale with measured averages in in-fill or twin drilling.

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Table 14-2 below gives a comparison of overall SiO₂ and MgO between diamond drilling data and calculated values based on the corrected Fe₂ O₃ contents of the related historic holes. There is good practical agreement and a small underestimate of the SiO₂/MgO ratio by the calculation in the historic holes. The overall ratio in saprolite is about 1.45, well within the range suitable for ferro-nickel smelting.

14.4.2

Calculated Ni and Co Values

The calculated nickel and cobalt values were derived by the method described in Section 13.4.2 for part of the historic database that relied upon titrated Ni+Co assays.

Verification of the historic nickel, cobalt and iron assays was undertaken by re-analysis of a limited number of assay pulps by CGN from about a dozen boreholes in the 217 and 251 deposits. The data showed acceptable agreement between calculated values and re-assayed nickel and cobalt. Transcription errors where cobalt falls in the range 0.5-0.01% were identified and corrected. Where available, the approximately 200 ‘ore-zone’ composites (K-composites) were also re-analyzed by SGS and found to be in excellent agreement with the value originally attributed to them by Exmibal.

Table 14-2: Large-scale Validation of SiO₂ and MgO Estimates


	Depth Interval Below (above) Top of Saprolite		MgO	SiO₂	MgO Calculated	SiO₂ Calculated	DMgO	DSiO₂
Twin Holes	-20	-10	0.87	4.17	0.87	6.63	0.00	2.46
Twin Holes	-10	0	1.09	5.08	0.90	5.56	-0.19	0.48
Twin Holes	0	10	25.74	37.95	26.07	39.37	0.33	1.42
Twin Holes	10	20	32.18	40.26	28.58	40.13	-3.61	-0.13
215-1	-10	0	4.93	13.31	8.25	18.65	3.32	5.35
215-1	0	10	26.44	37.94	26.26	38.39	-0.19	0.45
251	-20	-10	1.32	5.79	1.77	7.09	0.45	1.30
251	-10	0	4.19	10.31	5.66	14.04	1.47	3.73
251	0	10	27.03	37.73	26.81	37.92	-0.23	0.19
251	10	20	27.48	40.43	27.60	37.52	0.12	-2.91
217-1&2	-20	-10	1.15	5.09	2.47	7.76	1.33	2.67
217-1&-2	-10	0	4.04	12.53	5.92	15.75	1.88	3.22
217-1&-2	0	10	25.01	37.96	25.41	38.14	0.40	0.18
217-1&-2	10	20	24.12	39.60	25.47	36.97	1.35	-2.62

Twin Holes used all data. Other areas used a cut-off of 1.25% Ni in a 4 sample running average. The first percentile of individual sample Ni grades was 0.9-1%Ni. MgO & SiO₂ calculated for historic data using corrected Fe₂O₃.

14.5

Bulk Density

There are six sets of data bearing on the in situ bulk density and moisture content of the mineralization and these served to establish and verify the formulae used to calculate Dry Bulk Density (DBD) values for samples used in the 2006 resource estimates (Table 14-3).

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Table 14-3: Sources of Density Data


Program	Author	Date	# Data
Special pitting program 1971-2	F. Barnard	1971-2	160
Limonite pits in mine	R. C. Osborne	1980	17
Core drilling: Total recovery density core drilling	CGN	2005-6	12,748	¹
Water displacement volume density core drilling	CGN	2005-6	362
Bulk density pits	D. MacKenzie	2006	13
Mine reconciliation ore stockpile weightometer data / mined volume	Snowden	2006

Golightly Geoscience undertook the following studies:

•

A review of the data from the special pitting data and comparison with the factors used by Exmibal during actual mining. It was concluded that:

•

The DBD varies systematically as a function of iron, nickel and depth.

•

Exmibal had applied constant DBD’s for saprolite, 1.08 t/m³ , and limonite, 0.96 t/m³.

•

Exmibal’s constant factors underestimated the density of saprolite as suggested by Exmibal’s 1980 mine reconciliation and confirmed by Snowden’s re-analysis of the mining data.

•

Analysis of the variation in DBD with level in the laterite profile. It was found that dry bulk density decreases rapidly as a function of iron assay from bedrock to a minimum value in upper saprolite. It remains constant in the upper saprolite and transition but jumps upward in the limonite and again in the cover. There is a slight linear increase with iron content in the cover.

•

Analysis of the saprolite data. Comparison of information from several of the sources showed data for saprolite to be somewhat problematic. The core samples in the saprolite zone tend to be fractured and to have relatively poor integrity, underestimating density values compared to the pit samples. A small saprolite pitting program by CGN (MacKenzie 2006) confirmed the historic pit data.

•

Analysis of the relationship between bulk specific volume and moisture content for saprolite, transition and limonite.

•

Establishment of regression formulae based on pit data to calculate DBD values for each sample (Table 14-4).

Table 14-4: Regression Formulae to Calculate DBD for Drill Samples

Regression for Bulk Specific Volume (SV) m³/t

Specific Volume = a₀ + a₁ Fe₂O₃ + a₂ Fe₂O₃²

Note: DBD = 1/SV.

¹	At the time of writing, less than half of the complete weight data for >30,000 core samples were available in usable form in the database.

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14.6

Twin Drilling

CGN’s program of twin drilling at numerous sites and across several deposits verified the following aspects of the geological and sample data:

•

The general grade and thickness of the mineralization in the historic boreholes. The results of the overall verification are provided in Section 12.1.

•

Incidence of corestones. Corestone intervals were expected to be under-estimated in the historic drilling. Exmibal had recognized this and undertook testwork to identify the potential for screening of the saprolite feed to remove corestones. They found that there was no potential to upgrade the saprolite feed by screening (Barnard 1972a 1972b).

•

Extent of the saprolite profile. The full depth of the saprolite profile was not expected to be tested in all of the historic drilling. Sufficient evidence was obtained in the twin drilling to confirm that there were many examples of the historic drilling falling short of the bedrock position and a method was established to predict the ultimate depth (Section 10.6.2.1).

•

Unbiasedness of the grades in the historic drilling. Although instances of poor correlation were identified at several sites of twin drilling, there was sufficient evidence that the historic nickel data was globally unbiased and could be used in resource estimates. The iron data was found to be biased and correction factors were established and applied to the historic data.

14.7

Verification by Snowden

14.7.1

Data Flow Background

Data was provided to Snowden by Golightly Geoscience and CGN in several trenches during the 2005-6 drilling program, including batches of historic sample data.

Historic borehole data were provided as industry standard computer files containing the following information:

•

Record identifier.

•

‘Holeid, X, Y, Z, Ni, Co, Fe₂O₃_Data’ were from the historic Exmibal database without any changes other than accuracy checks against original documents to correct typographic errors.

•

Sample ‘from, to’ information.

•

‘Chem_Label’, which is a chemical classification of each 1 metre assay sample based directly on the raw data above.

•

‘Fe2O3_DEBIAS’, which is a correction for iron bias.

•

‘Ni4 & Fe4’, which are four metre down hole moving averages of Ni and Fe₂O₃_assays respectively.

•

‘Layer’, which is an integer from one to five indicating the material type classification derived from the chemistry. The classification is an interpretation of dominant materials using the technique as described in Section 10.6.2.2.

•

‘Layer_Change’ to monitor manual changes in layer interpretation.

•

Fields for calculated SiO₂, MgO and DBD (density) values.

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•

‘Distance to nearest sample’ field.

•

‘Priority’ field to indicate sampling method and program.

Data from the recent CGN core drilling programs were provided as industry standard computer files containing similar information as the historic files except that the only calculated field was ‘DBD’ (density), and the full suite of elements and oxide constituents were also provided.

14.7.2

Sample Data Validation

Snowden compiled all primary samples and assay data provided by Golightly Geoscience into a series of Gemcom mining software workspaces, one each for areas 212, 213, 215, 217, and 251.

The following process of validation was undertaken:

•

Initial searches for duplicate intervals, missing intervals, out-of-range values, and inter-table inconsistencies using the validation tools provided by Gemcom.

•

Visualization of sample data in 3D, section and plan-view to identify errors and inconsistencies in the data. Errors and questionable records were provided to Golightly Geoscience and CGN for checking and correction where appropriate.

•

Subsequently any corrected records were updated to Snowden’s primary Gemcom project database and the process of checking was repeated.

•

Plotting of borehole cross-sections and topography surfaces from Gemcom to compare directly with historic hand-drawn cross-sections. Snowden completed these checks for areas 212 and 213 and identified a number of historic ABTF borehole collars with incorrect elevations. Compiled borehole collar files for the other areas were provided to Golightly Geoscience and CGN geologists for final review and cross-checking with hand-drawn cross-sections and plans.

•

Snowden geologists verified the geological layer interpretations in 3D and in cross-section, taking the assay data into consideration. Inconsistencies were noted and discussed with Golightly Geoscience and CGN geologists, prior to modifying the interpretation.

14.7.3

Topographic Surfaces

14.7.3.1

Excavated Areas

Mining has occurred in Areas 212 and 213 subsequent to the historic drilling.

CGN provided Snowden with a digital terrain wireframe model (DTM) of the mine area at La Gloria (Area 212) that was compiled from mine survey plans. This was reviewed in cross-section and in 3-D and compared with the digital borehole information using Gemcom computer graphics. The pit DTM was also compared with the historic hand-drawn cross sections. Several inconsistencies were identified and most were rectified by CGN. Final checking revealed some errors remained on the southern margins of the DTM under the haul road. Any mineralization in this area remains an un-classified resource due to uncertainties in the topographic model.

Pre-mine surfaces at 213 and 213 were constructed as DTMs by simple triangulation of relevant surveyed borehole collars.

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14.7.3.2

Un-excavated Areas

No mining has occurred in Areas 215, 217 and 251.

LAND INFO recently provided a DTM for the entire Fenix Project property; however the vegetation canopy for most areas has resulted in an unrealistic set of topographic contours which were not therefore suitable for use in the resource estimates. An exception was Area 251 where most of the drilled area had been de-forested by agricultural activity. The LAND INFO contours and DTM were used to control the geological model and delimit the surficial extents of the laterite profile in this case.

Topography surfaces at 215 and 217 were constructed as DTMs by simple triangulation of all surveyed borehole collars. Thus the subtle landform features evident in the photogrammetric contours provided by Hunting Survey Corporation in 1961 were not integrated into the DTM. The surficial extents of the laterite profile were defined by a 10 metre boundary from the outermost pattern of boreholes.

14.7.4

Reconciliation with Mine Production Information

Snowden conducted a study to reconcile Snowden’s 2006 resource model (the exploration model) estimated from wide-spaced exploration data with the historical mine production in blocks 212-1 and 212-5. The within-pit portion of the resource model reconciled in terms of overburden volume and mineralized material grade and tonnes (contained nickel) to within 8% of the historical production data on a global basis. Bench-by-bench reconciliation, although poor, was as expected given the nature of the mining process and the wider spacing of the exploration data (Snowden 2005a).

Historical information at site includes tabulated bench reports of mined overburden and ore which was determined at a cut-off grade of 1.6% Ni. The Snowden resource model in the mined areas is compared with overburden and ore production in Table 14-5 and Table 14-6, respectively.

It should be noted that during the mining process, selection of overburden and ore was undertaken following grade control sampling at close spacings and detailed geological mapping of exposed benches and faces. Grade control included auger drilling of 1 metre deep boreholes at spacings of 5 to 10 metres, and samples were taken at 0.5 m intervals (MacKenzie 1980).

14.7.4.1

Overburden from Blocks 212-1 and 212-5

In Table 14-5 the overburden in the Snowden exploration model is shown as a combination of cover, limonite and transition blocks that are graded less than 1.6% Ni. On this basis the estimated overburden from exploration data was 1.48 million cubic metres compared with actual excavation of 1.41 million cubic metres.

It can be seen that the volumes of overburden material in the exploration block model generally do not reconcile well with the historical production data on a bench-by-bench basis. Such differences are expected given the nature of the mining process; for example, irregularities in the actual mined benches versus the regular nature of the block model, and more importantly, the difficulty in accurately predicting local estimates from exploration data alone.

The model and historical production data do, however, reconcile to within 5% on a global basis.

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Table 14-5: Overburden Volume from 2006 Pre-mine Resource Model Compared with Historical Records

from Blocks 212-1 and 212-5


Bench Below (m)	Within-Pit Volumes - Snowden Exploration Model				Historical Production
Bench Below (m)	Cover (m³)	Limonite (m³)	Subgrade Transition (m³)	Overburden Total (m³)	Overburden Total (m³)
465	—	—	—	—	—
460	780	390	—	1,170	—
455	1,950	14,840	1,170	17,970	—
450	390	20,310	3,130	23,830	—
445	4,300	22,270	3,900	30,470	44,800
440	14,450	21,090	8,590	44,140	32,900
435	42,580	19,530	4,690	66,800	18,000
430	71,090	71,880	3,910	146,880	122,300
425	44,920	139,060	12,500	196,480	96,500
420	42,190	176,950	17,580	236,720	287,000
415	39,060	153,520	44,140	236,720	253,300
410	35,550	108,980	37,110	181,640	182,410
405	15,630	96,090	21,880	133,590	193,040
400	3,520	79,690	19,530	102,730	63,960
395	3,520	21,480	10,550	35,550	71,320
390	2,730	7,420	1,560	11,720	20,600
385	1,560	2,340	1,560	5,470	16,570
380	—	1,560	3,130	4,690	—
375	—	—	780	780	—
370	—	—	—	—	—
365	—	—	—	—	—
360	—	—	—	—	2,900
Total	324,220	957,420	195,700	1,477,340	1,405,610

Note: Historical data sourced from Exmibal mine records (Production To Date, December 1980)

14.7.4.2

Mine Production from Blocks 212-1 and 212-5

In Table 14-6 the Snowden model is reported within the total pit volume excluding the overburden identified in Table 14-5. In this case the mined volume exceeds the modeled volume since additional ore was identified in the root zones of the saprolite profile where exploration drilling had not penetrated all areas.

The within-pit tonnes and nickel grade of the mineralized transition and saprolite material in the exploration block model do not reconcile well with the historical production data on a bench-by-bench basis. Potential inaccuracies in, and variations associated with, the calibration of the on-site weightometer and determination of moisture content may also have contributed to the discrepancies displayed on a bench-by-bench basis, notwithstanding that the exploration model is interpreted from data that is wider-spaced than the grade control sampling.

Globally, however, the resource model for the mineralized material reconciles favourably with the historical production, given the various assumptions used in the generation of the exploration model from wide spaced data:

•

The resource model indicates a total of 913,000 tonnes of mineralized transition and saprolite material (adding the void below model tonnes to the global tonnes in Table 14-6) at a grade of 1.97% Ni (the nickel grade of the global transition plus saprolite material was assumed for the void blocks).

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•

The historical production indicates a total of 927,000 tonnes of saprolite ore at an average grade of 2.13% Ni².

•

The exploration model and historical production data reconcile to better than 2% with regards to tonnes and 8% with regards to nickel grade on a global basis.

14.8

Statements Regarding Verification

From the checks made, Golightly Geoscience and Snowden believe that the data verification has been sufficient to permit the information to be used in resource estimates for a feasibility study. Despite the best endeavours of the geoscientists working on the project, errors may still remain in the historic database; however in the opinion of the qualified persons these are likely to be immaterial.

14.9

Reliance on Owner Supplied Information

14.9.1

Geology and Mineral Resource Estimates

Snowden, Hatch and Golightly Geoscience have not verified the commercial, environmental and legal aspects of Skye’s mineral tenure as disclosed in Skye’s public documents.

14.9.2

Metallurgical Testwork

Samples for hydrometallurgical testwork were taken by Skye and CGN. Details of the sampling program are summarized from D. Neudorf’s report (Skye 2005) and presented in Section 16.

Overall nickel and cobalt recovery values in the hydrometallurgical process were estimated by D. Neudorf and were applied by Snowden in the calculation of nickel equivalence for limonite resource estimates.

14.9.3

Other Information

Hatch has relied on certain data and information provided by Skye in the preparation of the capital cost estimate and operating cost estimate set out in Section 18.11 of this Report and the economic analysis set out in Section 18.12. In particular, Skye has supplied information regarding local salaries, labour costs and, through a contract with PACE Global Energy Services, the long term price for coal, petroleum coke and HFO (heavy fuel oil). Skye also provided the list of project opportunities presented in Section 18.14. Although Hatch has no reason to question the reasonableness of this information, Hatch has relied on this information without independent verification.

Note on production grade. The average production grade of 2.13% Ni was determined by the Exmibal mine and geology department. Ultimate reconciliation from the process plant indicated an under-call of 5% in recovered nickel that was attributed to unaccounted process losses eg “loss in fines”, during the metallurgical process. The under-call may, in fact, reflect a lower mined grade delivered to the plant; however there is no documentation to resolve this (C. McKenzie, pers.comm.)

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Table 14-6: Comparison of 2006 Resource Model with Historical Production from Blocks 212-1 and 212-5


	Snowden Resource Model – Within-Pit Mineralisation								Historical Production
Bench Below (m)	Global (Total)			1.6% Ni Cut-Off Grade			Void Below Model		Historical Production
Bench Below (m)	Volume (m³)	Tonnes	Ni (%)	Volume (m³)	Tonnes	Ni (%)	Volume (m³)	Tonnes*	Tonnes	Ni (%)
450	3,520	4,320	2.16	2,730	3,410	2.37	—	—	—	—
445	6,250	7,760	2.19	5,080	6,510	2.37	—	—	—	—
440	9,770	13,270	2.14	7,030	9,410	2.46	—	—	—	—
435	11,330	15,230	1.94	8,590	11,560	2.11	3,520	4,520	26,500	2.19
430	20,310	25,270	2.09	17,190	21,810	2.24	6,640	8,540	5,690	2.14
425	37,890	48,080	1.99	30,860	39,300	2.14	5,470	7,040	25,230	2.33
420	46,480	61,220	1.99	39,840	51,340	2.15	7,420	9,550	85,440	2.13
415	88,670	111,350	1.97	75,390	91,900	2.12	9,770	12,560	86,580	2.09
410	149,220	189,760	2.05	101,950	28,840	2.19	19,530	25,130	92,830	2.11
405	135,550	180,150	1.94	107,030	36,580	2.18	19,140	24,630	212,510	2.17
400	55,860	74,120	1.85	43,360	53,750	2.14	8,980	11,560	210,350	2.13
395	41,800	50,800	2.05	35,160	41,390	2.22	1,950	2,510	33,470	2.23
390	5,860	7,950	1.92	4,300	5,710	2.21	—	—	134,500	2.02
385	1,950	2,210	1.23	390	370	1.64	—	—	11,560	2.01
380	3,520	4,030	1.47	1,170	1,270	1.76	—	—	—	—
375	7,030	8,740	1.12	—	—	—	—	—	—	—
370	1,560	1,930	0.82	—	—	—	781	1,005	—	—
365	—	—	—	—	—	—	—	—	—	—
360	—	—	—	—	—	—	—	—	2,800	1.83
355	—	—	—	—	—	—	—	—	—	—
Totals	626,560	806,160	1.97	480,080	603,160	2.17	83,200	107,050	927,470	2.13

Note: Historical data sourced from Exmibal mine records (Production To Date, December 1980)

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15.

Adjacent Properties

There is no information from adjacent properties applicable to the Fenix Project for disclosure in this report.

16.

Mineral Processing and Metallurgical Testing

16.1

Introduction

Information in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included for completeness.

Skye has been investigating the application of both pyrometallurgical and hydrometallurgical processes to recover nickel from the saprolite and limonite fractions, respectively, of the Fenix mineral deposits. Summaries of the testwork results and processing are provided in Section 16.2 below.

Skye has developed a novel atmospheric sulphuric acid leaching process for the recovery of nickel and cobalt from limonite and saprolite. This process was studied as the basis of the hydrometallurgical options in Hatch 2005, and has been piloted extensively. Subsequently, Skye initiated a Feasibility Study to be based on refurbishment and expansion of the existing pyrometallurgical smelter, as well as a new Preliminary Assessment of a hydrometallurgical expansion of the smelter project to be described in the Feasibility Study. In the new Preliminary Assessment, initially, the so-called High Pressure Acid Leach (HPAL) Process was compared to Skye’s Sulphation Atmospheric Leach (SAL) Process. It was decided to complete the Preliminary Assessment on the basis that the HPAL Process would be used to process the limonite. Nickel and cobalt recovery was assumed to be by mixed nickel cobalt hydroxide precipitation using magnesia. This recovery option was tested extensively in conjunction with continuous pilot plant testing of the SAL Process. The HPAL process, which has been commercialized at several nickel laterite mines, was tested with Fenix limonite in bench-scale leach tests, and further pilot plant testing is planned for later in 2007. New bulk samples of limonite and transition material from Areas 215 and 217 have been taken for this pilot plant program.

This technical report deals with the project to recover nickel from the saprolite at Fenix by conventional pyrometallurgical processing. Description of testwork on hydrometallurgical processing of the limonite is included for the sake of completeness.

16.1.1

Sample Collection 2005

Approximately 12 wet tonnes of limonite and saprolite samples were taken from mine areas 212-1 and 213-1 on 14-16 June, 2005. The sampling was supervised directly by D. Neudorf and C. McKenzie of Skye Resources (Skye 2005).

16.1.1.1

Procedures

Sampling sites were selected by reference to mine and exploration assay data, face mapping and visual inspection of the exposed limonite and saprolite in the mine faces. The sites were first cleared of detritus and then sampled by back-hoe.

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Sample bags were loaded directly by the back hoe – shovel by shovel – at the sampling sites, using a large funnel fabricated previously. While laborious, this procedure eliminated the possibility of inadvertent sample contamination. The full bags were then loaded into a dump truck for transport to the plant site, where the bags were stored temporarily inside the maintenance shop. On 17 June, the day after completing the sampling, the sample bags on pallets were loaded into a single 40 foot container for transport to Puerto Santo Tomas and shipment to Canada. The shipment left Santo Tomas for Kingston, Jamaica and Halifax on 22 June.

Weather conditions at the time of sampling (14-16 June) were partly sunny skies, hot and humid. No rain fell during the actual sampling; however, the area experienced a very heavy rainfall overnight from 13-14 June (2.25 inches of rain were recorded at the plant site) and another heavy rainfall the evening of 15 June. The rain made equipment movement difficult at some of the sampling sites, however, the sampled material appeared to be relatively dry, especially that which was sampled from near-vertical bench faces. After sampling, pits were refilled with excess material and at bench-face sample sites the excess material was pushed up against the sampled face.

Sites were photographed prior to and during sampling, and after back-filling.

16.1.1.2

Sample Descriptions

The sample locations and qualitative descriptions are given in Table 16-1. In addition to the samples described here (PP9-12), samples PP6 (2 bags), PP7 (3 bags) and PP8 (2 bags) were taken on 9 and 10 May, 2005 under the direction of Luis Menegazzo. PP6 and PP8 were taken from the same pits dug earlier for samples PP1 and PP4, respectively, in December 2004. PP7 was taken from another exposed saprolite pinnacle about 10 m to the south of the PP2 sample site.³

The weights of the bagged saprolite samples ranged from 732 kg to 874 kg. The weights of the bagged limonite samples ranged from 630 kg to 807 kg.

The samples were intended to be typical saprolite and limonite material; however the samples are not necessarily representative of the expected run-of-mine feed.

³	An initial phase of sampling was conducted in 2004 when samples PP1 to 4 were taken. The sample sites were re-sampled in 2005 to obtain additional material.

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Table 16-1: Description of Pilot Plant Samples

Sample

No.

Area

Ore Type

Sample Site Coordinates

Description

Easting

Northing

Elevation

PP9

213-1

Saprolite

242,605

1,716,317

114 m

2 bags of mostly rocky saprolite in an earthy, ferruginous matrix. This sample was taken from depth in the south-west corner of the existing pit from which samples PP1 (Dec 3, 2004) and PP6 (May 10, 2005) were taken.

PP10

212-1-N

Saprolite

243,476

~ 1,717,852

399 m

3 bags from exposed pinnacle of mostly well-weathered yellowish, earthy saprolite. This pinnacle was located ~ 10 m south of the pinnacle from which sample PP2 was taken. Also, sample PP7 (May 10, 2005) was taken from the same spot as PP10.

PP11

212-1-N

Limonite

243,480

1,717,766
and

~ 1,717,746

399 m

11 bags of “lower” limonite (ferruginous saprolite) from an existing bench face, located between two saprolite pinnacles, rock textures evident (same site as PP3 taken Dec 3, 2004), and 3 bags of similar material taken from the same bench face about 20 m to the south. There were streaks of whitish-beige material visible in the face after cleaning.

PP12

212-1-S

Saprolite

~ 243,619

1,717,672

398 m

2 bags of very rocky saprolite in a hard, slightly ferruginous matrix. The sample site was immediately adjacent to the western side of the PP4 sample site. Sample PP8 (May 9 and 10, 2005) was taken from the PP4 sample pit.

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16.2

Pyrometallurgical Testwork

The feasibility of smelting the saprolite from the CGN property has been demonstrated through the commercial operation of the Exmibal facility. There is a proven data base for smelting saprolite over a 3 year period.

The process development history for the Exmibal facility is documented by Toomver 1979. The following sections are a summary of T.T. Toomver’s paper as it relates to Exmibal.

16.2.1

Summary

In the 1950s and 60s, Inco’s Process Research Department tested a variety of different nickel laterite ores in a range of extraction processes, pyrometallurgical, hydrometallurgical and combination of both pyro- and hydro-metallurgical. The pyrometallurgical processing route consisting of drying, selective reduction and smelting was selected for processing Inco’s Guatemalan ore. The selection process involved extensive theoretical, laboratory and pilot plant work to prove the process metallurgy and provide scale-up parameters for commercial design.

16.2.2

Physical Ore Beneficiation

Tests including gravity and size separation, magnetic and electrostatic separation and froth flotation were performed on the ores to determine if the nickel content could be concentrated before processing. While simple rejection of coarse size fractions was found to work on other ores tested, none of the physical upgrading tests were successful on the Guatemalan ores.

16.2.3

Selective Reduction

Various pyrometallurgical processes were tested and selective reduction (with and without sulphiding) followed by electric furnace melting was found to be the most desirable for the Guatemalan ores.

Selective reduction was developed using a 1.5 m diameter by 12.2 m long reduction kiln at Inco’s Research Stations in Port Colborne, Ontario. The targeted process involved heating the ore in a reducing atmosphere to selectively reduce first the nickel oxides and then the iron oxides to metallic Ni and Fe such that a matte or ferronickel grade of 30% would be produced in the subsequent smelting step.

To design the commercial reduction kilns for Exmibal and other Inco laterite projects at the time, a wide variety of kiln configurations, fuels, reductants and sulphur addition methods were tested over an 11 year test period.

16.2.4

Smelting

The smelting step involved melting of the partially reduced ore, completing reduction with any residual carbon and separating the metal/matte phases from the slag. Inco tested three vessels for the smelting step, a tilting electric furnace, a rotary melting kiln and a submerged arc furnace.

Electric furnace smelting was selected for commercialization because of its proven reliability and the good metallurgical results obtained from the 300 kW test furnace operated at the research station. Both the production of matte and ferronickel were successfully tested using Guatemalan ore. A sample of some of the metallurgical results obtained during the production of FeNi from Guatemalan ore is shown in Table 16-2.

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Table 16-2: Metallurgical Results During Production of Ferronickel from Guatemalan Ore


Item	Parameter	Value
Reduced Calcine	% Ni	2.43
	% Fe	26.4
	% C	0.3
Slag	Temperature (°C)	1560
	% Ni	0.14
	% Fe	25.1
Ferronickel	Temperature (°C)	1520
	% Ni	30.1
	% Fe	67.1
	% C	0.08
Energy Loss	% of power input	18
Power Required	Total kWh per tonne ore	425
Nickel Recovery	%	94.7

Although ferronickel was successfully produced in the test facility, Inco ultimately selected matte production for the Exmibal project. Additional testwork was undertaken to verify the converting step required to remove iron from the electric furnace matte.

The development, construction and first year of operation of the Exmibal project are described by Sopko, 1979. Mobilization in Guatemala started in July 1974 and was essentially completed by the end of 1976. Nominal plant capacity was 11,300 tonnes of nickel per year (25 Mlbs). The plant was officially inaugurated in July 1977 and a first shipment of 600 tonnes was made in February 1978.

Loan agreements required that a performance test be carried out to confirm the plant capacity and recovery. This test was to last 90 days and to achieve 90% of nominal capacity using ore at 2.10% Ni. A first performance test was initiated in September 1979 but was terminated after 56 days due to a labour dispute. A second, successful, performance test was carried out from 5 May to 2 August, 1980. This test achieved 106% of its target (i.e. about 97% of nominal capacity) with ore at 2.09% Ni and a recovery of 87.4%.

Overall, Exmibal produced some 11,000 tonnes of nickel from off and on operation over three years.

Based on the successful pilot plant production of ferronickel from the Fenix saprolite, the successful commercial production of nickel matte from the Fenix saprolite in the commercial plant, and the successful commercial production of ferronickel from saprolite ores in a number of other, similar rotary kiln electric furnace smelters, it is believed that the refurbishment and conversion of the commercial plant to produce ferronickel from the same saprolite feed can be commercialized without additional testwork.

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17.

Mineral Resource and Mineral Reserve Estimates

17.1

Mineral Resource Estimates

Information in this section has been previously disclosed in a Technical Report dated 4 July 2006 (Snowden 2006) and is included in a modified format to reflect revised cut-off grades for resource reporting. Subsequently, Snowden has also updated resource estimates in area 217 based on 35 m in-fill drilling completed in 2006. The revised resource estimates for this area have not been incorporated in the mine schedule or mineral reserve estimates.

Mineral resources are reported inclusive of mineral reserves.

17.1.1

Areas 212, 213, 215, 217, 251

The mineral resource estimates prepared by Snowden are presented at single cut-off grades of 0.8% Ni in Table 17-1 for transition-saprolite; and at 1.0% nickel equivalent (NiEq) in Table 17-2 for limonite.

In Table 17-1 Fe, SiO₂ and MgO are reported to characterize the chemistry of the saprolite and as a guide to the suitability of the material for pyrometallurgical processing. With the exception of a credit for iron in ferronickel, they do not have economic value as recovered products.

In Table 17-2 Fe and MgO are reported to characterize the chemistry of the limonite and as a guide to the suitability of the material for hydrometallurgical processing. They do not have economic value as recovered products.

The nickel equivalent grade is only applicable to the limonite mineral resource and not to the saprolite plus transition resources or to the mineral reserves.

•

Nickel equivalence (NiEq) is based upon: the formula NiEq = Ni + (3*Co) where:

•

Ni metal price of $US5 per pound is assumed.

•

Co metal price of $US15.31 per pound is assumed.

•

Ni recovery of 93.9% is assumed.

•

Co recovery of 92% is assumed.

The above metal prices were derived from a historical review conducted by Skye and Gander Consulting LLC. Recoveries are estimated from hydrometallurgical testwork and summarized by Skye (D. Neudorf, pers.comm).

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Table 17-1: Mineral Resource Estimates of Transition and Saprolite at 0.8% Nickel Cut-off for Selected Areas


Area	Tonnes (000’s), Measured	%Ni	%Fe	%SiO₂	%MgO	SiO₂/MgO
212	6,700	1.71	16.3	34.0	23.0	1.48
213	950	1.78	14.7	36.0	22.7	1.59
215	100	1.50	15.3	34.6	24.3	1.42
217	16,920	1.46	17.0	34.0	22.3	1.54
251	0	0.00	0.0	0.0	0.0	0.00

Total Measured	24,670	1.54	17.0	33.5	22.5	1.53

Area	Tonnes (000’s), Indicated	%Ni	%Fe	%SiO₂	%MgO	SiO₂/MgO
212	470	1.58	17.2	33.4	22.2	1.50
213	600	1.31	13.2	38.0	23.4	1.62
215	10,290	1.44	17.3	33.3	22.8	1.46
217	5,960	1.40	17.8	33.0	21.5	1.54
251	19,460	1.48	16.7	33.7	23.5	1.43

Total Indicated	36,780	1.45	17.0	33.5	23.0	1.46

Area	Tonnes (000’s), Measured + Indicated	%Ni	%Fe	%SiO₂	%MgO	SiO₂/MgO
212	7,170	1.70	16.4	34.0	22.9	1.48
213	1,550	1.59	14.1	36.7	23.0	1.60
215	10,390	1.45	17.3	33.3	22.8	1.46
217	22,880	1.45	17.2	33.8	22.1	1.53
251	19,460	1.48	16.7	33.7	23.5	1.43

Total Measured + Indicated	61,450	1.49	16.9	33.8	22.8	1.48

Area	Tonnes (000’s), Inferred	%Ni	%Fe	%SiO₂	%MgO	SiO₂/MgO
212	400	1.35	11.7	37.5	27.3	1.38
213	40	1.01	11.0	39.0	25.6	1.52
215	19,770	1.31	19.0	32.4	20.3	1.60
217	22,800	1.19	17.8	32.7	20.6	1.59
251	0	0.00	0.0	0.0	0.0	0.00

Total Inferred	43,010	1.25	18.3	32.6	20.5	1.59

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Table 17-2: Mineral Resource Estimates of Limonite at 1.0 % Ni Equivalent Cut-off (NiEq = Ni + 3Co) for Selected Areas


Area	Tonnes (000’s), Measured	NiEq	%Ni	%Co	%Fe	%SiO₂	%MgO
212	2,570	1.42	1.13	0.097	43.7	6.5	0.9
213	40	1.71	1.32	0.132	44.4	5.7	0.9
215	8,670	1.39	1.09	0.100	44.1	5.5	1.1
217	16,510	1.43	1.10	0.110	44.4	6.0	1.1
251	16,530	1.50	1.15	0.118	45.2	4.8	1.0

Total Measured	44,320	1.45	1.12	0.110	44.6	5.5	1.0

Area	Tonnes (000’s), Indicated	NiEq	%Ni	%Co	%Fe	%SiO₂	%MgO
212	6	1.21	0.96	0.084	37.2	9.1	0.9
213	—	—	—	—	—	—	—
215	950	1.34	1.05	0.096	43.0	6.3	1.1
217	1,250	1.28	0.99	0.100	43.0	7.7	1.1
251	—	—	—	—	—	—	—

Total Indicated	2,206	1.31	1.02	0.098	43.0	7.1	1.1

Area	Tonnes (000’s), Measured + Indicated	NiEq	%Ni	%Co	%Fe	%SiO₂	%MgO
212	2,580	1.42	1.13	0.097	43.6	6.5	0.9
213	40	1.71	1.32	0.132	44.4	5.7	0.9
215	9,620	1.38	1.08	0.100	44.0	5.6	1.1
217	17,760	1.37	1.09	0.110	44.3	6.1	1.1
251	16,530	1.50	1.15	0.118	45.2	4.8	1.0

Total Measured + Indicated	46,530	1.42	1.11	0.110	44.5	5.6	1.1

Area	Tonnes (000’s), Inferred	NiEq	%Ni	%Co	%Fe	%SiO₂	%MgO
212	—	—	—	—	—	—	—
213	—	—	—	—	—	—	—
215	9,570	1.28	1.03	0.085	41.2	7.1	1.1
217	5,130	1.22	0.97	0.080	42.0	8.5	1.1
251	—	—	—	—	—	—	—

Total Inferred	14,700	1.26	1.01	0.083	41.5	7.6	1.1

17.1.2

Other Areas

The mineral resource estimates for saprolite at other deposits within Skye’s licences are presented in Table 17-3. These deposits are located at areas 210, 211, 214, 215, 216, 218, 219, 245, 260 (Amate), Montúfar and Mantos 5-10.

Table 17-3: Mineral Resource Estimates – Other Saprolite Deposits


Saprolite Mineral Resources	Tonnes (000’s)	% Ni	Cut-off (% Ni)
Measured	8,710	1.79	1.6	%
Indicated	27,060	1.82	1.6	%
Measured + Indicated	35,770	1.81	1.6	%
Inferred	48,220	1.64	1.6% and 1.0	%

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Limonite also occurs in these areas and is reported at a 1.0% nickel equivalent cut-off in Table 17-4.

Table 17-4: Mineral Resource Estimates – Other Limonite Deposits


Limonite Mineral resources	Tonnes (000’s)	% Ni	%Co	Cut-off (NiEq)
Measured	2,300	1.10	0.093	1.0
Indicated	10,300	1.20	0.101	1.0
Measured + Indicated	12,600	1.18	0.100	1.0
Inferred	32,800	1.15	0.095	1.0

Note: NiEq = Ni + (3*Co)

17.1.3

Disclosure

Mineral resources reported in Section 17.1.1 were prepared by Snowden under the technical supervision of Andrew Ross, MSc., FAusIMM, (CP), P.Geo. Both Mr. Ross and Snowden are independent of Skye.

Mineral resources reported in Section 17.1.2 were compiled by Dr. Paul Golightly, PhD, P.Geo of Golightly Geoscience. Dr. Golightly is independent of Skye. Limonite mineral resources were estimated by Snowden using only the historic Inco data.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.

17.1.3.1

Known Issues that Materially Affect Mineral Resources

Snowden and Golightly Geoscience do not know of any issues that materially affect the mineral resource estimates. These conclusions are based on the following:

•

Environmental

The Fenix Exploitation Licence has an approved Environmental Assessment, Ministerio de Ambiente y Recursos Naturales Resolución No. 0190 - 2006. The Montúfar exploration licence has an approved Exploration Mitigation Plan. An Environmental Impact Assessment for the Process Plant and supporting infrastructure were approved by the Ministerio de Ambiente y Recursos Naturales on June 7, 2007.

•

Permitting

Skye has represented that both licences are in good standing, and that a construction permit from the Municipality of El Estor has been received.

•

Legal

Skye has represented that there are no outstanding legal issues; no legal action, and injunctions pending against the Project.

•

Title

Skye has represented that the mineral and surface rights have secure title. All mineral resources and mineral reserves estimated for the feasibility study are within the Fenix exploitation license and on CGN-controlled land.

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•

Taxation

No known issues.

•

Socio-economic

Skye has represented that the Project has strong local community support.

Access for exploration of mineral resources outside CGN’s surface rights but within its exploitation rights has not been achieved as CGN continues to consult with local communities and landowners to build a climate of trust and transparency. CGN sponsors a community development foundation in the area and has an active and continuing community relations initiative to build the good neighbour relationships needed for future access to the mineral resources required to extend the Project beyond 30 years.

•

Marketing

No known issues.

•

Political

Skye has represented that the current government is very supportive of the Project.

•

Other Relevant Issues

No known issues.

17.1.3.2

Factors that Materially Affect Mineral Resources

Snowden and Golightly Geoscience do not know of any factors that materially affect the mineral resource estimates. These conclusions are based on the following:

•

Mining

Exmibal successfully mined the limonite and saprolite deposits at Areas 212 and 213 in the period 1977 to 1980. The landforms and deposits in the other areas are generally similar.

•

Metallurgical

Exmibal successfully treated saprolitic material in the Process Plant to produce nickel matte (Section 16.2).

There are existing hydrometallurgical processes to treat limonitic material and these were assessed by Hatch for treatment of the Fenix deposits.

•

Infrastructure

No known factors.

•

Other Relevant Factors

No known factors.

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17.1.4

Assumptions, Methods and Parameters – Snowden Estimates

The Snowden estimates for areas 212, 213, 215, 217 and 251 were prepared in the following steps:

•

Data validation – this is discussed in Section 14.6.

•

Data preparation – this and subsequent steps are discussed below.

•

Geological interpretation and modeling.

•

Establishment of block models.

•

Compositing of assay intervals.

•

Transformation of composites.

•

Exploratory data analysis of the key constituents – Ni, Co, Fe₂O₃, MgO, SiO₂, Density (DBD).

•

Analysis of boundary conditions.

•

Variogram analysis.

•

Derivation of kriging plan.

•

Grade interpolation.

•

Validation of grade estimates.

•

Classification of estimates with respect to CIM guidelines.

•

Resource tabulation and resource reporting.

17.1.4.1

Software

The following specialist software was used in the generation of Mineral Resource estimates:

•

Gemcom software (GEMS version 5.55) was used to: conduct additional validation checks on the borehole database provided by Golightly Geoscience; prepare the database for the resource estimation process; generate 3-D wireframe surfaces for the different lithological layers; block model validation; resource classification and resource reporting; plotting and figure generation.

•

Datamine Studio (version 2.1) was used for the bulk of the resource estimation work including: desurveying of boreholes; block model generation, coding and validation; grade interpolation; resource classification and reporting for validation purposes; and model preparation for a mining study.

•

Snowden’s Supervisor (version 7.10) software was used for exploratory data analysis, declustering, topcut analysis, variography and model validation.

•

Snowden’s Utilisor (version 1.4) software was used for model validation.

•

Microsoft Excel (2002) was used extensively, in a supplementary role to the above software, for calculations, plots and validation. Various GSLIB programs (e.g. kt3d, declus) were used for statistical testing.

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17.1.4.2

Data Preparation

Redundant Data

A process to identify redundant boreholes was completed, since:

•

Boreholes and pits are frequently co-located; i.e. may be located within four metres of each other.

•

The laterite profile was found to be incompletely sampled by certain boreholes or pits in the historical programs and may have been superseded by recent boreholes.

Any grouping of boreholes or pits within four metres proximity was therefore evaluated to identify the highest quality and representative method of sampling. In terms of sample quality, the earliest hand-auger holes are superseded by Winkie auger, then McKinney ‘Sampling Tube’, tractor-mounted auger, pits and finally diamond core drilling.

The redundant boreholes or pits are identified in Appendix D of Snowden 2006 by the colour fill.

17.1.4.3

Heads and Tails

In the case where auger boreholes were superseded by pit samples, a check was made to ensure that deeper auger samples were not removed from the database. Deeper sample intervals from co-located auger holes were thus integrated as ‘tails’ to pit records.

In the case where CGN core holes were collared in excavations and superseded auger samples or pits, a process was undertaken to ensure that sample data from the ‘mined’ section was integrated as ‘heads’ to the CGN record. This was done so that pre-mine estimates were based on the most reliable data.

All geological interpretation and subsequent grade estimates were based on a database that excluded redundant borehole or pits records, and included ‘heads’ and ‘tails’ information to ensure, where possible, sampling across the full laterite profile.

17.1.4.4

Desurveyed Sample File Preparation

Sample assay and borehole collar files were imported into Datamine Studio. A desurveyed sample file was generated from the data following basic validation checks. No down-hole survey information was needed for the desurveying process as all of the boreholes are vertical and only occasionally exceed 35 m depth.

17.1.4.5

Geological Interpretation and Modeling

•

Method

The layer codes in the Gemcom borehole database for cover, limonite, transition, saprolite and bedrock were used to construct wireframes of the laterite units in the following way:

Simple triangulation of layer elevations for areas 212, 213, 215 and 217.

At Area 251, the borehole collars were ‘pressed’ to register with the LAND INFO topographic surface (DTM). Layer thicknesses were interpolated by inverse distance weighting into a two dimensional grid model. The final layer wireframes at 251 thus resulted from an integration of the land form and layer thickness.

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•

Bedrock

The bedrock position was either intersected in boreholes or, in the case of incomplete drill penetration, given an assumed elevation. The assumed elevation was either the end of the borehole or the predicted depth (Section 10.6.2.1).

•

Modifications to Layer Wireframes

The wireframes were inspected in cross-section view to check for any inconsistencies or overlaps. In several instances where boreholes had not intersected bedrock, the bedrock wireframe was adjusted locally to account for off-section information that indicated a deeper repositioning was realistic. Any additional volumes of saprolite that resulted from this process were subsequently classified as ‘Inferred’.

•

Additional Wireframes

Surface topography wireframes for each area were generated by simple triangulation of the borehole collars with the exception of area 251, where the LAND INFO DTM was used.

Excavations in Areas 212 and 213, and stockpiles on Area 213 are identified by wireframes.

•

Model Coding

The Gemcom surface, excavation, stockpile and layer wireframes were used to code block models in Datamine’s Studio software.

17.1.4.6

Block Model Set-up

Block models were generated using the geological model wireframes provided. Wireframes were filled with regular sized blocks and no sub-celling was undertaken as this is difficult to handle in z-transformations. Block sizes were 12.5 m (X) x 12.5 m (Y) x 2.5 m (Z) for Area 212, and 25 m (X) x 25 m (Y) x 2.5 m (Z) for all other areas. Resulting block models were coded according to geological layer and the 50% in-out rule. The ensuing block model volume for each layer was compared to the wireframe volume by way of validation. The block model volume was found to be within ±1% of the wireframe volume, indicating good volume representation.

17.1.4.7

Compositing of Assay Intervals

The raw sample data in the desurveyed borehole file displayed a range of different sample lengths and consequently needed to be composited to similar intervals to ensure consistent sample support during estimation. A composite interval of 1.0 m was considered suitable following analysis of the raw data on an all data and on a by-layer basis. Density weighted compositing was conducted within lithological layers to account for the variable nature of the density (particularly important in the transition and saprolite layers).

The compositing was conducted in such a way so as to ensure that all samples were used whilst keeping the compositing interval as close to 1.0 m as possible: between 98-99% of the composites in the limonite layer, 82-97% of the composites in the transition layer, and between 95-99% of the composites in the saprolite layer occur in the 0.9-1.1 m interval.

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17.1.4.8

Transformations

The undulating nature of the topography and the contacts between the different layers complicates grade estimation in real space. Multiple small search volumes of multiple orientations would normally be required on a localized basis to ensure that grade interpolation is conducted using borehole information from the relevant part of the profile. Whilst theoretically possible, this is extremely time-consuming and difficult to validate. Consequently it was decided to transform the data into horizontal units of equal thickness. This transformation was conducted on the elevation data only (z-coordinate), thereby retaining the spatial relationships between the data in Northing and Easting (y-, and x-coordinates, respectively). The transformation was performed by layer for each of the five areas in the following way:

•

Data composites were selected for the layer of interest (e.g. limonite data for the limonite layer etc.).

•

The maximum thickness of this layer was determined from the selected borehole data.

•

The original z-coordinate was stored in a new field (Z_ORIG).

•

All borehole data (for each relevant layer) were “stretched” over the maximum thickness, whilst retaining the original composite length – this process effectively inserted gaps into those boreholes for which the layer thickness was less than the maximum thickness, with the gap length a function of layer thickness, original composite length and number of samples in the borehole.

•

A nominal Elevation of 0 m was selected for the top of the profile, and sample FROM and TO values were transformed accordingly, resulting in three fields being changed during the transformation – Z, FROM and TO.

•

All transformed borehole composites from the relevant layer and area were visually validated.

•

The transformation of the z-coordinate was conducted following the exploratory data analysis, as part of the preparation for grade estimation. The transformed data was used for variography analysis and estimation.

17.1.4.9

Exploratory Data Analysis

Exploratory data analysis was conducted on the composited borehole data for each of the five areas. The analysis was focused on the Ni, Co, Fe₂O₃, SiO₂, MgO and density (DBD) data in the limonite, transition and saprolite layers, and the following aspects relevant to the resource estimation process were investigated for each constituent:

•

Inspection of aerial distribution plots to identify anomalies. The plots in Appendix A of Snowden 2006 represent length- and density-weighted averages for each borehole, facilitating the two dimensional interpretation of aerial trends.

•

Inspection of vertical grade trends to identify anomalies. The plots in Appendix B of Snowden 2006 are vertically stacked grade versus depth profiles and provide the best representation of the average down-hole grade profile through the limonite, transition and saprolite layers in each area.

•

Correlation of data. Correlation matrices were determined for Ni, Co, Fe₂O₃, MgO and SiO₂ data in the limonite, transition and saprolite layers. Density was determined on the basis of its relationship to Fe₂O₃, and was therefore not included.

•

Characterization of the grade populations to identify statistical outliers and to identify appropriate methods of grade interpolation.

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Based on the results of the exploratory data analysis, Snowden elected to:

•

Apply topcuts to the MgO and SiO₂ data from the limonite layer for all five areas.

•

Not to apply topcuts to any of the other grade data (Ni, Co and Fe₂O₃).

•

Use ordinary kriging to interpolate Ni, Co, Fe₂O₃, MgO, SiO₂ and Density values by layer into block models for each of the five areas.

•

Restrict the search ellipse (to be used in the grade interpolation) in the vertical direction due to the presence of down-hole trends in some of the variables.

17.1.4.10

Topcuts

Topcuts were applied to MgO and SiO₂ grade data from the limonite layer for all five areas after assessing the impact of outlier values on the coefficient of variation (COV = standard deviation/mean). The topcuts were assessed using, and applied to, the composited data, and are presented in Table 17-5 and Table 17-6.

Table 17-5: Topcuts Applied to MgO Data Composites


Grade Component	Area	Topcut (%)	Percentile
MgO	212	1.05	93.6
	213	1.07	90.8
	215	2.38	96.7
	217	1.80	97.3
	251	1.60	95.9

Table 17-6: Topcuts Applied to SiO₂ Data Composites


Grade Component	Area	Topcut (%)	Percentile
SiO₂	212	14.00	95.2
	213	9.00	93.1
	215	11.20	95.1
	217	16.61	97.3
	251	9.94	95.2

17.1.4.11

Boundary Conditions

Limonite was estimated using only limonite borehole data. Nickel and cobalt estimates in the Transition and Saprolite Layers used both transition and saprolite borehole data (i.e. a soft boundary), due to the closer similarity in these two layers than between transition and limonite, and due to the thin nature of the transition layer and limited number of data in this layer, and the continuation of grade trends from this layer into the saprolite.

Iron, density, silica and magnesia estimates in the Transition and Saprolite Layers honoured hard boundary conditions, where only transition composites informed the Transition blocks and only saprolite composites informed the Saprolite blocks.

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17.1.4.12

Variogram Analysis

Variography was conducted on the uncut z-transformed composite data for Ni, Co, Fe₂O₃, SiO₂, MgO and Density for the limonite and combined transition and saprolite layers in a given area. The transition and saprolite layer data was combined for practical reasons due to the limited number of data available in the relatively thin transition layer. The greater degree of similarity between the transition and saprolite layers than between the transition and limonite layers resulted in the decision to combine the former rather than the latter for continuity analysis purposes.

Continuity was modeled in three dimensions. Horizontal continuity directions were generated for each of the variables of interest using variogram fans in the horizontal plane. The variogram in this direction was fitted with a mathematical model, consisting of a nugget and generally two spherical structures. Continuity directions and traditional variograms were iteratively modeled using the z-transformed data, log-normal z-transformed data, and normal scores of the z-transformed data. Lag distances were selected on the basis of drill spacing in the layer of interest (with a lag tolerance of half the lag distance). Angular tolerances of between 10° and 15° were used, depending on data density. Given the relatively tight angular tolerances used and the average borehole spacing in each of the five areas, it was decided not to use a bandwidth restriction on the search for suitable data pairs. The modeled major, semi-major and minor directions were used to generate a three dimensional continuity ellipse.

Variogram fans and models for each variable for each of the three layers were generated for each of the five areas.

Based on the correlation results from the exploratory data analysis, and the results of the variography, it was decided to use the variogram parameters in the following way for grade estimation:

•

Ni variogram parameters (for the relevant layers) were used to model Ni continuity in all areas.

•

Fe₂O₃ variogram parameters (for the relevant layers) were used to model Fe₂O₃ and Density continuity in the limonite layer for all areas; and Fe₂O₃, Co and Density continuity in the combined transition and saprolite layers for all areas.

•

Co variogram parameters were used to model Co continuity in limonite in all areas.

•

SiO₂ variogram parameters (for the relevant layers) were used to model SiO₂ and MgO continuity in the limonite, and in the combined transition and saprolite layers.

•

Due to the limited number of data available for the limonite layer in Area 213, limonite variogram parameters for Ni, Co, Fe2O3, MgO, SiO2 and Density from Area 212 were assumed for Area 213.

17.1.4.13

Kriging Plan

Kriging neighbourhood analysis (KNA) was conducted on all areas, confirming block size, discretisation of 3x3x3, and minimum-to-maximum sample restriction of 12 to 20.

17.1.4.14

Grade Interpolation

Grade estimation was conducted in z-transformed space using ordinary kriging with a restricted search to minimize the impact of the down-hole trends noted in several of the components (particularly in the transition layer).

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An expanding search method was selected:

•

The first pass concentrated on a distance of 75 m x 75 m x 5 m.

•

The search was expanded to between ³/4 variogram range and variogram range for the second search pass.

•

The final search was expanded to many times the variogram range on the third search pass to fill all blocks (grades in search volume 3 to be classified inferred at best).

A maximum number of samples (3) per borehole constraint were used for Fe₂O₃, SiO₂ and MgO grade interpolations.

Block data was back-transformed to original space following estimation.

Several grade estimation runs were conducted for each area – nearest neighbour (NN), inverse distance weighting (IPD², IPD³) and ordinary kriging with variable search volume parameters (wider and less restrictive searches to tight searches) for sensitivity testing.

17.1.4.15

Model Validation

Block models were validated in several ways:

•

Visual inspection of block grades against input data composites in cross-section.

•

Model validation slices by northing, easting and elevation whereby grade trends in the model are compared with grade trends in the topcut and declustered input data. (Appendix I of Snowden 2006 contains an example for all constituents in Area 212; Ni, Co, density in limonite and Ni, density in saprolite in the other deposits).

•

Comparison of global statistics of block model (by layer and area) with declustered composite data used as input, with topcuts applied to SiO₂ and MgO for limonite (refer to Appendix J of Snowden 2006).

•

Reconciliation of exploration model to historic production data in area 212, blocks 1 and 5 (Snowden 2006a).

•

Checks on the degree of nickel grade smoothing were conducted by a uniform conditioning method and found to be adequate (Snowden 2006b).

17.1.4.16

Resource Classification

The Mineral Resource estimates were classified as Measured, Indicated or Inferred based upon the spacing of boreholes and an assessment of the continuity in nickel grade and thickness (Snowden 2005, 2006c).

Snowden completed a borehole spacing study on the area 212 data with the aim of the analysis to determine minimum borehole collar spacing that would be required to classify estimates from drilling as Indicated and Measured Mineral resources for the transition-saprolite and limonite layers. Snowden considers that an appropriate error for Indicated Mineral resources is ±15% at 90% confidence limits for annual production increments and for Measured Mineral resources the error should be ±15% for quarterly production increments.

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For this study Snowden used conditional simulation of thickness, and metal accumulation (the product of nickel grade, thickness, and density) to create detailed two-dimensional point datasets conditioned to the known data, which honoured the histogram and variogram as if sample information was available at a very close spacing.

The study assumed a production rate of 1.5 Mt of saprolite material per year, an average thickness of the simulation area of 9.5 m, and average density of 1.35 t/m³. One 100 mE by 100 mN panel is equivalent to one month of production.

Although the actual values for the grid nodes represented by the simulations remained unknown (except where actual data is coincident), an exhaustive sample model on a very detailed grid is valuable for the analysis of likely estimation error associated with different drilling patterns and spacing. Two dimensional panel grades estimated by kriging sample grids extracted from the simulation on different patterns, were compared to the known ‘true’ panel grades determined by averaging the exhaustive data in each panel from the simulation. Using this model of reality, the likely estimation error for different borehole patterns and spacings was quantified. Additionally, estimation of panels of different sizes allowed for analysis of the effect of change of support.

The findings from this study were:

•

The accumulation estimation error in the transition-saprolite layer for quarterly production increments is within ±15% at 90% confidence limited when the borehole spacing is 35 m or better. Snowden recommends classifying areas of the transition-saprolite drilled to 35 m centres or better as Measured Mineral resources.

•

The accumulation estimation error in the transition-saprolite domain for annual production is within ±15% at 90% confidence limited when the borehole spacing is 100 m or better. Due to some uncertainty regarding bias in auger drilling, the nature of small scale thickness variation, and the reliability of results at 100 m spacing, Snowden recommends classifying areas of the transition-saprolite drilled to 75 m centres or better as Indicated Mineral resources.

•

Comparison of the transition-saprolite and limonite variograms for accumulation indicates that the drilling spacing requirements for limonite can be relaxed to 75 m for Measured and 100 m for Indicated. Snowden also considers that the transitional nature of the basal limonite contact would reduce both thickness and accumulation variability.

The drill spacing patterns for each area were reviewed and polygons were generated to delimit pattern spacings of 35 m, 75 m and 100 m. The block models were then coded to identify limonite blocks within the 75 m and 100 m polygons as Measured and Indicated classes respectively; and to identify transition and saprolite blocks within the 35 m and 75 m polygons as Measured and Indicated classes respectively.

In practice the areas of Measured transition and saprolite are drilled at spacings of 25 m or closer.

Inferred limonite mineral resources were identified where the drill spacing was greater than 100 m.

Inferred saprolite mineral resources were identified where the drill spacing was greater than 75 m or where saprolite mineralization was interpreted beneath the limit of drilling.

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17.1.4.17

Resource Reporting

All estimates and reports were prepared on a ‘dry tonnage’ basis.

The classified Mineral Resource estimates were reported from Datamine Studio and the tabulations checked by re-reporting from Gemcom software. Areas 212 and 213 have been subject to mining and partial reclamation - consequently the block model resource reports reflect the current post-mining topography exclusive of any reclaimed areas such as Cerro 400.

A stockpiled resource at area 213 is not included in the resource tables and represents an additional 46,000 tonnes of saprolite (C. McKenzie, pers.comm.).

17.1.5

Assumptions, Methods and Parameters – Previous Estimates on Other Areas

The Mineral Resources reported in Table 17-3 were estimated using the Laterite Evaluation System (LES) by Inco in 1998 and subsequently reviewed and re-classified by Golightly Geoscience in 2005 and disclosed in Hatch, 2005.

The LES method is a 2-D inverse squared distance method developed by Inco, and includes a dilution allowance but not a “mineability allowance” of about 85% that was “based on mining experience”.

The LES program uses an arbitrary search radius to select samples for estimating each cell. Declustering is accomplished by quadrant averaging and screening far data by near in each quadrant. A constant density was used for saprolite. The following constraints were used for saprolite estimates:

•

Limonite (defined as Fe>36%) was treated separately.

•

A constant density for saprolite (1.08 t/m³) and limonite (0.96 t/m³) was applied.

•

Samples from test pits, rotary and auger holes were used.

•

Deposit boundaries are the same as CGN’s 1980 MRI and subsequently used for estimates in 1998.

•

Cell size was by zone at around ¹/2 borehole spacing, and the search radius two times that spacing. Where the borehole spacing is 25 m as it is in most of the deposits in areas 212, 213 & 217, this is consistent with a variogram based range.

•

A minimum ore thickness (MOT) of 2 m and an initial borehole/pit grading cut-off of 1.6% nickel were used to define intercepts down the holes and refined iteratively or manually accommodating additional material down to a preset minimum grade to resolve the problem of multiple intercepts and give representation to low grade zones within the deposit outlines.

•

Contact dilution was incorporated (0.25 m at the top and 0.5 m at the bottom).

The limonite mineral resources in Table 17-4 have been estimated and classified by Snowden using Inco data and a 2-D estimation method similar to the LES method.

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17.2

Mineral Reserve Estimates

17.2.1

General

The Mineral Reserve estimate was re-evaluated to reflect updated process costs, mine costs, and nickel price assumptions. A new optimisation was completed applying the new parameters and a pit shell was chosen using the same selection criteria as in the October 2006 Feasibility Study. The new pit shell selected came within 2% of the October 2006 Feasibility Study pit shell indicating that the changes to cost and revenue assumptions had a negligible effect on the size of the pit shell selected. As such no changes were made to scheduling or the Mineral Reserve Estimate.

The selection process used in the October 2006 Feasibility Study is summarized below with an updated table containing the new, updated costs and nickel price assumptions. Had the updated costs and metal price assumptions been within the range of the original sensitivity analysis completed in the October 2006 Feasibility Study a new optimisation would not have been necessary. However the new process costs and the metal price estimates have surpassed the October 2006 Feasibility Study sensitivity analysis and a sensitivity analysis was completed using the new costs and metal price assumptions and is included in this report for further detail and support for the current pit shell selection.

The accelerated process facility ramp-up does not affect mine production rates as mine operations were already estimated at a higher rate and ore stockpiled the accelerated ramp-up will simply reduce the amount of ore stockpiled.

Mineral Reserve estimates reported in this section were prepared by Snowden under the technical supervision of Dick Matthews, Divisional Manager, Mining. Mr. Matthews and Snowden are independent of Skye.

The Mineral Reserve estimate is derived from mineral resource estimates for areas 212, 213,215,217 and 251 as estimated in the October 2006 Feasibility Study. These mineral reserves have not been updated to reflect the increased mineral resources from area 217. All mineral reserves have been appropriately modified to account for:

•

Dilution, contamination, and ore losses.

•

Metallurgical process recovery.

•

Economic evaluation by means of resource optimization, and the scheduling of phased pits in an order to meet processing requirements and blending constraints.

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Table 17-7: Mineral Reserve Design Criteria


Criteria		Explanation
Mineral Resource Estimate for Conversion to Mineral Reserves		The mineral reserve is based on the mineral resource estimate block model and incorporated only measured and indicated mineral resources

Cut-off Grades		Incremental cut-off grades are a function of process plant costs, recovery and metal price. They are used to segregate material within the selected pit shell into mineral and waste. The incremental cut off grade varies from 0.57% to 0.58% for area 251 due to the extra long ore haul.

Mining Factors and Assumptions		Whittle Four-X pit optimization was used to determine economic pit limits. Phased pits were scheduled in a order to meet processing requirements and blending constraints. The ore reserve inventory includes dilution and mineral loss and takes into account practical mining widths, and geotechnical factors. Waste dump locations and haul roads are incorporated into the mine costs and schedule. Selection of a mining method has been subject to an analysis of alternatives, and supplemented with an alternative transportation study.

Metallurgical Factors		Blending constraints, milling throughput, and recovery were provided by Hatch

Cost and Revenue Factors		Mining capital costs were estimated in detail by Snowden based on recent supplier submissions. Overall mining costs were estimated by Snowden from first principles based on recent supplier submissions, and incorporating cost centers/structures and allocations relevant to the project. Processing cost parameters and revenues were provided by Hatch

Others		Analysis has been carried out to assess the sensitivity of the project to metal prices and cost parameters. Mine scheduling has taken into account economic as well as social and environmental considerations.

Market Assessment		Metal price assumptions were provided by Gander Consulting.

Classification		The classification of proven and probable mineral reserves is based on the confidence categories of measured and indicated mineral resources. Mining assumptions commensurate with these levels confidence except for the resource material beyond year 24. Mining after year 24 involves working around backfill and revisiting previously mined out areas and as such all mineral was classified as probable.

Reviews		The dilution and optimization parameters prepared by Snowden have been reviewed by Skye and their consultants. The reserve estimate was also subject to a thorough peer review.

No mining rate limit was applied in the optimization process because the key constraint of the operation will be the process rate limit and mining rates would be adjusted to meet process requirements. No capital costs were considered in order to find break even cut-off grades from only the cash flow of revenue from nickel recovered less the mining, process and associated selling costs.

Thirty-three pit shells were generated using Whittle Four-X software by using 33 revenue factors on the base metal price of $6.70/lb. Revenue factors ranged 0.3 to 1.1 in increments of approximately 0.025. For example, a revenue factor of 1 would equate to a metal price of $6.70/lb and a revenue factor of 0.5 would imply a metal price of $3.35/lb.

Pit shell 11 (revenue factor 0.55) was the optimum pit shell selected, containing approximately 44 million ore tonnes. This pit was selected based on NPV and risk management. Selecting a larger pit would involve mining significantly more waste for relatively minor increases in NPV. For scheduling purposes a staged approach was taken and pit shell 8, containing approximately 33 million tonnes grading 1.70% nickel, was selected as stage 1 so as to ensure higher nickel grades to mill in the earlier years. Stage 2 completes mining from the 44 million tonne pit shell at an average grade of 1.31% nickel. Due to backfilling and access requirements the 44 million ore tonne pit shell has been reduced to 41.4 million ore tonnes mineable. MineMAX software was used interactively to schedule waste stripping and ore movements to the plant and stockpiles to meet the plant feedstock requirements.

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Other financial and operational assumptions used in the pit optimization iterations are presented in Table 17-8. The changes made for the purposes of this report are highlighted.

Table 17-8: Fenix Project Optimization Parameters


Item	Units	Amount
Financial Data
Currency units		US $
Discount rate	(%)	10%
Price – base case – ferro nickel	($/lb)	6.70
Price – base case – ferro nickel	($/10kg)	147.67
Selling – costs (includes royalty, $/tonne metal cost)	($/10kg)	14.17
Production Factors
Dilution and ore loss		Assigned to blocks
Pit slope angle	(degrees)	35
Ore processing limit	(Mtpa)	1.47
Mining rate limit	(Mtpa)	N/A
Processing recovery – nickel	(%)	89.5%
Mining Costs
Base cost (for both ore and waste)	($/tonne)	3.15
MCAF (mining cost adjustment factor)	($/tonne)	Assigned to blocks
Processing Costs
Base Cost	($/tonne)	68.04
Incremental processing ore cost applied to Area 251 ore only	($/tonne)	1.60

17.2.2

Sensitivity Analysis

A sensitivity analysis was completed using the updated operating costs and metal price assumptions. The results are similar to the original feasibility study indicating the project’s value is most sensitive to metal prices and least sensitive to mining costs. The NPV results are simplified and are used for pit selection purposes only and are not meant to determine the project’s estimated value. NPV estimates were calculated using a 10% discount rate.

The graphs below summarize the results of the sensitivity analysis. The uniform shape of all the curves is indicative of a shallow deposit, where the pit size is insensitive to variations in mine costs, process costs, and metal price. The optimization mines everything above cut-off and there is no waste hurdle to expose new mineralization as costs decrease or metal prices increase. Snowden has selected a pit that maximises the NPV for the least amount of material processed. By selecting in this manner the total projected land disturbance is minimized and modeling risk reduced. The shallow nature of the deposit also allows for simple and easy mine expansions should economics and mineral resources be favourable at the estimated end of mine life.

Figure 17-1, Figure 17-2 and Figure 17-3 display mine costs, process costs, and metal price sensitivities for the entire range of pit shells.

Although pit size is insensitive, the value of the project can vary. For example, depending upon metal price, the resultant NPV for the selected pit can change by +/- 30% from the updated base case assumptions, by varying process costs by +/-10%. However, by varying mine costs by +/- 10%, the selected pit’s NPV has only a 2% variation.

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LOGO

Figure 17-1: Mine Cost Sensitivity Using Updated Feasibility Estimates

LOGO

Figure 17-2: Process Cost Sensitivity Using Updated Feasibility Estimates

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LOGO

Figure 17-3: Metal Price Sensitivity Using Updated Feasibility Estimates

Figure 17-4 summarizes the sensitivity analysis for mine costs, process costs, and metal price on the selected pit. The results indicate that metal price has the greatest impact on the value of the selected pit and increases or decreases to mining costs are negligible. This point was taken into consideration during the mine planning process, where grade control and waste management had a much higher priority than simply trying to reduce mining costs.

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Figure 17-4: Sensitivity Using Updated Feasibility Estimates On Selected Pit Shell

Mineral reserves were derived from a life of mine schedule that exploits the mineralization within the selected Whittle shell. There is a 7% discrepancy between the selected Whittle shell and the final mineral reserves as stated in Table 17-9. This is a result of some ore losses due to backfilling requirements and pit access.

17.2.3

Reserve Summary

Table 17-9 summarizes the mineral reserves for the Fenix Project.

Table 17-9: Fenix Mineral Reserve Estimates – 25 September 2006


Category	Tonnes	%Ni	%Fe	%SiO₂	%MgO
Proven	8,674,000	1.81	17.2	33.0	21.3
Probable	32,678,000	1.58	18.4	31.8	20.7
Proven + Probable	41,352,000	1.63	18.2	32.0	20.8

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The mineral reserves have been derived from only measured and indicated mineral resources. Proven reserves have only been based on measured mineral resources. Probable reserves have been based on indicated mineral resources and some measured mineral resources that enter the mine plan after Year 24. Further details relating to mineral reserves are provided in Section 18.1. As indicated in Sections 17.1.3.1 and 17.1.3.2 there are no known environmental, legal, title, taxation, socio-economic, marketing, political or other issues and no known mining, metallurgical or infrastructure issues which materially affect the estimation of the mineral reserves. Nevertheless the execution of the project is subject to various risks and uncertainties which may materially affect the economics of the project and therefore the mineral reserves. The project risks are identified in Section 18.15.

18.

Other Relevant Data and Information

18.1

Mining

18.1.1

Mining Methods and Dilution

The general mining sequence will begin with removal and storage of soil and organic cover for future rehabilitation of the mine areas. Waste cover and limonite will be stripped by excavator or track dozer and placed in either temporary external or in-pit storage areas. Limonite with potential for future processing will be stripped and stored separately from the waste. Mining will begin at the top of the hill and proceed to lower elevations. Ore will be extracted by conventional surface mining methods with no blasting expected. Excavators will work in either shovel or backhoe configuration depending on ore thickness and bench height and will load rigid-frame, off-road haul trucks. Wherever possible the backhoe will employ the double benching method to limit in-pit road building. The minimum selective mining unit is 2.5 m high.

Ore will be transported to the plant or stockpiles for future plant feed. Waste will be moved to temporary storage areas or dumped in mined out areas for long term storage. The waste dumps and stockpiles will be maintained and re-claimed by track dozers and wheel loaders.

The mineralized zone considered suitable for processing is a combination of saprolite and transitional (i.e. high iron) saprolite material. Dilution was applied along the top and bottom of this mineralized zone. To reduce dilution, mining methods will be varied to best suit the dip and thickness of the mineralized zones. Steeply dipping and narrow mineralized zones will be assigned a more selective mining method therefore production rates would be reduced, and costs will be factored up from the base rate. In flat-lying, thick mineralized zones, production rates increased and costs will be factored down. Along the top of the mineralized zone 0.5m of dilution and ore loss is assumed, and along the bottom of the mineralized zone 0.25m of dilution and ore loss is applied. Dilution for each area correlates directly to the vertical thickness of the mineralized zones. Area 215 has the highest dilution and on average has the thinnest mineralized zones. Overall dilution for all mining areas within the pit shell averages less than 4% by tonnes.

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18.1.2

Mine Production Schedule

Mining was scheduled within two pit shells to ensure high grades early in the mine life. A smaller pit shell, the Stage 1 pit, containing approximately 33 million tonnes of saprolite and transition material, provided the first 24 years’ production. Quarterly and annual mine production schedules were completed for Stage 1 using MineMAX software. The software enables automated scheduling while satisfying practical mining and all production and grade constraints. The larger and lower grade Stage 2 pit was scheduled in an Excel spreadsheet. To account for ore losses due to backfilling in Stage 1, 20% of the remaining ore tonnes from Stage 2 were sterilized, and mining costs were increased. Mill feed and mill head grade for the life of the project are summarized in Figure 18-1.

At steady state operation the mine will provide annually 1,464,000 tonnes of saprolite and transition ore to the plant. Total average annual mine production is approximately 3,100,000 tonnes at an average stripping ratio is 1.1. High grade and low grade stockpiles will be created in order to optimize nickel grade delivered to the plant. The high grade stockpile will contain ore above 1.5% nickel and reach a maximum of about 200,000 tonnes. The maximum low grade stockpile (material less than 1.5% nickel) is approximately 1,500,000 tonnes. Some waste material will be placed in temporary storage facilities outside the pit areas. However as internal space is created in the pits, waste will be stored within the previously mined out areas. Eventually the temporary storage facilities will be re-claimed and moved into the pits.

LOGO

Figure 18-1: Mill Feed Tonnes and Grade

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18.1.3

Mine Equipment Requirements and Costs

The key production mine equipment selected for the Fenix project comprises hydraulic excavators and rigid rear dump mining trucks (40 tonne capacity). The mine fleet will also include track dozers, front end loaders, core drills and portable lighting. Support equipment includes graders, water trucks, soil compactors, service and fuel trucks, a low bed transport truck and a crane. For ore haulage from area 251, 40 tonne highway trucks will also be required. The additional haulage will add $1.60/tonne to mine operating costs in this area.

The mine capital estimate is approximately $US15 million. The sustaining capital averages $US3 million per year for the life of the project. Start up capital costs were reduced by incorporating as much of the existing infrastructure into the project as possible. Labour costs were determined using Guatemalan pay standards except for the first three years where the senior mine personnel such as the manager and chief geologist are to be expatriates. Mine unit operating costs average $0.29/lb nickel.

18.2

Existing Facilities and Equipment

18.2.1

Overview

A key feature of the Fenix project is to maximise the use of the facilities previously owned and operated by Exmibal. There are some special considerations in that Exmibal’s product was nickel sulphide matte while the Fenix product will be ferro-nickel, and that not all of Exmibal’s physical assets will be suitable or available for modification and re-use.

A great deal of information from the Exmibal pilot plant, capital project and commercial operation is available to the Fenix project team and this information was considered in the development of the current project.

18.2.2

Plant Mothballing and Upkeep

In September of 1980, Exmibal’s mining and smelting operations were stopped. The emergency diesel generators and all mine, plant and maintenance mobile equipment were sold.

Considerable care was taken in mothballing the processing plant. All major rotary equipment was blocked off, and bearings and oil lines flushed and filled with diesel oil. Since then, maintenance personnel have routinely carried out necessary protective actions. Upkeep of the plant was audited by an Inco professional engineer every six months over the period.

18.2.3

DAVY Recommissioning Report (1993)

In 1993, Inco Exploration and Technical Services contracted Davy International to prepare a bankable-level report for the recommissioning of the Exmibal plant to the operational state of 1980. The report included projected capital and operating costs, and an implementation plan and schedule for the recommissioning. The site inspection team was comprised of:

•

Personnel from the architectural, civil, electrical, instrumentation and mechanical engineering groups of Davy International.

•

Representatives from the manufacturers of the ore dryer, reduction kiln, electric furnace, boiler and turbine generator. Babcock & Wilcox of Canada (B&W) and Turbinas y Generadores C.A. of Venezuela (Turgenca) were engaged to inspect the boiler and steam turbine equipment respectively.

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18.2.4

Fenix Refurbishment Requirements

The plant refurbishment requirements were assessed through a combination of vendor inspections and Hatch inspections. The inspections were primarily visual in nature, though labour was available to remove inspection covers, etc. Electricity and compressed air were not available to energize equipment drives or actuators.

18.2.4.1

Process Plant

18.2.4.1.1

Mechanical Inspection

Major equipment was inspected by vendor representatives and/or companies with related expertise. The vendors provided inspection reports including estimates of the cost and manhour requirements for refurbishment. The equipment that was inspected by vendors includes apron feeders, crushers, stackers, belt conveyors, rotary kiln, furnace electrode columns, the electrostatic precipitator and all bridge cranes.

Secondary equipment was inspected by Hatch engineers. Most of the equipment was found to be in good condition considering the long period of stand-down. The equipment exhibits little deterioration due to corrosion or other environmental conditions, except at locations where rain water collected such as in some insulated, un-drained enclosures, concave dimples, such as on tanks, piping, fans and ductwork, located outside.

18.2.4.1.2

Structural Inspection

The Hatch Structural Assets group completed a site audit of the existing Fenix facility. The scope of the audit included inspection of the civil, structural and architectural features of the process plant, power plant, utilities, in-plant roads and non-process buildings.

Based on the visual structural audit and condition assessments the facility was found to be structurally sound. In general the foundations and concrete works were found to be in good condition. Corrosion of steel structures is a problem in only a few areas and can be repaired by blasting and repainting or by paint touch-up depending on the condition of the steel.

Architectural inspections revealed that most of the non-process buildings have mainly cosmetic problems. Most of the architectural refurbishing effort will be concentrated on cleaning, painting, replacing ceilings/floors where required, installation of new and/or repair of existing fixtures, doors and windows.

During the Detailed Engineering phase of this project, core samples of existing concrete and structural steel specimens will be obtained and tested to confirm material properties.

18.2.4.1.3

Controls, Automation and Electrical (CA&E)

The following assumptions were made for the power distribution and the CA&E estimates:

•

A majority of the main supply feeder cables will be replaced with new 5 kV cables. It is assumed that the existing cable trays and underground cable ducts for the supply feeders will be reused.

•

New feeders from the existing overhauled switchboards will be provided for the supply of the new MCC’s.

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•

It is assumed that all existing MCC’s will be replaced with new ones. As a consequence, it is assumed that all existing LV power cables will be replaced with new ones.

•

It is assumed that 10% of the total cost of the existing grounding will be added for refurbishment/replacement.

•

It is assumed that all existing cable trays will be reused as is.

•

It is assumed that 50% of the existing lighting and LV distribution equipment will be replaced with new equipment.

•

It is assumed that all the existing local control stations (including cable) and electrical junction boxes will be replaced with new components.

•

It is assumed that all the existing instrumentation (including cable and tubing) and the control systems will be replaced with new components.

18.2.4.2

Power Plant

18.2.4.2.1

Inspections

The assessment of the boiler and steam turbine generator were sub-contracted to specialist companies: Babcock & Wilcox Company (B&W) and Siemens (Westinghouse) Power Generation Services, respectively.

Generally the inspections were carried out without major disassembly of the equipment. Where possible, access covers were removed, rotating equipment was turned over by hand and valves and dampers were operated to determine the equipment condition. Electric power and compressed air were not available during the inspection. The boiler inspection included nondestructive testing; boroscopic inspections were performed on some turbine components.

18.2.4.2.2

Boiler

The Phase 2 project does not include the refurbishment of the existing boiler. B&W’s recommendations, from their Condition Assessment Report, are noted below for reference.

•

Replace the following items for the Forced Draft Fan: electric motor drive, all variable inlet vanes, all variable inlet vane bearings, FD fan shaft bearings, vane linkage actuator drive.

•

Replace drum and superheater safety valves, including the EBV valve.

•

Replace all boiler sootblowers, the sootblower control system, the motor starters, and the pressure reducing station and the drains for the sootblower piping.

•

Replace the steam coil air heater and its steam control valve and condensate drain valves/traps.

•

Replace the flue gas inlet and fresh air outlet duct section of the air heater, the windbox roof and the penthouse casing.

•

Replace expansion joints.

•

Replace and upgrade the Burner Management System.

•

Replace igniters, flame scanners, cooling air blowers, oil guns, and burner/igniter valve trains, oil gun and air register actuators.

•

Refurbish all large valves (>2”).

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18.2.4.2.3

Steam Turbine Generator

The refurbishment planned for the steam turbine and generator are also consistent with the recommendations made by Siemens and can be summarized as follows:

•

Dismantle the steam turbine, clean, inspect, replace parts as necessary and re-assemble.

•

Dismantle the generator, clean, inspect, test, replace parts as necessary, and re-assemble.

•

Test the generator insulation after drying it out.

•

Inspect and clean all auxiliary systems and equipment. Replace filters, gaskets and seals, as required. Refill lube oil system.

•

Commission turbine systems - check, verify and calibrate components and systems.

18.2.4.2.4

Ancillary Equipment and Systems

All equipment and piping requires a thorough cleaning and most equipment will need to be disassembled for an internal inspection and reconditioning /refurbishment. All valves have to be inspected and reconditioned , some will have to be replaced. Some equipment has corroded substantially due to weather exposure and some parts may have to be replaced. Pressure parts will need to be subjected to a NDT inspection or pressure tests to verify the integrity of the equipment.

18.2.4.2.5

Instrumentation and Control

The existing control system in the control room will be replaced with a Distributed Control System (DCS). Where possible primary elements and drives will be retained or modified as necessary for compatibility with the new DCS. Functional equipment, e.g., pressure and temperature transmitters, gauges and indicators will be retained. Incompatible or deteriorated components will be replaced.

18.3

Process Plant Description

18.3.1

Geotechnical Assessment

Trow International Ltd. conducted a geotechnical investigation of the plant site. The fieldwork was performed between July 24, 2005 and August 14, 2005 and consisted of drilling 25 boreholes to depths varying from 6 to 30 meters.

In general, many boreholes encountered the following:

•

Layer of fill materials with thickness varied from 1.2 m to 8 m. The fill consists of materials with variable composition of mine waste, crushed rock, slag, sand, gravel and some topsoil.

•

Underlying the fill, a layer of brown to grey silty clay to sandy silt with boulders to depths of about 7.3 m to 16 m. The thickness of this layer varied from 1 m to 10.1 m.

•

Saprolitic silty clay with boulders with thickness from about 2.5 m to 9 m.

•

All boreholes encountered highly fractured weathered rock at various depths (except for a few boreholes located farthest away from the site), borehole drilling terminated in this rock layer.

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Note that the borehole information for the proposed areas of the power plant and coal/petroleum coke receiving, storage, reclaim was extrapolated due to the limited number of boreholes in the vicinity of these structures.

Groundwater conditions in the open boreholes were observed throughout and immediately after the drilling operations. Standpipe piezometers were installed in selected boreholes to permit long term monitoring of groundwater levels at the site. The groundwater at the site will fluctuate seasonally and can be expected to be somewhat higher during the wet season and in response to major weather events. For design purposes the average groundwater levels were assumed at about 2 m below existing grade. If shallow rock is encountered at depth of less than 2 m, the groundwater level could be assumed at the top of rock surface. Under severe weather conditions (heavy rain-storm), the groundwater level could be located near to the existing ground surface.

Based on the results of the site investigation and laboratory testing program, Trow provided design recommendations for the following:

•

Site Grading

•

Fill Suitability

•

Slope Stability

•

Foundation Design

•

Temporary Shoring

18.3.2

Process Operations

A description of the process plant operations, from receipt of ore from the mine to product packaging, is provided is the following subsections. The process block flow diagram is presented in Figure 18-2.

18.3.2.1

Ore Preparation

Ore is trucked from the mine and fed into a 180 tonne dump pocket through a horizontal, adjustable bar, stationary grizzly set at 750 mm. Oversize rocks are rejected while the undersize is fed from the dump pocket by an apron feeder to a double roll crusher. The ore is crushed to -100 mm and transferred by belt conveyor to the wet ore stockpile tripper conveyor. Ore is conveyed to the stacker which is mounted on rails. Ore is transferred to the boom conveyor of the stacker and moved to the discharge point. The stacker and conveying equipment are rated for up to 545 wet t/h.

Each of the two stockpiles is 103 m long, 64 m wide and 24 m high, contains 92,000 wet tonnes of ore and can provide 12 days of dryer feed. The stacker moves back and forth along the long axis of the stockpile, progressively building a blended pile. The stacker builds up one stockpile while the other is being reclaimed as feed to the dryer.

18.3.2.2

Drying and Dry Ore Storage

Blended ore is reclaimed from the wet ore stockpiles and deposited into the dryer feed dump pocket by front-end loader. The ore is fed from the dump pocket at a controlled rate by an apron feeder equipped with a variable speed drive to the dryer feed belt conveyor. Ore enters the dryer at a nominal moisture content of 34%, although the dryer is designed to partially dry ore with moisture up to 38%. Pugged dust, at 20% moisture, that has been collected from various areas of the plant is also added to the dryer feed conveyor.

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LOGO

Figure 18-2: Process Block Flow Diagram

The co-currently fired rotary dryer is 5.5 m in diameter and 40 m long, and has an unlined carbon steel shell equipped with internal lifters. The barrel is supported on two riding rings on the shell, which rest on roller assemblies fixed to the top of concrete piers. A 633 kW drive rotates the barrel at 3 rpm via a toothed gear ring on the shell, located between the two riding rings. Pulverized bituminous coal is combusted in a refractory lined combustion chamber attached to the feed-end housing. A primary fan is used for coal combustion and a secondary fan provides tempering air to the combustion chamber so that the hot combustion gases enter the dryer at 1000ºC.

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The ore is dried to 20% moisture and passed through a trommel screen located at the end of the dryer shell. Ore that is less than 38 mm discharges directly to the dryer discharge conveyor. Ore that is -152 mm +38 mm is diverted to an impact crusher and crushed to –38mm before being discharged onto the dryer discharge conveyor. Oversized ore is rejected.

Dryer off-gas leaves at 150ºC and is ducted to an electrostatic precipitator (ESP) for particulate collection. The gases are cleaned to achieve a particulate emission rate below 20 mg/dNm³, in accordance with World Bank standards. The clean gases are vented to atmosphere through a 40 m high stack by an ID fan. Dust collected in the ESP is pneumatically conveyed to the dust recycle area for return to the process. SO₂ and NOx levels will be below the established environmental emission limits of 2,000 mg / dry Nm³ and 750 mg / dry Nm³, respectively.

The partially-dried, crushed ore (-38 mm) is conveyed to the covered dry ore storage building. A travelling, slewing, stacker operating within the building stacks the ore into four stockpiles, which are built in horizontal (windrow) layers to promote blending. The dry ore storage capacity of 36,000 tonnes provides close to 5 days of inventory for the reduction kilns. A front end loader reclaims the ore and transfers it into the 27-tonne dry ore reclaim hopper, from which it is fed by a variable speed belt feeder to the kilns’ feed bin feed conveyor. This feeds the reduction kilns’ feed bins, each having 250 tonnes storage capacity.

18.3.2.3

Calcining and Reduction

There are two reduction kilns operating in parallel. Kiln No. 1 is 5.5 m in diameter and 100 m long, while Kiln No. 2 is 6.0 m in diameter and 115 m long. Both kilns are inclined at 3.1 percent.

Kiln No. 1 is supported on three riding rings on the shell each of which rests on a pair of roller assemblies fixed to concrete piers. The kiln is driven by two 340 kW variable speed drives which drive pinion gears transmitting power to a girth gear located near the centre riding ring. The normal speed of rotation of the kiln barrel is 0.9 rpm. Kiln No. 2 is supported via four riding rings on the shell, which rest on roller assemblies fixed to concrete piers. Two 750 kW variable speed drive units rotate the barrel at 1.0 rpm via the toothed gear ring, which is located close to the kiln feed end. Both kilns are equipped with an emergency drive which turns the barrel at 0.1 rpm.

Each kiln has the following features that help to maximize productivity and efficiency:

•

There are four dams along the length. The dams’ height above the refractory lining for Kiln No. 1 and Kiln No. 2 will be 1,000 mm and 1020 mm respectively. Two dams will be located in the cooler middle zone, while the other two dams will be towards the hotter discharge zone.

•

Lifters will tumble the solids and enhance heat transfer from the gas and refractory lining, and prevent segregation of different size fractions in the bed. Spiral lifters along the first two metres help to propel feed into the kiln; the remainder are straight bar-shaped lifters.

•

A coal scoop will feed coal into the kiln 2/3 along the barrel from the feed end. Coal is metered from a bin into a trough that lies underneath the kiln. The scoop, mounted on and through the shell, opens as it rises through the coal-filled trough, closes as it rotates further, and then deposits coal into the kiln as it continues rotating to the uppermost position. Coal is thereby added to the kiln where the temperature is hot enough to ensure (in conjunction with tertiary air addition) that the volatile component of the coal is fully combusted, providing maximum calorific value.

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•

Four shell-mounted, on-board air blowers are distributed along the length of the kiln to allow a substantial proportion of the total air requirement, termed tertiary air, to be introduced away from the burner end of the kiln. This allows greater control, and hence optimization, of the energy profile along the kiln and (in conjunction with the coal scoop) ensures maximum utilisation of the energy in the volatile component of the coal.

The partially reduced, calcined ore reaches a maximum temperature a short distance from the kiln discharge, then falls slightly, so that it leaves the kiln at a target temperature of 900ºC. The calcine discharges from Kiln No. 1 to a 45-tonne capacity, refractory-lined surge bin at a nominal rate of 79 t/h. Calcine from Kiln No. 2 discharges to a separate 55-tonne capacity, refractory-lined surge bin at a nominal rate of 98 t/h. The surge bins are mounted on load-cells for level indication. Any oversize material caused by partial sintering of the calcine is diverted to grade by means of an air cooled grizzly in each surge bin. Calcine is batch-discharged from each surge bin to 19-tonne transfer containers through a slide gate.

The burner end of the kiln is maintained at a slight negative pressure by means of the exhaust fan damper. Fuel and total air input are controlled to maintain the target calcine temperature and to regulate the off-gas composition to ensure neither oxygen nor carbon monoxide levels are too high, avoiding inefficient air addition and potentially explosive gas mixtures, respectively.

The kiln off-gas leaves the feed-end housing at approximately 340ºC and passes through an ESP (one for each kiln) where the dust is removed and pneumatically conveyed to the dust recycle area. The cleaned gas is exhausted via induced draft fans through two dedicated stacks designed to achieve the World Bank particulates emission rate of below 20 mg/dNm³. The SO₂ and NOx levels will be below the established environmental emission limits.

18.3.2.4

Calcine Transfer and Furnace Feed System

Each kiln discharges calcine into a calcine transfer system consisting of two containers, one transfer car and one bridge crane. The hot calcine is moved from the kiln surge bins to the electric furnace feed bins located above the furnace. The furnace feed system delivers hot calcine from the feed bins to the electric furnace in a controlled manner which ensures that a consistent layer of calcine is maintained on the molten slag layer. Dust emissions during calcine transfer are captured and sent to the secondary ventilation baghouse. The principle components of the calcine transfer and furnace feed systems are:

•

Refractory-lined 19 tonne capacity transfer containers (3 per kiln, 2 operating, 1 standby).

•

Calcine transfer car capable of holding two fully loaded containers (1 per kiln).

•

One 36 tonne capacity bridge crane per kiln.

•

Nine 31 tonne capacity furnace feed bins mounted on load-cells. The nine feed bins are situated above the electric furnace in a circular arrangement around the electrode area. Each feed bin is suspension mounted via tension style load-cells located below the charging floor.

•

The furnace has twenty-four refractory lined feed pipes, each with an actuated discharge valve, to provide a controlled feedrate to the furnace. Three pipes feed calcine into the central zone between the three electrodes, nine pipes feed calcine to the zone immediately surrounding the electrodes, and twelve pipes distribute calcine near to the sidewalls.

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18.3.2.5

Electric Furnace Smelting

18.3.2.5.1

Overview

The 18 m diameter, three-electrode AC electric arc furnace, supplied by 3 x 45 MVA transformers, smelts 174 t/h of calcine at 90 MW. The furnace operates in high-voltage, low-current, shielded-arc mode. Hot calcine fed through the roof of the furnace is smelted and separates into slag and metal phases. Residual carbon in the calcine completes the reduction of nickel and iron to achieve the targeted ferro-nickel grade of 35% nickel. Carbon monoxide evolved from the bath is combusted in the furnace freeboard with infiltration air.

Crude ferro-nickel is tapped at 1,525°C between five and six times a day from one of two metal tapholes and flows down a refractory-lined launder to a 47-tonne capacity refractory lined ladle. After tapping is complete, the metal ladle is withdrawn from the tapping area by rail car and transferred by crane to the refining station transfer car. Slag is tapped semi-continuously from one of two slag tapholes down a water-cooled launder into the granulation system.

The furnace off-gas is sent to a spray cooler and an ESP. Dust collected in the ESP is pneumatically conveyed to the dust recycle area. The gas is drawn through the system by an induced draft fan and discharged to atmosphere through a stack. Dust from the spray cooler is collected in a tote box and sent back to the dry ore reclaim hopper.

The design parameters of the furnace are consistent with those in the laterite smelting industry, as summarised below:


Parameter	Fenix	Other Laterite Furnaces
Slag Silica:Magnesia Ratio	1.6	1.3 - 2.9
Bath Power, kW/m ²	130	50 - 215

The furnace power is in accordance with the established trend of the industry, as shown below in Figure 18-3 and Figure 18-4.

LOGO

Figure 18-3: Furnace Power Levels vs. Time

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LOGO

Figure 18-4: Nickel Laterite Furnace Sizing Parameters

18.3.2.5.2

Furnace Description

Furnace power will be supplied through water-cooled low voltage bus from three single phase transformers arranged in-line. The furnace will be equipped with Soderberg electrode columns and a state-of-the-art control, monitoring and diagnostic system. The furnace design is in accordance with Hatch’s philosophy of building robust furnaces and incorporates the following features:

•

The refractory layout ensures adequate brick bonding both horizontally and vertically with features including bricks keyed to the cooling elements, tongue and groove bricks and hold back clips to ensure a well-sealed, robust lining. The furnace bottom, which is forced-air cooled, is dish shaped, constructed of wedge-shaped, tongue and groove bricks in the working lining to minimise the risk of floating the hearth bricks. The sidewall and hearth bricks are magnesia for good chemical resistance and thermal conductivity. Some alumina-chrome material is used to help reduce the risk of hydration. The furnace roof refractory consists of hung alumina brick, with pre-cast shapes of similar composition around the openings. The roof’s steel beams are forced-air cooled.

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•

The furnace shell is equipped with the Hatch Wall Hold-Down System that maintains a vertical compressive force on the refractory bricks in the sidewall. The system consists of 18 spring loaded devices mounted to the shell plate that transfer loads through a circular ring beam onto the top of the refractory wall. This compressive force is key to the thermal design of the wall as it ensures that good contact, and therefore good thermal conductivity, is maintained between the brick layers and between the bricks and cooling elements. The vertical pressure also minimizes any ingress of metal or slag into the horizontal brick joints of the sidewall and thus prevents upward growth of the walls due to thermal ratcheting. Thermal ratcheting occurs on the refractory hot face when metal and slag penetrate the brick joints during successive furnace cool-down/heat-up cycles.

•

To minimize refractory wear in the slag zone, three staggered rows of water-cooled copper plates are installed in the bricks in this zone. These plate coolers are located such that the lowest row is above the maximum metal level. The water passages in the plate coolers are located outside of the furnace, at the plate cooler cold-face.

•

The slag/metal interface zone below the maximum metal level has historically been the highest wear area in nickel laterite smelting furnaces. The 342 mm high copper cooling elements installed in this zone have a waffled hot-face (refractory-filled) and cast-in pipe for cooling water that extends almost all the way to the block hot-face. The cooling water systems for the plate and waffle coolers are designed for 200% of the average expected heat flux. The cooling elements themselves are designed for 4 times the average expected heat flux.

•

At the skew level, forced-air cooled copper fins are installed on the lower furnace sidewall shell. The fin system works in conjunction with the copper coolers to ensure that a frozen slag crucible is formed, and also arrests penetration of metal into cracks between the bricks.

18.3.2.6

Refining, Shotting and Product Drying

The refining area processes approximately 69,500 t/y of crude ferro-nickel from the electric furnace, which corresponds to approximately five 47 t heats per furnace per day. The refining line consists of a chemical heating station, a de-slagging station and an 11 MVA ladle furnace arranged along one common rail track. Additional stations are provided for non-process tasks including ladle relining, drying and preheating, and for the old linings to be removed. There are a total of six refractory-lined refining ladles. Refining reagents for each line are kept in eight storage bins and delivered to the heating station and ladle furnace by means of a rail-mounted travelling weigh hopper.

18.3.2.7

Secondary Off-Gas System

The purpose of this system is to improve in-plant hygiene levels and reduce dust losses from the plant, by collecting calcine transfer dust releases, fumes generated by furnace tapping and refinery process gases. These gases are collected in hoods and a network of ducts and brought to a single duct running the length of Kiln No. 1 to a venturi scrubber. The gases are cooled and cleaned prior to being vented to atmosphere. Particulate emissions of 50 mg/dNm³ are expected. Scrubber slurry will be thickened and the underflow solids pumped to the dust recycle area.

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18.3.2.8

Hygiene Systems

Five separate fugitive dust control systems are proposed for the process and power plants to improve in-plant hygiene by collecting fugitive dust emissions. Typical control points include conveyor transfers, crushing operations and handling of materials by mobile equipment. Systems are grouped based on exhaust point proximity and type of dust. The coal dust collection system is isolated from other dust systems for safety reasons. Separate systems are required for control points physically removed from each other to limit the length of duct runs. Water sprays will be used to suppress dusting from the wet ore, dry ore, coal and coke stockpiles.

18.4

Power Supply

18.4.1

Summary

The October 2006 Feasibility Study contemplated the construction of an on-site dedicated power plant to meet the project power requirements. However, the lead times for key components of the power plant have increased to such an extent that timely completion is not feasible. In order to maintain the overall project schedule, CGN is negotiating an interim supply of power through a Power Purchase Agreement (PPA). The PPA counterparty is expected to not only provide the project power requirements, but also develop the requisite transmission infrastructure to deliver contracted power to site over the national grid through a related Transmission Tolling Agreement (TTA). Securing interim project power through a PPA/TTA not only removes constraints on the ramp-up of the process plant, but also reduces the initial project capital requirements. However, to secure long-term low power costs beyond the expected five year term of the PPA/TTA the development of an on-site power plant is planned. The power plant would be constructed after the process plant is in service.

18.4.2

Power Purchase Agreement (PPA)

18.4.2.1

Guatemalan Electrical System

The Guatemalan electrical power sector is characterized by the following:

It has an established Wholesale Electricity Market (WEM) that has operated with private participation since 1992.

Thermal and renewable (including small hydro) generators are largely privately owned whereas large hydro is publicly owned.

Transmission networks are largely publicly owned whereas distribution networks are privately owned.

It has recently had a surplus of energy production as evidenced by increasing power exports to neighbouring countries – principally Honduras and El Salvador.

Public ownership of hydro generation and transmission infrastructure is through Institute Nacional de Electrificacion (INDE). Private ownership of generation and distribution companies is more broad.

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The function of the WEM is to produce a market driven optimization of the electric system through contract and spot transactions between buyers and sellers of capacity and energy. In Guatemala, all final consumers of electricity – whether it be distributors or large industrial users – must be backed by commitments from generating plants. Capacity contracts are the principal means of fulfilling these obligations.

Table 18-1 provides an overview of generating capacity on the Guatemalan Electrical Grid. The installed capacity represents the sum of the nameplate capacities of the respective generating plants whereas the firm capacity is the sum of the capacities that can be committed to contracts and/or the spot market as determined by the Market Administrator (AMM).

The overall generating capacity is characterized by a large contribution from reciprocating engine plants which are primarily Heavy Fuel Oil (HFO) based. The combustion turbine units are also primarily oil distillate based given the lack of natural gas infrastructure in Guatemala. There is only one coal-fired power plant of any significance (the 139 MW San Jose plant). The system is also characterized by a significant capacity contribution from bagasse generators. Bagasse is residue from sugar cane processing that is available during the harvest season only. However, as this coincides with the dry season this offsets some of the reduced hydroelectric generating capacity. Otherwise, despite its large potential there is limited geothermal generating capacity in Guatemala.

Table 18-1: Overview of Generating Capacity on the Guatemalan Electrical Grid


	Units	Total Capacity [MW]		Comments
	Units	Installed	Firm	Comments
Hydroelectric	48	742.5	653.5	About 1/3 of installed/firm capacity.
Reciprocating Engine (RE)	71	649.1	624.6	Primarily HFO based.
Bagasse (B)	Variable	306.5	276.1	Sugar cane harvest season only.
Steam Turbine (ST)	2	140.5	132.9	Primarily San Jose.
Combustion Turbine (CT)	4	172.9	116.8	Primarily distillate based.
Geothermal (G)	8	29.0	20.7	Primarily Zunil.
Total Thermal	85 + B	1298.0	1171.1	About 2/3 of installed/firm capacity.
Total System	133 + B	2040.5	1824.6

18.4.2.2

Process Plant Load Requirements

The power supply must be able to meet the nominal and peak power demand of the process plant. Table 18-2 summarizes the process plant electrical power demand requirements. The capacity requirement is defined by the peak power demand over a 15 minute window. This is provisionally estimated to be 135 MW pending a more detailed load study based on certified vendor data.

Table 18-2: Process Plant Electrical Power Demand


Load Description	Nominal Demand [MW]	Peak Demand [MW]
Smelting Furnace	90	96
Refining Furnace	9	10
Process Plant Motive Power	21	25
Lighting and Building Services	3	3
Distribution Losses	1	1
Total	124 MW	135 MW

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18.4.2.3

Specific Generating Plants

The onus on all final consumers to secure commitments from generating plants to meet their capacity requirements is the basis of the contract market. The contract market is made up of contracts – generally Power Purchase Agreements (PPA’s) – between users and generators for the provision of capacity, energy or both. In Guatemala, the contract market predominates over the spot market. The contracts market is executed by the contracting parties. The only requirements imposed by the Electricity Law are:

•

Contracts be made public.

•

Consumers are required to have capacity contracts covering their firm demand requirements.

•

Supplier’s contract commitments cannot exceed the firm generating capacity under their control.

•

Contracts cannot contain take-or-pay clauses requiring the physical dispatch of plants.

The latter insures that generating units are dispatched according to their marginal costs rather than their contractual commitments. The Market Administrator (AMM) is responsible for establishing generators firm capacity and marginal costs. Table 18-3 depicts the largest generating stations on the Guatemalan electrical grid. The installed capacity represents the nameplate capacity whereas the firm capacity is the maximum capacity that can be committed to contracts and/or the spot market as determined by the Market Administrator (AMM).

The largest generating plants in the country – namely the 300 MW Chixoy Hydroelectric Generating Station (Chixoy GS) and the 160 MW Arizona Reciprocating Engine Plant (Planta Arizona) – were visited by representatives from Hatch in the course of a grid connection due diligence review.

Table 18-3: Largest Generating Stations on the Guatemalan Electrical Grid


	Units	Total Capacity [MW]		Date of Installation	Location (Department)
Generating Plant	Units	Installed	Firm	Date of Installation	Location (Department)
Hydroelectric
Chixoy	5	300	272	1983	Alta Verapaz
Aguacapa	3	90	80	1982	Santa Rosa
Jurun Marinala	3	75	60	1970	Escuintla
Renace	3	63	60	2004	Alta Verapaz
El Canada	2	48.1	47.4	2003	Quezaltenango
Las Vacas	3	45.7	39	2002	Guatemala
Secacao	1	16.5	16.3	1998	Alta Verapaz
Los Esclavos	2	15.0	14.0	1966	Santa Rosa
Thermalelectric
Arizona (R)	10	160.0	155.7	2003	Escuintla
San Jose (T)	1	139.0	132.4	2000	Escuintla
Poliwatt (R)	7	129.4	125.5	2000	Escuintla
Puerto Quetzal Power (R)	20	118.0	114.6	1993	Escuintla
Tampa (G)	2	80.0	78.1	1995	Escuintla
Las Palmas (R)	5	66.8	66.4	1998	Escuintla
Stewart & Stevenson (G)	1	51.0	23.7	1995	Escuintla
Genor (R)	4	46.2	41.4	1998	Puerto Barrios
Sidegua (R)	10	44.0	38.0	1995	Escuintla
Magdalena (B)		40.0	40.0	1994	Escuintla

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18.4.2.4

PPA Description

The Fenix projects expects to secure its electrical capacity and energy requirements through a Power Purchase Agreement (PPA) with an Independent Power Producer (IPP). The IPP owns and operates a portfolio of generating facilities located south of Guatemala City. These facilities include the 160 MW Arizona and the 85 MW Las Palmas plant.

With firm capacity of almost 156 MW, Arizona will be instrumental in fulfilling the obligation of the PPA. Arizona was developed by an IPP as a dual fuel (HFO/Orimulsion) plant. It first went into operation in 2003 using HFO and converted to Orimulsion in 2004. It operated successfully on Orimulsion until all supply was diverted to China in 2005 at which point it was converted back to HFO. The plant presently operates on HFO.

The PPA is expected to take the form of a capacity contract with associated energy – that is Fenix will purchase a pre-determined block of firm capacity and associated energy. The contracted capacity is expected to cover a period of five (5) years and three (3) months from July 1, 2009 to September 30, 2014 as described in Table 18-4. Note that contracted capacity includes provision for ramp-up of the processing plant. The contracted capacity stipulates a range with a final capacity values to be ratified by June 30, 2008.

CGN intends to negotiate an option of renewing the PPA for an additional five (5) year period commencing July 1, 2014. However, as the energy component of the PPA will likely be indexed to the price of HFO, the plan is to redevelop the on-site power plant as originally contemplated.

Table 18-4: Contracted Capacity Covering the Term of the PPA


	Start	Finish	Contracted Capacity
Period	Start	Finish	Min.	Max.
3 months	1/July/2009	30/Sept/2009	25 MW	25MW
3 months	1/Oct/2009	31/Dec/2009	42.5 MW	60 MW
3 months	1/Jan/2010	31/Mar/2010	70.5	85
3 months	1/Apr/2010	30/Jun/2010	85.5	103
3 months	1/Jul/2010	30/Sep/2010	94.5	112
3 months	1/Oct/2010	31/Dec/2010	98.5	116
3 months	1/Jan/2011	31/Mar/2011	101.5	120
3 months	1/Apr/2011	30/Jun/2011	105.5	125
3 months	1/Jul/2011	30/Sep/2011	112.5	131
36 months	1/Oct/2011	30/Sep/2014	117.5	135

18.4.3

Transmission Toll Agreement (TTA)

18.4.3.1

Guatemalan Interconnected System

The Guatemalan Interconnected System is a network of 69 kV, 138 kV and 230 kV transmission lines and associated substations and switching stations that interconnect generators with distributors and large users. There are also interconnections to adjacent countries including El Salvador and Honduras. An interconnection to Mexico will also be implemented shortly.

The Electricity Law of 1996 among other things legislated open access to the transmission system (note that the legalization of private participation in generation preceded in 1992). Though the transmission network remains under the ownership of INDE, it’s operation is regulated by the National Electrical Energy Commission (CNEE) with day-to-day control and administration executed by the Market Administrator (AMM).

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18.4.3.2

Project Fenix Requirements

The project power requirements demand connection at the 230 kV transmission level. However, the nearest transmission substation in El Estor is supplied by a 69 kV line from Puerto Barrios. The nearest 230 kV connection point is the Tactic substation some 160 km to the west of the plant site. This had been previously identified as a candidate grid connection point during the feasibility study (PRH319112.0015 Grid Connection Preliminary Assessment).

Tactic is an important substation in the integrated system as it handles all of the power flow from the 300 MW Chixoy Hydroelectric Generating System (the largest generator in the system) as well as supports the 69 kV networks supplying the centre of the country (Coban and environs).

The Tactic Substation was constructed in 2002 by Union Fenosa to permit the interconnection of a hydroelectric development. The substation is owned and operated by INDE. The substation 230 kV switchgear is encapsulated indoor Gas Insulated Switchgear (GIS) supplied by Siemens. This is state-of-the-art switchgear not normally seen in remote utility substation. As per typical utility reliability requirements, the 230 kV line-up features a double bus with a tie-breaker.

18.4.3.3

Tactic-El Estor 230 kV Transmission Line

To support the execution of a Power Purchase Agreement (PPA) between an IPP and Project Fenix, the IPP through an affiliate will develop a 230 kV transmission line between the Tactic and El Estor substations. The IPP will subsequently recover the costs of the transmission line development through a Transmission Toll Agreement (TTA) with Project Fenix. The IPP has developed similar 230 kV transmission line infrastructure to connect its 160 MW Arizona Reciprocating Engine Plant to the national grid.

To accelerate route selection and right-of-way acquisition, the transmission line route will follow national highways. This is a similar approach to that used for the Arizona 230 kV connection. Specifically, the Tactic-El Estor line follow national highways CA-14 and RN-7E passing through the municipalities of Tamaru, La Tinta and Panzós. The length of the line will be approximately 160 km.

As the Tactic-El Estor transmission line is a necessary prerequisite to the PPA, the IPP has initiated permitting, basic engineering and contract negotiations in order to achieve an in service date of July 2009. This is three (3) months before first delivery of contracted capacity and energy per the PPA. The IPP has also submitted and received the environmental approval required for construction of the Tactic-EL Estor line.

18.4.3.4 Interconnection Studies

Due to the relative size of the Project Fenix capacity requirements, there was a perception that the project might adversely impact the stable and reliable operation of the national interconnected grid system. Hence, interconnection studies were commissioned to assess the impact of Project Fenix on the national grid. The studies demonstrate that the project can be served by the existing and proposed infrastructure without adversely impacting the interconnected grid system. The studies have been submitted for review and approval to the grid regulatory and operating authorities, namely CNEE and AMM respectively.

Note that interconnecting of the Tactic and El Estor substations will create a new electrical loop in the national grid and therefore improve the reliability and stability of the interconnected system.

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18.4.3.5 TTA Description

The PPA counter party will construct, own and maintain the Tactic-El Estor 230 kV transmission line. The PPA counter party will apply to include the line in the primary transmission system, which may lead to cost recoveries from the AMM. However, there is no certainty as to the likelihood or amount of any recoveries from the AMM. Consequently, The PPA counter party will recover its construction and operating costs through a Transmission Toll Agreement (TTA) with CGN and CGN will benefit from any recoveries from the AMM through a reduction in transmission tolls payable to the PPA counter party. An initial term of twenty (20) years commencing at the same time as the PPA (July 1, 2009) is expected to take effect. During this period, the PPA counter party will invoice CGN in equal monthly amounts such that over the term of the agreement the cost (with interest) of the transmission line is recovered.

Note that CGN will also be invoiced for transmission tolls, transmission losses, spinning reserve and other ancillary services associated with wheeling the project power requirements through the interconnected system. These costs are regulated by the CNEE and invoiced by the AMM.

18.4.4

On-Site Power Generation

For the provision of power supply after the initial five-year term of the PPA, a dedicated power plant has been selected as the basis for the Fenix facility power requirements. The facility peak power requirements of 135 MW include a 90 MW (nominal) smelting furnace load. Smelting furnace load is characterized as a difficult load and several provisions are incorporated in the furnace and power plant to minimize it’s impact on the power supply system.

The existing thermal power plant consists of a single boiler-turbine island rated at 60 MW. It was commissioned in the late 1970’s with ratings commensurate with operations at the time. The existing thermal power plant features a Heavy Fuel Oil (HFO) fired boiler, single-casing turbine and once-through cooling system drawing water from Lake Izabal. The reliance on HFO as a primary fuel source and the rejection of heat into Lake Izabal are undesirable features that were not adopted in the new power plant configuration. However, an assessment of the existing turbine generator indicated that it is fully operable and can be put into operation with modest refurbishment.

Among a host of options, the optimal boiler-turbine configuration was found to include a full-rated (150 MW) boiler supplying the existing 60 MW turbine and a new 90 MW unit. This configuration represents the lowest total annual cost, with operational flexibility to meet the range of load requirements and acceptable reliability.

A key feature of the full-rated boiler is that it will be of Circulating Fluid Bed (CFB) design rather than of conventional Pulverized Coal (PC) design. CFB boilers are well proven in this size range and offer a number of important features:

•

Variability of fuel type including high sulphur fuels.

•

Low emission of nitrous oxides.

•

Better turn down capability.

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The variability of fuel type, including those with high sulphur content, is a key strategic asset for a merchant plant of this type where shorter-term fuel supply contracts (relative to their utility counterparts) are the norm. The power plant is engineered to burn petroleum coke (petcoke), a by-product of oil refining whose supply is anticipated to outstrip demand in the coming years (as per Pace Global’s “Coal and Petcoke Strategy Supply Plan Report,” dated July 11, 2006). Due to its high sulphur content, petroleum coke cannot be consumed in as large amounts in conventional utility PC-based thermal plants as it can in CFB-based plants, unless expensive flue gas desulphurization is employed. The CFB boiler does not require flue gas desulphurization because sulphur can be captured in the fluidized bed by adding limestone to the boiler. The power plant will also accommodate other fuel sources including both low and high sulphur coals. The only serious consequence of sulphur in the feed fuel is increased limestone consumption and associated fly and bottom ash handling and disposal. It should be noted that natural gas is not a viable fuel source as neither pipeline or Liquified Natural Gas (LNG) facilities exist in Guatemala.

Nitrous oxide emissions from CFB-based boilers are low due to the low temperature of the fluidized bed. In fact, they meet World Bank standards without any special mitigation measures.

Finally, the better turn down capability allows the CFB boiler to accommodate a wide range of plant operating conditions without resorting to a secondary fuel such as heavy fuel oil.

In addition to the full-rated CFB boiler, the upgraded power plant will include the following features:

•

Single-casing turbine design for maximum robustness in supplying industrial and smelting furnace load.

•

Steam-bypass valving on the larger turbine unit to accommodate severe furnace load reductions.

•

SPLC compensated furnace load to limit maximum furnace power demand and associated power plant rating. SPLC is a state-of-the-art smelting furnace compensation technology.

•

Closed-circuit condenser cooling water system with cooling tower rated for full heat rejection by the entire power plant.

•

Baghouse to capture fly ash from the CFB boiler.

•

Use of existing, upgraded stack for flue gas produced by both boilers.

18.4.5

Process Plant Load Characterization

18.4.5.1

Power Demand

As with the grid connection, the on-site power plant must be able to meet the nominal and peak power demand of the process plant depicted in Table 18-2.

18.4.5.2

Turndown

In addition to nominal and peak power demands, the process plant will operate in a number of turndown conditions having more modest power requirements. The process plant turndown conditions are summarized in Table 18-5. The most demanding turndown conditions are associated with the smelting furnace load whose power requirements vary from zero to 90 MW.

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Table 18-5: Process Plant Turndown Power Demand Requirements

Planned/
Forced

Frequency

Duration

Nominal Demand [MW]

Turndown Condition

Smelting
Furnace

Other

Total

Smelting Furnace Off

Both

Few hours

Smelting Furnace Idle

Both

Several days

Single Kiln

Both

43-52

Furnace Cold Start

Weeks 1-2

Every few

2 weeks

0-10

5-15

Weeks 3-4

Planned

years

2 weeks

10-20

20-30

Weeks 5-12

8 weeks

20-90

35-105

18.4.5.3

Upset Conditions

In addition to meeting the process plant power demand requirements, the power supply must be able to handle upset conditions. The most demanding upset conditions are those imposed by the smelting furnace loads as outlined in Table 18-6.

Furnace load rejections will occur due to electrical faults or in the event of emergency shutdown. Though it may not be feasible to keep the power plant on-line during such an extreme load rejection, provision should be made for this condition in the design and specification of power generation equipment. As a minimum the power generating equipment should not sustain damage due to such an upset condition.

Electrode breakages also represent an uncommon occurrence. Nonetheless provisions must be made in the power plant for this upset condition. Special provisions are required for the sustained negative sequence current that results from the smelting furnace electrode breakage. The condition may persist for some time until the electrode is replaced.

Loss of arc under an electrode produces a condition akin to an electrode breakage. This condition will not be sustained very long as the electrode hydraulic control will respond to restore the electrode current.

Table 18-6: Smelting Furnace Upset Conditions


Upset Condition		Impact on Power Plant
Furnace load rejection		Instantaneous 100 MW load rejection
Electrode breakage		Instantaneous 50 MW load rejection
		100% negative sequence current on 50 MW base
Loss of arc		Instantaneous 50 MW load rejection
		100% negative sequence current on 50 MW base

18.4.6

Fuel and Boiler Selection

18.4.6.1 Fuel Selection

Several candidate fuels were considered for the on-site power plant including light oil (#2), low sulphur heavy fuel oil (HFO), low sulphur coal and high sulphur petroleum coke (pet coke). These fuels are all available in the region. Natural gas was not considered due to lack of existing pipeline or Liquefied Natural Gas (LNG) infrastructure. Both light and heavy fuel oils were considered due to compatibility with the existing power generation equipment and existing supply and storage systems.

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Table 18-7 summarizes the fuel costs on a MW-h basis for candidate power generation methods including Rankin cycles, reciprocating engine and combustion turbines using typical heat rates. It is clear from this table that a Rankin cycle is the economically preferable option for primary power as it is the only cycle that can accommodate the least cost fuels (coal and petroleum coke). Consequently, despite the existing fuel oil-based power generation equipment, both light oil and HFO are eliminated for prime power generation due to their high cost. However, it can be seen that HFO-fired reciprocating engines represent a viable option for standby power generation.

The outcome of the fuel selection narrowed down the candidate fuels for primary power generation to coal and petroleum coke.

Table 18-7: Fuel Cost as a Function of Cycle on a kW-h Basis


Cycle	Heat Rate kJ/kW-h	Fuel Type	Fuel Cost US $/MW-h
Rankin (no reheat)	11,250	Light Oil (#2)	126.5
		HFO	77.7
		Coal	37.8
		Petroleum coke	20.3
Rankin (with reheat)	9,920	Light Oil (#2)	111.5
		HFO	68.5
		Coal	28.1
		Petroleum coke	17.9
Reciprocating Engine	9,000	Light Oil (#2)	101.2
		HFO	62.2
		Coal	N/A
		Petroleum coke	N/A
Combustion Turbine	12,162	Light Oil (#2)	136.7
(Simple Cycle)		HFO	N/A
		Coal	N/A
		Petroleum coke	N/A
Combustion Turbine	8,500	Light Oil (#2)	95.6
(Combined Cycle)		HFO	N/A
		Coal	N/A
		Petroleum coke	N/A

18.4.6.2

Solid Fuel Boiler Technology

The two proven solid fuel boiler technologies are Pulverized Coal (PC) and Circulating Fluid Bed (CFB). Pulverized Coal (PC) is the more established technology, finding application in boilers rated up to hundreds of megawatts. Circulating Fluid Bed (CFB) technology is a more recent technology commercially proven up to 250 MW. PC boilers employ conventional burners to combust coal at very high temperatures whereas CFB boilers combust coal or petroleum coke in a fluidized bed of fuel, sulphur dioxide sorbent and sand at relatively low temperature. The two technologies were compared on various technical points, as follows:

•

Sulphur Emissions

•

Nitrous Oxide Emissions

•

Fuel Variability

•

Capital Cost

•

Turndown and Load Following

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18.4.6.3

Boiler Selection

The recommended solid fuel boiler technology for facility primary power generation is the CFB type. This is based on its superior emissions control, fuel variability, turndown and load following capability relative to PC type boilers. Given CFB type boiler’s flexibility in solid fuel selection, it stands to reason that the preferred solid fuel is petroleum coke as it represents the lowest cost alternative.

18.4.7

Configuration Selection

18.4.7.1

Existing Power Plant

The existing 60 MW rated power plant was commissioned in 1977 along with the process plant to supply process electrical power requirements. The power plant is based on a non-reheat thermal cycle using heavy fuel oil (HFO) as its primary fuel. Secondary fuel for turndown operation was light oil (#2).

18.4.7.2

Boiler-Turbine Configuration

Given the selection of petroleum coke-fired Circulating Fuel Bed (CFB) boiler-based thermal power generation as the preferred fuel/cycle combination for facility primary power generation, it remained to select a configuration of power generation equipment. The optimal configuration should provide low total annual cost while maintaining reliability commensurate with the process requirements.

Table 18-8 summarizes the boiler/turbine configuration options. The configurations are designated according to the operating boiler and turbine quantities. The operating capacity refers to the rating of the operating equipment on a MW-electrical basis. The equipment is rated to supply both process power and power plant auxiliary requirements.

The backup equipment generally refers to the existing boiler used on a standby basis. In addition emergency generation will be provided by Reciprocating Engine (RE) gensets. This is common to all configurations.

The single contingency capacity refers to the power plant capacity available after an outage of the largest rated operating component. This gives a measure of the impact of a power plant major equipment outage on the process plant.

Generally a common header steam configuration was adopted. This is required to allow integration of the existing 60 MW HFO-fired boiler as a backup unit. The only exception is the 1B/1T configuration where the existing boiler/turbine island would operate as a standalone backup unit.

Table 18-8: Summary of Boiler/Turbine (B/T) Configuration Options

Operating Boiler/Turbine Configuration

Operating Equipment
Capacity (MW)

Total
Operating
Capacity
(MW)

Backup Equipment
Capacity (MW)

Single
Contingency
Capacity
(MW)

Boiler

Turbine

Boiler

Turbine

1B/1T

1 x 150

150

1 x 60

1B/2T

1 x 150

1x 90, 1 x 60

150

1 x 60

N/A

2B/2T

1 x 90, 1 x 60

150

1 x 60

N/A

120

3B/3T

2 x 90, 1 x 60

210

1 x 60

N/A

150

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18.4.7.2.1

Single Boiler/Double Turbine (1B/2T)

This option includes a new full capacity rated boiler coupled to two partial rated turbines including a new unit and the refurbished existing unit. This configuration demands a common steam header. Back-up is provided by the existing HFO-fired boiler coupled to the common steam header. Consequently the new boiler is constrained to operate at output steam conditions commensurate with the existing boiler/turbine. This configuration cannot accommodate reheat as the existing boiler/turbine island does not have it.

Again, this configuration provides low power capacity in the event of an outage. Furthermore, the thermal cycle is constrained by the requirements of the existing turbine. However, the reduced turbine capacity requirement results in a significant capital savings.

18.4.7.2.2

Double Boiler/Double Turbine (2B/2T)

This option includes two new partial rated boilers coupled to new and existing partial rated turbine units. Back-up is again provided by the existing HFO-fired boiler. All units are coupled to a common steam header. Again, the new boilers are constrained to operate at the same output steam conditions as the existing system.

The principal advantage of this configuration is that it provides higher capacity in the worst-case single contingency. However, the installation of two new part rated boilers carries a very high cost premium relative to a full rated boiler.

18.4.7.2.3

Three Boiler/Three Turbine (3B/3T)

This option includes three new partial rated boilers coupled to commensurately rated turbine units including the refurbished existing unit. The existing HFO-fired boiler again provides backup, however it would only be required to operate in the event of a simultaneous outage of two operating boilers – a so-called double contingency. This results in reduced standby use of HFO relative to the other options.

In this option, full power plant capacity is available in both single and double boiler outages resulting in a very high availability. However, this comes at the cost of additional power generation equipment and reduced cycle efficiency as this configuration will normally be operating at a approximately 70% of power plant MCR.

18.4.7.2.4

Alternate Configurations

Alternate configurations that include power generating equipment that deviate from the optimal fuel and cycle selection were studied. These include light oil-fired Combustion Turbine Generators (CTGs) and Reciprocating Engines (REs). Generally these alternate configurations were not cost competitive or were overly sensitive to fuel costs.

18.4.8

Selected Configuration

A total lowest annual cost analysis determined that the optimal power plant configuration is the Single Boiler/Double Turbine (1B/2T). Despite having lower reliability than some of the other options this did not significantly degrade the process plant operating factor of 85%.

Incorporating the refurbished existing turbine into the operating boiler/turbine island resulted in reduced capital cost relative to the Single Boiler/Single Turbine (1B/1T). In addition, the ability to burn low cost solid fuels such as coal and petroleum coke resulted in low operating costs.

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Table 18-9 summarizes the major equipment ratings of the selected power plant configuration. Note that the discrepancy between nominal and actual ratings is associated with the extraction of process steam. However, the overall power plant rating remains approximately 150 MW.

Table 18-9: Major Equipment Gross Ratings of Selected Power Plant Configuration


Major Equipment	Actual Capacity [MW]	Nominal Capacity [MW]	Fuel Type	Comments
New CFB boiler	148.3	150	Solid fuel (coal or petroleum coke)	Will supply both new and existing steam turbines
New Steam Turbine Generator (STG)	84.5	90	N/A	Non-reheat type for compatibility with existing STG.
Existing Steam Turbine Generator (STG)	63.8	60	N/A	Refurbished
Emergency Power Generation	6	5	Heavy Fuel Oil (HFO)	New diesel gensets

18.5

Process and Power Plant Infrastructure, Utilities and Services

18.5.1

Site Preparations and Earthwork

Minimal earthmoving is required, since the site was prepared in 1975 when the Exmibal project was initiated and the site has been reasonably maintained since that time. Structural excavation is limited to new equipment and facilities.

All in-plant roads are granular. The existing roads will undergo some minor and major regrading and resurfacing depending on location. New granular roads will be required around the coal and petroleum coke stockpiles.

18.5.2

Administration and Service Facilities

All of the existing non-process buildings will be refurbished. In addition to the refurbished buildings, the following new ones will be constructed:

•

New canteen (approximately 25 m x 50 m) located east of the smelter building.

•

New fire truck building (the existing one will be demolished to allow room for the SPLC). The new fire truck building will be located west of the first aid room.

An extension to the existing shops building will be constructed to accommodate the mine mobile equipment fleet. The building extension will consist of a four bay workshop, 36m x 16m in plan with an apron on both sides. The extension is located on the east side of the existing building.

A comprehensive list of the non-process buildings is provided in Table 18-10.

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Table 18-10: Non-Process Buildings


Building Name	New/ Existing	Area (m²)	Configuration
Agglomeration	New	130	Structural steel with metal siding
Boiler (small)	Existing	50	Structural steel with asbestos siding
Brick Storage	Existing	1,450	Concrete block
Canteen	New	1,300	Structural steel with metal siding
Carpenters Shop	Existing	280	Structural steel with asbestos siding
Change Room	Existing	210	Concrete block
Construction Camp	Existing	220,000	Various
Dry Ore Storage	Existing	7,560	Concrete walls, siding
Fire Truck House	New	100	Concrete block
First Aid Room	Existing	180	Concrete block
Flammable Liquid Storage	Existing	80	Concrete block
General Office	Existing	1,610	Brick
Guard House (there are 3)	Existing	10 (each)	Concrete block
Laboratory and Office	Existing	980	Concrete block
Maintenance and Utility Office	Existing	290	Concrete block
Mine Office and Changehouse	Existing	570	Concrete block
Mobile Equipment Shops (extension to shops building)	New	580	Structural steel with metal siding
Shops Building	Existing	2,790	Structural steel with asbestos siding
Warehouse	Existing	2,410	Concrete block
Waste Water Treatment	Existing	200	Structural steel with metal siding

18.5.3

Electrical Power Distribution

The electrical power distribution will be based on a grid connection with provision for a future on-site power plant. An air-insulated, 230/34.5 kV substation will connect the Fenix Project to the national grid. The facility primary distribution voltage (34.5 kV) will be used to supply the smelting furnace, the refining furnace, the motive power step-down transformers as well as any compensation equipment.

The motive power step-down transformers will develop the facility secondary distribution voltage of 4.16 kV. This voltage level will also serve large loads. From the main secondary distribution switchboard, power will be supplied to the process areas via radial 4.16 kV feeders.

Each process area will include 4.16 kV switchgear, 4160/480 V unit substation transformers and 480 V switchgear and Motor Control Centres (MCC’s).

18.5.4

Compensation Equipment

In order to facilitate the connection of furnace load to the national grid and the future on-site power plant, compensation equipment is included in the substation design. The compensation equipment includes an SPLC and an SVC.

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The SPLC is in principal an instantaneous demand management system for large smelting furnace loads. It operates as a variable reactance in series with the furnace load that adjusts itself on a cycle-by-cycle basis to limit peak power to a set-point value. Figure 18-5 presents typical SPLC performance data as measured at a reference nickel smelting furnace installation⁴.

The SVC – or Static VAR Compensator – is shunt connected device that performs power factor correction and load balancing on a cycle-by-cycle basis. In addition, the SVC compensates harmonic currents.

LOGO

Figure 18-5: Falcondo Furnace Power With and Without Compensation

18.5.5

Plant Utilities

The utilities required for the process plant include the following systems:

•

Compressed air: Three existing compressors will be refurbished and one new compressor purchased to supply compressed air to the process and power plants.

•

Diesel fuel: Diesel fuel will be required on site for process plant, power plant and mining operations. Diesel will be delivered to the plant in trucks and stored in the existing 660 m³ diesel storage tank. From the central storage tank, diesel will be delivered to four separate storage areas serving each of the main users.

•

Heavy Fuel Oil: Heavy fuel oil will be required on site as an auxiliary firing fuel (for dryer and kiln burners), for warm up of the new 150 MW boiler, and for firing the 5-MW Reciprocating Engine (RE) providing Emergency back-up power.

⁴	SPLC – A power supply for smelting furnaces, 44^th Annual CIM Conference, Nickel and Cobalt 2005

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•

Nitrogen: Nitrogen will be required in the coal mill as inert gas for the emergency system and cleaning of the baghouse. Nitrogen will also be used in the refinery for ladle stirring and the power plant shut-down system. Two 34 tonne liquid nitrogen tanks will be installed with evaporators to meet daily average requirements. Nitrogen will be transported to site by truck once a week.

•

Oxygen: The primary oxygen users are the ferro-nickel refinery and the two rotary kilns. Other miscellaneous oxygen users include furnace taphole lancing and metal launder burners. Two 47 tonne liquid oxygen tanks will be installed with evaporators and excess flow safety systems. Oxygen will be transported to site by truck every other day.

•

Steam: Steam will be required in the process plant for heating and atomization of heavy fuel oil used in the dryer and two kilns. The steam will be available from a package boiler.

•

Fire protection: Fire protection water is drawn from the water inlet channel by a dedicated set of pumps and distributed via the plant fire protection ring main. Electrical and control rooms are provided with smoke detectors, alarms and portable extinguishers, in accordance with current industry practice. The process plant is provided with a fire truck.

18.5.6

Information System (IS) and Telecommunication Infrastructure

18.5.6.1

Information Systems

The various systems are summarized in Table 18-11.

Table 18-11: Information Systems Software Packages


IS & IT Layer		System – Software Packages
Business Planning and Management		IFS Enterprise System
Operations/Production:
Production Information Management		OSISoft
Environmental Management		EQWin
Laboratory Information Management		LabWare
Scheduling		Preactor
Shared Services:
Servers OS		Windows 2003 Server
Relational Database		Microsoft SQL Server
Email Server		Microsoft Exchange
Desktop and Office		Windows XP, MS Office, Internet Explorer, Adobe Acrobat
Documents Management		Microsoft Sharepoint
Engineering		AutoCAD
Development Environment		Microsoft.NET

18.5.6.2

Telecommunications Solutions

The various telecom systems and services are summarized in Table 18-12.

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Table 18-12: Telecommunication Systems


Type		Hardware Description
Plant Site and Harbour facilities		118 Handsets connected to the VoIP phone system
External Telecoms		Satellite Link for 10 Voice Trunks including 256 kbps access for data links GSM Cellular for 30 Cell Phones using Guatemala GSM Network accessible at El Estor (Comcel)
Internal Telephone		VoIP and PBXs (PA System linked in)
Internal Radio		Trunked UHF System connected to PBX, 3 Transmitters, 50 Portable units at the plant. 4 Portable units at the harbour facility
Internal Network		Optical Fibre between areas and between Buildings for Data
		Copper Cabling for Phones and Data within Buildings

18.5.7

Water and Waste Management

18.5.7.1

Water

The site water management plan has been designed such that under dry weather conditions, the plant will be a zero discharge facility, excepting only the treated sewage effluent. During the wet season, stormwater will be collected and recovered for plant consumption, to the maximum extent possible. Stormwater runoff that exceeds the stormwater pond capacity, will overflow to the Fenix water inlet channel.

Lake Izabal water will flow into the inlet channel to the raw water treatment plant, where the raw water will pass through strainers and will be treated with biocide (chlorine). This treated raw water will then be supplied as make-up water to the cooling water system, the potable water treatment plant, the demineralization plant and the Slag Cooling Pond.

The October 2006 Feasibility Study site water balance (refer to Table 18-13) indicates that the process water demand will be on average 693 m³/h. 3.5 m³/h (nominal) will be required for potable water (this includes the requirements for the new power plant). The potable water plant will supply potable water to both the Process and Power Plants and the Mine Site. Water will be trucked from the potable water plant to the Mine Site on a regular basis.

Similarly, all sewage generated at the mine site will be trucked back to the sewage treatment plant located at the Process/Power Plant for treatment. The treated effluent from the sewage treatment plant will flow to the outlet channel.

To minimize the water makeup required by the slag granulation system, blowdown streams from the various cooling water systems and ferro-nickel shotting system will be pumped to the Slag Cooling Pond, as makeup water.

Runoff from the Plant Site will be collected in Stormwater Pond No.1 where suspended solids will settle out. The treated run-off will then be returned to the inlet channel. Runoff from coal/petroleum coke/lime storage piles at the Plant Site will be collected in Stormwater Pond No.2 then recycled for dust suppression on the storage piles, or pumped as make up to the Slag Cooling Pond. Run-off from the Slag Pile will be collected in Stormwater Pond No.3 where suspended solids will settle out. The treated run-off will then be returned to the natural watercourse. The stormwater pond at each location is sized to provide total containment of stormwater runoff for rainfall events up to and including the 10-yr, 24-hr duration, storm event. For storms greater than the 10-year event, a safe flow path to the pond will be ensured.

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Plant site grading, roads and ditches will be designed to ensure that there is a safe flow path for storm events up to the 1 in 100 yr event, such that the runoff will not flood key process areas, cause significant erosion, or cause other problems.

The Fenix Site facility design will incorporate secondary containment of fuel oil, acid, caustic, waste liquid storage tanks, as well as any other process liquid that poses an environmental risk if released to the environment.

At the CGN Santo Tomas trans-shipment facility, precipitation run-off from the coal/petroleum coke storage piles will be collected in Stormwater Pond No.4 then recycled for dust suppression on the storage piles. Excess runoff will be neutralized before it is returned to the sea.

Table 18-13: Summary of Water Users (excludes Backwash Flows and Other Losses)


	Average flow (m³/h)	Peak Flow (m³/h)	Design Flow (m³/h)
*Raw Water*
To Firewater system	0	1136	1136
Total Raw Water	0	1136	1136
*Filtered Raw Water*
Slag Granulation Makeup	110	225	225
Fe Ni Shotting Makeup	20	92	92
To Potable Water Treatment	9	9	9
To Demineralization	21	42	42
Process Plant Cooling Makeup	33	53	53
Power Plant Cooling Makeup	445	445	445
Washdown Water	21	34	34
To Thickener Overflow Tank	29	131	131
To Spray Cooling	2	20	20
Dust Suppression	4	12	12
Total Filtered Raw Water	693	1063	1063
Total Water Demand	693	2199	2199

18.5.7.2

Non-Process Waste Disposal

The waste management system for the Fenix Feasibility Study includes the collection, transportation and disposal of various wastes types from the Plant Site, Power Plant and Mine Site over the 25-year design life of the project. The disposal facilities include the sanitary landfill, used tire storage, used battery storage and a hazardous liquids storage facility. In addition, efforts will be made to use local recyclers when the industry develops.

The sanitary landfill location is adjacent to the western side of the slag repository, downwind of the populated working areas (see Hatch drawing 319112-0000-G-003). A detailed hydrogeologic and geotechnical study is in progress to confirm the suitability of the landfill location.

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18.5.7.3

Process Waste Disposal

Three types of process waste materials will be generated at the Fenix site:

•

Electric Furnace Slag: Granulated slag will be trucked from the truck loading bins to the slag repository located immediately west of the process plant. The haul truck will operate 24 hours per day, making approximately 4 trips per hour. A bull dozer will be used to stack the slag in benches. The slag pile will be built up in sections of five year’s capacity each. Once one section has been filled it will be blanketed with top soil and reforested.

The plant will produce approximately 1,188,000 tonnes of slag per year.

•

Refining Slag: Approximately 7700 tons of refining slag will be produced per year. It will be processed through the metal recovery system and the residual slag transported to the slag repository by truck for disposal along with the electric furnace slag.

•

Power Plant Ash: Ash will be transported from the storage silos to the ash disposal site by dump truck. The construction plans for the new power plant include the construction of an ash disposal cell sized for 3 years of ash. During the life span of the nickel plant, additional cells for ash disposal will be constructed as needed. Once the ash has been dumped it will be compacted and periodically sprayed with water to keep the surface moist to manage dusting. Once a cell has been filled, it will be capped with common fill, covered with topsoil and reforested.

Groundwater monitoring wells will be installed to monitor groundwater quality. Drainage from the cells will be directed to a drainage pond that will collect water for recycling for dust control purposes. Surface water run-off from the ash disposal cells will be directed to a stormwater management pond by gravity flow using open channel drainage ditches. Clean surface runoff from external (i.e. upstream) areas will be intercepted by perimeter ditches and conveyed around the disposal cells to avoid potential contamination of clean runoff water.

18.6

Operations Logistics and Transportation

18.6.1

Transport Options

Several alternatives for the shipping of major bulk consumables (primarily coal and petroleum coke) to the project site were considered. In all cases, it was assumed that these consumables would arrive in Handy-size⁵ vessels via the Atlantic port of Santo Tomas de Castilla, located on Amatique Bay. This is because: 1) Santo Tomas is much closer to the project site than the Pacific port of Quetzal; 2) the major regional sources of petroleum coke and coal ship their product via Atlantic or Caribbean ports, and; 3) vessels are limited to Handy-size due to draft restrictions at Santo Tomas. The options considered included the following:

•

Trucking of the bulk materials from the port of Santo Tomas to the plant site via highways CA9, CA13 and 7E.

•

Barging of bulk materials in shallow-draft barges (to cross the shallow bar at Livingston) from ships at anchor in Amatique Bay up the Rio Dulce, through the Golfete, and across Lake Izabal to the plant site.

⁵	Cargo ships of 28,000 to 40,000 dwt capacity

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•

Trucking of bulk materials to a trans-shipment terminal on the shore of Lake Izabal and ferrying the trailers across the lake to the plant site. Two sites were considered for the trans-shipment terminal: Finca Colorado on the eastern shore of the lake; Mariscos on the southern shore.

The first option, trucking from Santo Tomas to the plant site, was adopted, due to perceived environmental and socio-economic risks associated with river and/or lake transport. Hatch Mott MacDonald (HMM) completed a site audit of the existing public highways (CA9, CA13 and 7E) between Santo Tomas and the plant site. A preliminary cost estimate was developed including the realignment of a section of highway 7E that passes through the town of Rio Dulce.

CGN is negotiating an agreement with the Guatemalan government whereby in return for the government upgrading the highways as per the HMM recommendations, CGN will pay for the maintenance of Highway 7E from Panzos (west of the Fenix plant) to Fronteras (junction 7E and CA13) and the maintenance of a gravel highway from 7E to Cahabon, during the operational life of the Fenix facility. In addition, the government will divert the funds it would normally have spent on maintaining the sections of 7E that CGN will be maintaining, to improved maintenance on those parts of highways CA13 and CA9 that the Fenix operation will be using.

18.6.2

Material Transport Requirements

The Fenix plant site is located approximately 130 km (by road) from the nearest commercial port, the port of Santo Tomas de Castilla. A significant challenge is the delivery of consumables to the plant site and the delivery of product to the market.

The major Fenix consumables are:

•

Coal: Low-sulphur thermal coal is required in the process plant principally as a source of thermal energy (for firing the dryer and kilns) as well as providing a source of carbon (reductant) for the metallurgical process. The coal will be sourced in bulk, most likely from Colombia and/or Venezuela.

•

Petroleum Coke (petcoke): Petroleum coke is required as a source of thermal energy for on-site power generation. It will be sourced in bulk at oil refineries in the U.S. Gulf and/or in Aruba and Venezuela. Petroleum coke will not be required until the new power plant is ready for service (expected to be October 2014)

The Fenix mine and plant will produce high-grade ferro-nickel (FeNi) shot that will be exported to Asian, European and North American markets. The FeNi will be packaged in two-tonne Flexible Intermediate Bulk Containers (FIBCs, or ‘bulk bags’). These will be consolidated in 20-tonne shipping containers for export to market.

The Fenix transport requirements will gradually increase during the first three years of operation, as the plant ramps up to its design capacity. After the third year of production the process plant transport requirements should be relatively consistent, with minor variations occurring according to equipment downtime. After the new power plant begins operation in year five, however, the transport requirements will increase significantly as petroleum coke is delivered to site.

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Table 18-14 lists the estimated Fenix transport requirements.

Table 18-14: Fenix Consumable and Product Transport Requirements


			Fenix Annual Requirements
Description	Units	Year 1	Year 2	Years 3 - 5	Years 6+	Source of Supply
Bulk Consumables
Pet Coke	Wet t/y	0	0	0	305,000	US Gulf Coast, Aruba
Coal	Wet t/y	109,850	224,900	274,350	275,000	Colombia, Venezuela
Secondary Consumables
Heavy Fuel Oil	t/y	830	840	840	4,000	Santo Tomas
Diesel	t/y	1,630	2,150	2,380	2,400	Santo Tomas
Lime (refinery)	t/y	2,212	4,056	4,950	5,000	Guatemala City
Dolime	t/y	360	667	810	900	Tampico, Mexico
Fluorspar (CaF₂)	t/y	452	829	1,010	1,100	Tampico, Mexico
Al	t/y	425	779	950	1,000	Santos
Al2O3	t/y	500	928	1,130	1,200	China
Ferrosilicon (FeSi)	t/y	54	100	120	120	China
CaSi wire	t/y	71	131	160	160	China
Electrode Paste	t/y	800	1,651	2,000	2,010	Vitoria, ES, Brazil
Oxygen	t/y	1,417	2,785	3,390	3,700	Guatemala City
Nitrogen	t/y	962	1,954	2,380	1,400	Guatemala City
Product
Ferro-nickel	t/y	31,175	57,150	69,760	70,500	Fenix Plant Site

18.6.3

Fenix Bulk Consumable/Product Transport System - Description

A summary description of the petroleum coke, coal and FeNi transportation system is provided below.

•

Petroleum coke (or coal) will be transported by Handy-size freighter (27,500 tonne cargos) to the Guatemalan port of Santo Tomas. Ships will tie up at the port’s Berth #1, designated for bulk solids cargos.

•

Using the ship’s cranes, the freighter will transfer the cargo to four mobile wharf hoppers. The wharf hoppers are designed for truck loading. The freighter will discharge its cargo in approximately 3.5 days.

•

For the 3.5 day freighter unloading period, a fleet of 16 tractor-trailers (25 tonne capacity, rear dump) will operate 24 hours/day to transfer the petroleum coke (or coal) from the wharf hoppers to the CGN/Santo Tomas trans-shipment facility.

•

The CGN/Santo Tomas trans-shipment facility is located approximately 3 km (road distance) from the wharf. The trucks will transport the petroleum coke (or coal) to the trans-shipment facility where bulldozers will push it into stockpiles.

•

Front-end loaders will load petroleum coke (or coal) from the trans-shipment stockpile into trailers (25 tonne capacity, rear dump). A fleet of trucks will operate 6 days/week, 24 hours/day to transfer full trailers from the CGN/Santo Tomas trans-shipment facility to the Fenix plant site.

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As shown in Figure 18-6, the transportation route to the site consists of three distinct segments. The first segment extends west of Santo Tomas, along highway CA09 towards Guatemala City for a distance of 51 km. The vehicles will then turn north onto the second segment, highway CA13 at La Ruidosa and continue over the Rio Dulce River into the community of Fronteras. The second segment, from La Ruidosa to Fronteras, is 30 km long. The third and final segment of the journey extends west across the north shore of Lake Izabal, on road 7E, and continues to the plant site, just west of El Estor for a distance of 49 km. The total distance from Santo Tomas to the Fenix site is approximately 130 km.

LOGO

Figure 18-6: Bulk Material Transport Route

•

Ferro-nickel, packaged in 2-tonne bulk bags, will be loaded onto returning trailers (after year 5, about 10% of the trailers will be required for backhauling FeNi). The trailers containing FeNi will be delivered to a near-port export distribution warehouse. Export orders will be loaded into ocean containers, each with 20 metric tonnes of product. The containers will be dispatched via the Atlantic ports of Santo Tomas and Puerto Barrios for North American and European destinations, and to Puerto Quetzal on the Pacific coast of Guatemala for exports to Pacific Rim markets.

18.6.4

Bulk Consumables/Product Transport Logistics

A schedule for the transport of bulk materials to site, based on the quantities shown in Table 18-14, is presented in Table 18-15 and Table 18-16. The trucking requirements for backhauling ferro-nickel are presented in Table 18-17. The trucking values are based on 310 day/year truck and barge operation.

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Table 18-15: Petroleum coke and Coal Freighter Requirements


		Annual Shipping Requirements (# Ships)
Description	Freighter Capacity	Year 1	Year 2	Years 3-5	Years 6+
Petroleum coke	27,500 tonnes	0	0	0	11
Coal	27,500 tonnes	4	8	10	10

Total		4	8	10	21


Table 18-16: Daily Trucking Requirements (Santo Tomas - Fenix)

		Daily Trucking Requirements (# Truckloads)
Description	Truck Capacity	Year 1	Year 2	Years 3-5	Years 6+
Petroleum coke	25 Tonnes	0.0	0.0	0.0	39.4
Coal	25 Tonnes	14.2	29.0	35.4	35.4

Total		14.2	29.0	35.4	74.7


Table 18-17: Trucking Requirements for Backhauling Ferro-nickel

		Daily FeNi Trucking Requirements (# Truckloads)
Description	Truck Capacity	Year 1	Year 2	Year 3-5	Years 6+
FeNi (backhauled)	24 tonnes	4.2	7.7	9.4	9.5

18.6.5

Transport of Secondary Consumables

Secondary consumables, including heavy fuel oil and diesel, will be transported to the plant site by tractor-trailer via public roads CA9, CA13 and 7E. Daily trucking requirements, based on the quantities provided in Table 18-14 and assuming trucking 310 days/year, are provided in Table 18-18.

Table 18-18: Daily Trucking Requirements for Secondary Consumables


Description	Truck Capacity	Year 1	Year 2	Years 3-5	Years 6+	Comments
Power Plant Heavy Fuel Oil	22 tonnes	0	0	0	0.5	Tanker truck
Process Heavy Fuel Oil	22 tonnes	0.1	0.1	0.1	0.1	Tanker truck
Diesel	20,000 L	0.2	0.3	0.3	0.3	Tanker truck
Lime (refinery)	20 tonnes	0.4	0.7	0.8	0.8	2 tonne bulk bags
Dolime	20 tonnes	0.1	0.1	0.1	0.1	2 tonne bulk bags
Fluorspar (CaF₂)	20 tonnes	0.1	0.1	0.2	0.2	2 tonne bulk bags
Al	20 tonnes	0.1	0.1	0.2	0.2	2 tonne bulk bags
Al₂O₃	20 tonnes	0.1	0.1	0.2	0.2	2 tonne bulk bags
Ferrosilicon (FeSi)	20 tonnes	0.01	0.02	0.02	0.02	2 tonne bulk bags
CaSi wire	20 tonnes	0.0	0.0	0.0	0.0
Electrode Paste	20 tonnes	0.1	0.3	0.3	0.3	2 tonne bulk bags
Oxygen	21.5 tonnes	0.2	0.4	0.5	0.5	Tanker truck
Nitrogen	21.5 tonnes	0.1	0.3	0.4	0.4	Tanker truck
Miscellaneous Supplies		1	1	1	1
Total Trucks per Day		2	4	4	5

Trucking requirements based on 310 days/year truck operation

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18.6.6

Off-Site Infrastructure and Facility Requirements

18.6.6.1

Upgrades to Public Highways

Hatch Mott MacDonald (HMM) completed a visual inspection of the roads and bridges along the proposed transport route (GISystems also completed a survey of highway 7E). Structural drawings were obtained for 23 of 31 bridges and preliminary structural evaluations completed. Improvements to the road geometry and pavement structure were considered.

The evaluation considered load requirements of the capital project (i.e. transport of equipment and materials required for the construction of the Fenix plant) and the operating facility (i.e. transport of bulk materials such as coal and petroleum coke that are required for plant operation).

18.6.6.1.1

Atlantic Highway CA9

The CA9 highway is the main truck route between Guatemala’s Atlantic ports of Santo Tomas and Puerto Barrios and Guatemala City. It is a paved, two lane, well graded, highway and is in good condition. The section of road between the CGN Santo Tomas trans-shipment facility and the intersection with CA13 at La Ruidosa is 47 km long and includes thirteen bridges.

HMM concluded that the CA9 pavement structure is acceptable and suggested only minor improvements (i.e. additional signage, pavement marking, rumble strips). Three bridges will require significant upgrades; it is anticipated that the upgrades can be completed without disrupting traffic by strengthening steel girders in situ.

Skye/CGN contracted the services of GISystems Soluciones Inteligentes (GISystems) consultants to conduct several road studies of highways CA9 and CA13 and roads RD4 and 7E. A survey of the existing traffic density on highway CA9 was completed. Based on this information the impact of the proposed CGN truck/trailer traffic of 75 roundtrips per day, 6 days per week, between Santo Tomas and Mariscos is calculated to increase traffic density on CA9 by about 8%.

18.6.6.1.2

Highway CA13

Highway CA13 is the major north-south arterial route along the border of Belize. It is a paved, two lane, reasonably graded, highway and is in moderate condition The section of CA13 from the intersection of CA9 near La Ruidosa to Rio Dulce is approximately 30 km long and includes three bridges. One of the bridges, the 860m long Rio Dulce Bridge, is a high-level crossing that spans a navigable water way.

HMM concluded that the CA13 pavement structure is acceptable and suggested only minor improvements (i.e. additional signage, pavement marking, rumble strips). All three bridges will require minor upgrading. A geometric upgrade of the intersection of highways CA9 and CA13 may be required to improve truck mobility.

18.6.6.1.3

Regional Highway 7E

Highway 7E is a two lane regional road running along the north shore of Lake Izabal. The road from Rio Dulce to the Fenix Site at El Estor has two distinct sections:

•

7E/1 is 13 km long and connects Rio Dulce, located at the junction with CA13, with the town of Sumache. The road surface is chip sealed, and the alignment is narrow and winding with limited sight lines. The road is in poor condition. There are four bridges.

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•

7E/2 is 36 km long and connects Sumache to the Fenix site. The road has a gravel surface and is wide enough for two lanes of traffic in most places. During heavy rains it is the road near the plant site that could become impassable. There are eleven bridges along this stretch of road.

A Japanese funded project to upgrade highway 7E is in the design phase. The project includes:

•

A bypass of the town of Rio Dulce and a new geometric intersection with highway CA13.

•

Widening the road surface (3.25m wide lanes).

•

Upgrading the pavement structure.

HMM concluded that the pavement structure should be upgraded beyond that proposed for the Japanese funded contract and the road widths should be increased to a minimum of 3.35m. Furthermore, most of the bridges will require modifications including the replacement of one bridge and major reconstruction of five bridges.

18.6.6.2

Trans-shipment Terminal at Santo Tomas

CGN has purchased a lease option for one year renewable for a further year for a near-port site for the bulk trans-shipment operation. The 73,657 square meters (7.4 hectares) site, is located a distance of 3.2 km from the port gate. The terminal will receive petroleum coke and coal from the ships for stockpiling and re-loading to bulk trailers to be delivered to the Fenix Plant.

The trans-shipment terminal facility will accommodate the following structures and activities:

•

The bulk storage area will consist of two 60,000 tonne capacity stockpiles with a geomembrane (HDPE) liner under the stockpiles and a 1.0 meter backfill cover separated by a settling pond collecting the surface water run-off during periods of rain. Dust control will be by water sprinkler system.

•

A truck-scale will be located inside the terminal at a location which facilitates the flow of truck-trailer units to be weighed as they return empty to the terminal and to be re-weighed as they depart loaded from the terminal.

•

A truck tire washing pond will be located at the exit before the truck scale.

•

A perimeter block-wall of 3 meter height will provide security and help minimize dust being carried outside the storage area during high winds.

•

A terminal entry and exit gate, access and egress and a road system will facilitate traffic within the terminal area.

•

Parking area for a total of 35 trailers will be provided inside the terminal.

•

Office building (or office-trailer) will have communications links (phone/internet), wash room, change room, lunch room with an adjacent tool shed.

•

An additional area of 8,000 m² opposite the terminal entry gate is reserved for additional truck/trailer parking and holding area.

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18.7

Metallurgical Process and Recoverability

The product is dry ferro-nickel shot containing 35% by weight of nickel, packaged in bags for transport to customers and is produced utilising the well-established Rotary Kiln–Electric Furnace (RKEF) technology. The processing line of the plant, shown as a process flowsheet block diagram in Figure 18-2, consists of one dryer, two reduction kilns, one electric furnace and a ferro-nickel refinery. Average ore assays from the mine plan developed by Snowden are shown in Table 18-19.

Table 18-19: Average Ore Assay, dry wt%


Component	Avg. Y1 – 30	Avg. Y3 – 22
Ni	1.63	1.72
Fe	18.18	18.5
Co	0.04	0.04
SiO₂	32.0	32.2
MgO	20.8	20.8
Loss on ignition (LOI)	11.7	11.7
Cr₂O₃	1.1	1.1
Al₂O₃	3.8	3.7
S	0.02	0.02
Other	2.0	1.4
SiO₂/MgO	1.54	1.54
Fe/Ni	11.2	10.7
Moisture Content	34	34

Run-of-mine (ROM) ore is trucked from the mine to the process plant site, where it is crushed and stacked in one of two wet ore stockpiles. A travelling stacker builds the stockpiles in a manner that promotes ore blending, thereby providing a consistent feed to the downstream processes. The blended ore is reclaimed from the wet ore stockpile by front-end loader and transferred to the coal-fired rotary dryer. Dryer product at about 20% moisture is conveyed to the dry ore storage building and laid down into one of four stockpiles, which provide inventory ahead of the reduction kilns and blending capability to ensure that a consistent feed satisfying the process ore chemistry requirements is delivered to the process plant. Dryer off-gas is cleaned in an ESP before being released to atmosphere and the collected dust is recycled.

Ore from the dry ore stockpiles is reclaimed by front-end loader and transferred to the two rotary reduction kilns operating in parallel. The kilns, which are fired with pulverised coal, are equipped with coal scoops, dams, lifters and on-board tertiary air fans, all of which improve kiln productivity and efficiency. In the kilns, the remaining free and crystalline moisture is removed, the solids temperature is raised to 900ºC and the calcined ore is partially reduced.

Hot calcine is discharged from the kilns into transfer containers mounted on rail cars. The containers are driven horizontally from the kiln to the smelter building and hoisted by cranes to the electric furnace feed bins. Nine covered, insulated feed bins discharge calcine into the furnace through 24 feed pipes. Calcine is smelted in the 90 MW, 18m diameter electric furnace and separated into ferro-nickel and slag. The selective reduction of the iron and nickel oxides is completed by residual carbon in the calcine to give a crude ferro-nickel containing 35% Ni. The ferro-nickel is tapped intermittently into ladles and transferred to the refinery. Slag is tapped from the furnace semi-continuously, water-granulated, loaded onto trucks by clam shell crane and transferred to the slag repository.

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In the refinery, impurities are removed from the crude ferro-nickel to the levels specified for the final product. The refining line consists of a heating station, a slag removal station and a ladle furnace. The crude ferro-nickel is chemically heated and sulphur is removed using a highly basic lime-based slag. Molten refined ferro-nickel is sent to a shotting tank where it forms a solid product with a size range of 3 mm to 30 mm. The ferro-nickel shot is dried and packaged for shipment to market. The refining slag is sent to a metal recovery area where the slag is crushed and ferro-nickel recovered using magnets.

All dusts from process streams, secondary emissions and plant hygiene systems are collected and pneumatically conveyed to the dust recycle area for recycling into the process. The dusts are mixed with water in a pug mill to form a paste which is returned to the rotary dryer along with wet ore.

Estimated nickel losses, expressed as a percentage of nickel input to the process plant, are shown by source in Table 18-20 for the first 20 years after ramp-up, and are based upon evaluations of operational data from ferro-nickel plants. A 0.16% nickel content in electric furnace slag was estimated for the production of a 35% Ni ferro-nickel. The material handling and fugitive emissions losses are predicted to be 3.0% of the nickel input during the first four operating years. This value is anticipated to gradually decrease to 2.0% by the seventh year of plant operation based on increasing operational experience and plant optimisation.

Table 18-20: Sources of Nickel Losses


Source of Loss	Nickel Loss, % of Nickel Input to Process Plant
Source of Loss	Years 1 - 4	Years 7 - 22
Furnace Slag (varies with ore grade)	5.5 – 6.7	7.0 – 8.3
Refining Slag	0.1	0.1
Fugitive Emissions and spills	3.0	2.0
Emissions from Stacks	<0.02	<0.02
TOTAL LOSSES	8.6 - 9.8	9.1 – 10.4
NICKEL RECOVERY	90.2 – 91.4	89.6 – 90.9

The operating factor for the Fenix ferro-nickel facility is predicted to be 85%, and was derived from Hatch’s in-house knowledge of similar laterite smelting operations.

Nickel production in the early years of operation of the RKEF processing line requires adjustments for production losses that will occur as the equipment is commissioned and plant personnel are learning how to control and maintain the process. The ramp-up allowance accounts for all events that will lower production in the early years of operation, including, throughput restrictions, increased maintenance outages and reduced nickel recovery. A 2-year ramp-up was estimated for both kilns to achieve full nominal production with the second kiln beginning operation 3 months after the first kiln. The ramp-up factors expressed as a percent of full production for each kiln are shown in Table 18-21.

Table 18-21: Ramp-up for Each Kiln


Year	Kiln No. 1		Kiln No. 2		Total
1	50	%	32	%	40	%
2	85	%	80	%	82	%
3	100	%	99	%	99	%
4	100	%	100	%	100	%

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The data from the mine plan, nickel recovery, ramp-up schedule and operating factor were used to calculate the production levels with time, shown in Figure 18-7. Estimated total nickel production over the life of the project (years 1-30) is approximately 604,000 tonnes with an estimated peak nickel production of approximately 26,400 tonnes in year 4. The average nickel production over years 3 to 22 (20 years after initial ramp-up) is 22,000 tonnes.

LOGO

Figure 18-7: Nickel Production

(Note the 3 sharp reductions in production reflect planned furnace rebuilds in those years.)

18.8

Markets and Sales Agency Agreement

18.8.1

Ferro-nickel Marketing

The Skye Fenix Project plans to begin production of ferro-nickel in 2009. Production will ramp-up to 58 million pounds (26,400 metric tons) of nickel contained in ferro-nickel in 2012. The Fenix project’s capacity is the equivalent of approximately 60,000 metric tons of ferro-nickel containing 35% nickel per year.

18.8.1.1

Demand for Primary Nickel

Nickel is a critical component of materials used in modern society. As an important component of stainless steels, alloy steels, high temperature-resisting, corrosion resisting, cryogenic, and electronic alloys, nickel use has grown at a rate exceeding that of the major metals including iron and steel, copper, and aluminium over the last several decades. The properties that nickel imparts to the metals in which it is incorporated are of increasing importance to consumers ranging from the most sophisticated equipment designer to the household buyer of home appliances. Over the last twenty years, from 1985 to 2005, overall consumption of primary nickel has increased at 5% per annum.

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Total demand for primary nickel for all uses in 2015 is forecast to be approximately 2.01 million metric tons per year, of which approximately 70%, or 1.38 million tons will be consumed in the production of stainless steel. The Fenix project’s output of 24,312 tons of nickel in 2015 will represent approximately 1% of nickel demand in that year.

18.8.1.2

Nickel in Stainless Steel

Ferro-nickel is used by specialty steel companies across the globe in the production of stainless steel. Production of austenitic stainless steel is by far the single largest consumer of nickel in all forms, using an estimated 70% of the world’s total in 2005.

Stainless steel production has been the most important growth stimulant for nickel in the last ten to fifteen years. Excluding the former Soviet Union, stainless steel production grew at almost 6% per annum from 1990 to 2005. In this period the traditional industrialized areas, the US, Japan, and Western Europe experienced modest growth in stainless steel production, while the Asian region, led by Korea, Taiwan, India and China achieved dramatic growth. China alone increased stainless steel production at a rate of 25% per annum from 1990 to 2005, while Taiwan grew at an 18% rate and Korea at 14%.

Growth in stainless steel production of 5% per annum globally is expected to continue for the next decade, with China the most important factor. Even with Chinese stainless steel growth projected to decline from a 2000-2006 rate of 46% to 15% between 2005 and 2015, its impact on the demand for primary nickel will be very large.

18.8.1.3

Nickel Demand Growth

Overall, the demand for primary nickel from 2005 to 2015 is projected to grow at 4.8% per annum. Nickel demand from the non-stainless markets, for aerospace and corrosion resisting alloys, plating and foundries, and in developing applications like batteries will grow, but on average, at rates expected to be lower than for stainless steel.

Between 2005 and 2015, an additional 750,000 tons per annum of primary nickel over and above the 1.28 million tons produced in 2005 will be required to satisfy the projected demand.

18.8.1.4

New Capacity Required

A prolonged period of low nickel prices in the 1990’s resulted in a low rate of investment in new projects in the first half of the current decade. Development of new sources of nickel was further delayed by technical difficulties in the construction and start-up of two laterite nickel projects in Australia (Murrin Murrin and Bulong) and by development delays of the major discovery made by Inco at Voisey’s Bay in Newfoundland, Canada.

As a result of strong growth in demand in recent years and high nickel prices since the start of 2004, nickel mine development has received a boost, and a substantial amount of new mine capacity is now due to come on stream in 2007-11. The net forecast addition to production in that period is 524,000 tonnes of contained nickel, equivalent to 36% of world mine production in 2006. The major projects that will contribute to this growth include CVRD Inco’s Goro and Onca Puma projects, BHP Billiton’s Ravensthorpe project and Anglo-American’s Barro Alto project.

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18.8.1.5

Fenix Favourable Cost Position

The projected ferro-nickel cost curve shows that once a solid-fuel based power plant is in service, Fenix should be in the lowest quartile in cash production cost among all planned and existing producers. Based on interim power supply under the PPA, Fenix will fall into the third quartile of cash costs. CRU Strategies estimated the position of Fenix on the cost curve based on data provided by Skye.

LOGO

Figure 18-8: Forecast Nickel Cost Curve

18.8.1.6

Pricing

Nickel price forecasts typically consider the industry cost curve, the expected supply-demand balance, and the price required to justify new investments in capacities to replace exhausted deposits and meet growing market needs. Historical prices may also be a guide to the performance of future nickel prices, given the projected changes in the basic price factors.

Nickel producers, investors, and a number of financial institutions and industry consultants make and maintain nickel price forecasts. In the last few years the long term nickel price forecasts (in constant US dollar terms) made by the financial community in particular have increased substantially. This presumably reflects a combination of the positive long term market growth outlook, especially for stainless steel, the high price levels in the last several years which appear to be driven by a lack of new supply in the face of relatively strong demand, the substantially higher cash operating costs caused by increasing energy prices, the weaker US dollar relative to currencies of several of the main nickel producing nations (Canada and Australia, for example), and the high capital cost per unit output of those expansion projects which are under way and under discussion by several of the large industry participants.

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The following table summarizes historic pricing (expressed in $2006).

Table 18-22: Historical Nickel Prices


$/lb Ni	1925-2006		1925-1979		1980-2006		1997-2006		2002-2006
Lowest Year	$	2.60	$	3.01	$	2.60	$	2.60	$	3.44
Lowest 5 yrs	$	3.40	$	3.35	$	3.40	$	3.40	$	3.40
Lowest 10 yrs	$	3.75	$	3.61	$	3.75	$	3.75
Lowest 20 yrs	$	4.70	$	4.21	$	4.70
Average	$	5.12	$	5.07	$	5.22	$	5.00	$	6.49
Highest 20 yrs	$	6.46	$	6.18	$	6.46
Highest 10 yrs	$	7.16	$	7.11	$	7.16	$	5.72
Highest 5 yrs	$	7.34	$	7.47	$	7.34	$	6.49	$	6.49
Highest Year	$	10.65	$	7.97	$	10.65	$	10.60	$	10.60

Source:

Data, LME

Over the longest term, from 1925 to 2006, the price of nickel averaged $5.12 per pound in 2006 dollars, while over the past 27 years; from 1980 to 2006 it averaged $5.22. In 2006 the average cash nickel price on the LME was $10.99 per pound; prices in 2007 to date have exceeded $20.00 per pound as a result of extreme tightness in the nickel market resulting from a shortfall of supply against a backdrop of growing demand. In mid 2007 prices have fallen back to the $15.00 per pound range.

18.8.1.7

Fenix Feasibility Study Price Assumption

As a result of growth in nickel supply together with moderation in demand growth for nickel, CRU Strategies expects a steady decline in nickel prices, from a peak in 2007 to an average of $6.79/lb in real 2007 terms by the year 2011. Despite the major forecast decline in prices from current levels, CRU Strategies expects the LME nickel price to converge towards a level that will be well above the long run average that prevailed up till 2003. CRU Strategies’ trend forecast of $6.50/lb in real terms is based on the price that they believe will be necessary to encourage and sustain investment in the new generation of laterite leaching projects that the industry will need in order to generate the supply required to meet long term demand. This price includes capital charges, in addition to cash operating costs. Capital cost escalation is expected to be a persistent feature of the industry, and capital costs per lb of nickel production can well average $3.00/lb. CRU Strategies’ price forecast is set out in the chart below. Given this forecast a constant price of $6.50 per pound of nickel has been selected for the purposes of the Fenix economic analysis.

Table 18-23: Forecast Real 2007$ Nickel Prices


	$/lb Ni
2007	$	17.35
2008	$	10.47
2009	$	8.15
2010	$	6.31
2011	$	6.79
Long Term	$	6.50

Source: Data, CRU Strategies (August 2007)

In addition, a constant dollar scrap steel price of US $300 per tonne will result in an estimated long term iron credit for the Fenix project’s ferro-nickel of $.20 per pound of contained nickel.

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18.8.2

Sales Agency Agreement

Under the terms of the Master Agreement with Inco Ltd. Skye has agreed that Inco will be its exclusive sales agent for the ferro-nickel produced at Fenix. In return for receiving a commission of 2.75% on the net invoice value of the sales contract Inco will provide marketing, contract administration and other services related to product sales.

18.9

Environmental Considerations

18.9.1

Environmental Bonding

In compliance with Guatemalan environmental legislation, CGN posted environmental bonds as part of the approval of the Mine EIA of $US 100,000 and more recently for the four Fenix Project EIA’s totalling Q 17,828,000.

18.9.2

Mine Reclamation Plan

Limited waste storage external to pits will occur during the initial years of mining but ultimately all pits will be backfilled over the life of the mine. This includes compacting and grading backfill slopes, restoring any exposed slopes created for ditching and refilling any unused in-pit runoff controls such as sumps and ditches. Stockpile and backfill heights will not exceed 10m and a slope of 27 degrees. Compaction is to be carried out by routing haul traffic over the fill; in some areas a sheep’s foot compactor will be used.

The suspended solids generated by mining will be contained before they leave the property. World Bank standards for discharging suspended solids will be complied with for the selected design event. The pond sedimentation will be sized to remove suspended sediment for the 10-year return period 24-hour flow. Runoff collector ditches are sized for peak flows from the 10-year to 24-hour storm:

•

The mine storm water handling system will essentially have two levels for sediment control, but with the flexibility to include a third if required. Level 1 sediment control involves diverting clean water from disturbed areas while reducing sediment in runoff water in the active mining areas. Level 1 structures include in-pit sumps, clean water diversion and runoff collector ditches.

•

Level 2 sediment control involves collecting sediment-laden water in larger ponds to settle out the majority of suspended sediment.

•

Level 3 sediment control structures are tertiary ponds where flocculant is added to bind and settle clays under gravity, thus further clarifying water discharging from silt-trap ponds. Level 3 structures will be built only if adequate settling is not achieved in Level 2 ponds. Two Level 3 polishing ponds have been proposed at this time. All structures have limited capacity to store sediment so regular cleaning will be needed.

Progressive and on-going reclamation of mined out areas will lead to the potential for decommissioning of some sediment ponds and ditches and the potential for relocating collector ditches and freshwater diversions.

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Areas 212 and 85% of 217 will drain into the Sechoc II impoundment, while three additional Level 2 silt-trap ponds will contain runoff from remaining areas. Twelve runoff-collector ditches and one diversion channel will help to isolate flows into those impoundments. For the Sechoc dam a 3 m raise is considered necessary to meet freeboard requirements. At this height the dam would initially have a 3 m deep pond at the dam face and 3 m of freeboard. Side slopes of 2.5H:1V were determined to be stable for this height. The spillway will need to be relocated in rock at the west abutment. The spillway will be 10 m wide and the invert would be 3 m below the crest of the dam.

Areas 251 drains south to the Manto-4 valley. One runoff-collector ditch and one diversion channel are proposed. A 5 m high dam will be constructed to settle sediment. A tertiary pond is tentatively located immediately downstream of the silt-trap pond if needed. A second sediment pond is located beneath the proposed external waste stockpiles.

Sediment control structures will be decommissioned and reclaimed when mining in a particular area has ceased, the area has been re-vegetated and suspended material in runoff has returned to background levels.

18.9.3

Social and Environmental Assessment

18.9.3.1

Socioeconomic and Environmental Effects Assessment

Social and Ecosystem components for the Fenix Project were identified during the scoping and information collection processes. The processes comprised biophysical and socioeconomic baseline studies including reviews of literature, standards and regulations, and historical and current data, communications and consultations with local indigenous groups, regulators, and other stakeholders. The issues were then subject to a screening process, during which each component’s ‘valued’ status was rationalized using available information.

This screening process identified the ‘Valued’ Social and Ecosystem Components, known as VSCs and VECs respectively. These VSCs and VECs form the basis of the impact assessment. They are elements of a project’s socioeconomic and ecological surroundings that are of value and have importance to a number of interested and affected parties. The purpose of identifying these valued components is to focus the impact assessment on the most vital social and environmental issues.

The impact assessment identified VECs and VSCs likely to be affected by the project during the construction, operation and post-closure phases of the project. Potential effects were then subjected to mitigation measures and the resulting residual impacts were identified.

18.9.3.2

Social and Environmental Management Plans

Social and Environmental Management Plans have been developed to address potential residual effects to the key VSCs and VECs noted above and other pertinent issues. Table 18-24 summarizes key aspects of the management plans. Plans developed to date include:

•

Sediment Control and Stormwater Management Plan

•

Air Quality and Noise Impact Management Plan

•

Water Quality Management Plan

•

Community Development Plan

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•

Transportation Safety Management Plan

•

Hazardous Materials Handling and Storage Plan

•

Biodiversity Management Plan

•

Occupational Health and Safety

•

Waste Management Plan

•

Emergency Response and Spill Management Plan

Table 18-24: Summary of Social and Environmental Management Plans


Plan	Activity Description	Responsibility

Sediment Control and Stormwater Management	Sediment dams built as per engineering design	Construction Manager
	Ongoing water quality monitoring (including sample collection, analysis and reporting)	Environment, Health and Safety Manager
	Regular maintenance program to maintain hydraulic facilities to specification	Geotechnical Engineer
	Monitor and inspect facilities for structural integrity, capacity etc (including annual reporting)	Geotechnical Engineer
	Ongoing training, as required	Training Manager

Air Quality and Noise Impact Management Plan	Install systems as per engineering design	Construction Manager
Air Quality and Noise Impact Management Plan	Operate facilities according to manufacturer’s recommendations and air quality and noise impact management plan	Process Manager / Environment, Health and Safety Manager

	Regularly maintain facilities and equipment to ensure efficiency and noise suppression	Maintenance Manager / Environment, Health and Safety Manager

	Monitor stack gas quality to ensure compliance with permits	Environmental Health and Safety Manager

	Annual inspection of facilities to ensure integrity, or as needed	Process engineer

Water Quality Management Plan	Reduce contact with (clean) water into the mine development area	Construction Manager / Mine Manager
	Optimise process cooling water recycling and reduce fresh water demand	Process Design Engineer
	Manage water on-site until it is suitable for discharge	Operation Managers / Environment, Health and Safety Manager
	Manage the removal of riparian vegetation	Mine Manager

Community Development Plan	Continue development of small- scale, community- based education and health projects by Raxche’	Executive Director, Raxche Foundation
	Improve donor coordination and leverage resources of agencies such as USAID, EU and other multi-lateral and bilateral financial and aid organizations	Executive Director, Raxche Foundation
	Coordinate with local municipal government to facilitate efficient execution of infrastructure initiatives	Community Relations Manager / Executive Director, Raxche Foundation

Transportation Safety Management Plan	Training of drivers and operators in best practices and transport safety to minimize the risk of accidents	Training Manager
	Vehicle speed limits as per the transportation safety plan	Environment, Health and Safety Manager
	Inform new employees and consultants working on site of speed limits; post and repair signs indicating speed limits	Environment, Health and Safety Manager

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Plan	Activity Description	Responsibility
Hazardous Materials Handling and Storage Plan	Review, follow and update (as required) MSDS and hazardous materials handling and storage plan for PPE, storage and handling of hazardous materials	Environment, Health and Safety Manager/ Operation Managers

	Regularly monitor and inspect storage facilities	Environment, Health and Safety Manager

	Keep and maintain an up-to-date inventory list of hazardous materials stored on site as per standard operating procedures	Environment, Health and Safety Manager/ Operation Managers

	Ongoing training, as required	Training Manager

Biodiversity Management Plan	Sediment control and stormwater management structures built, operated and maintained as per engineering design and the Sediment Control and Stormwater Management Plan	Construction Manager / Operation Managers and Environment, Health and Safety Manager

	Clearance of aquatic and terrestrial vegetation and dredging activities as per the mine plan.	Mine Manager

	Monitor and report clearing and dredging activities to ensure conformance with mine plan	Environment Health and Safety Manager / Mine Manager

	Promote no hunting, no fishing and no harassment of wildlife on the mine site, process plant or in transit to the mine site by mine employees, and contractors; post signs stating mine policies on hunting, fishing and harassment	Mine Manager

	Inform new employees and consultants working on site of speed limits; post and replace signs indicating vehicle speed limits	Mine Manager

	Monitoring and reporting of human/ wildlife interactions.	Environment, Health and Safety Manager

	Reclamation as per the reclamation plan. Progressive reclamation on disturbed areas to reduce erosion and surface runoff	Environment, Health and Safety Manager

	Monitoring, inspection and reporting of reclamation progress and make adjustments to the reclamation plan as required	Environment, Health and Safety Manager

	Ongoing training, as required	Training Manager

Occupational Health and Safety Plan	Monitor and inspect workplace equipment and machinery to ensure that hazards to workers are eliminated or controlled, as per the OHSP	Environment, Health and Safety Manager

	Report unsafe acts or incidents and results of equipment and machinery inspections as they occur	Environment, Health and Safety Manager

	Keep an up-to-date inventory list of hazardous materials stored on site as per standard operating procedures	Environment, Health and Safety Manager/ Operation Managers

	Use and upgrade PPE and MSDS in accordance with hazardous materials handling and storage plan	Environment, Health and Safety Manager

	Ongoing training, as required	Training Manager

Waste Management Plan	Maintain waste segregation and recycling program	Mine Manager

	Incinerator and / or landfill operated as per design recommendations	Mine Manager

	Wastewater disposal system constructed to manage expected volume	Construction Manager

	Hazardous wastes managed on site	Mine Manager

	Meet water quality discharge criteria	Mine Manager

Emergency Response and Spill Contingency Plan	Emergency response planning	Environment, Health and Safety Manager
	Emergency response	Mine Manager
	Training and simulation	Training Manager

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18.10

Taxes

18.10.1

Taxes

The following Guatemalan taxes were applied to the Fenix Project on its revenue or asset values:

•

Income taxes (ISR) determined each fiscal year (the year ending December 31) on the basis of one of two regimes referred to as the “general income tax regime” (5% of gross revenues) and the “alternative income tax regime” (31% on taxable income). CGN has the right to elect early in December of each year which of the income tax regimes it wishes to have applied to its revenue in the following fiscal year. ISR instalments are payable monthly for the 5% general tax regime and quarterly in advance for the alternative income tax regime.

•

IETAAP (see definition below) is a tax only payable in a fiscal year when the “31% Alternative income tax regime” is in force. It is not payable when the 5% flat rate General tax regime is in effect. The IETAAP law is due to expire on 2007-12-31. For one time only, IETAAP is not payable for four fiscal quarters following “two consecutive years of operating losses (see definition below) for income tax purposes”. This one-time exemption from IETAAP for the Fenix Project will begin in January 2007, providing approval is received from the tax authority (SAT). IETAAP paid is creditable (i.e. not a deduction, but an offset) against income tax otherwise payable under the 31% Alternative income tax (but not under the 5% general regime) in respect of the same fiscal year and the next three fiscal years. There are two options for crediting IETAAP, as follows:

IETAAP paid during the four quarters of a fiscal year, can be credited to the Income Tax (quarterly or annual returns) determined in the following fiscal year. It can be credited for the following three fiscal years until it is exhausted. Any IETAAP payment remaining at the end of the third succeeding year is no longer creditable against alternative income tax otherwise payable, but the amount is a mandatory deduction in calculating the alternative income tax in the fourth succeeding fiscal year.

Since 2004, quarterly income tax payments (ISR) may be credited to IETAAP liability of the same fiscal year.

•

IETAAP is payable quarterly at the annual rate of 1% of the greater of net assets and gross revenue recorded on the Balance Sheet as of the prior fiscal year-end. If the net assets are greater than 4 times gross revenue, then the IETAAP is based on gross revenue.

•

Property tax (IUSI) is payable quarterly based on the registered value each quarter-end. It is determined on the basis of 0.9% of such values and is a deductible cost for alternative income tax purposes. At the commencement of construction of the Fenix Project, CGN has a market value of its assets for IUSI purposes of 44 million Quetzales (approximately US$5.8M). IUSI is treated as a site operating cost in the CFM. This tax should also be considered as a deduction in the IETAAP determination.

•

Value added tax (IVA) at 12% is charged on all purchased goods and services. IVA paid can be credited against remittances of IVA collected on sales and as export sales are zero-rated, IVA paid is refundable. Refunds of IVA cannot be applied for until CGN is established as an export company, that is, when it is in production and has export sales for at least one year. For the construction period, CGN has been granted exoneration of IVA and duties on imports of equipment and materials under Guatemalan Drawback Law (Decree No. 29-89) and will manage other purchases to minimize the IVA paid. It is assumed in this study that any IVA paid is refundable once CGN has export sales. In the CFM an estimate is made on the IVA recoverable per annum. For the same reason IVA and import duties were not included in the capital costs, as in the CFM these would be a ‘flow through’ cash flow stream, if fully recovered.

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18.10.2

Payments to Inco on Tonnes Mined (Production Interest), on Sales and on Net Smelter Returns (NSR)

As part of the agreement with Inco, Inco is to be paid a production interest, a sales agency commission, and a net smelter return, calculated as follows.

18.10.2.1

Production Interest

•

Payable on 70% of dry tonnes mined from Fenix.

•

Rate per tonne mined:

•

0 when average LME price is below US$ 2.50.

•

50% of the average LME price for a quarter when the average price is between US$ 2.50-7.00.

•

When the average LME for a quarter is above US$ 7.00, US$ 3.50 plus 25% of the amount by which the average exceeds US$ 7.00.

•

Payments are made quarterly, 60 days after the end of the quarter for which payment is being made.

•

Liability accrues from first quarter after quarter in which commencement of commercial production is achieved.

•

So if 300,000 tonnes are mined in a quarter and the average LME for the quarter is US$ 6.50, production interest of 300,000*70%*3.25 = $682,500 is payable 60 days after the quarter end.

18.10.2.2

Sales Agency Commission

•

Inco is appointed exclusive sales agent for all finished products, including ferro nickel.

•

Commission – 2.75% of net invoice value.

•

Net invoice value – net amount received for each shipment.

•

Payment – deducted by Inco from net invoice value paid to CGN.

18.10.2.3

Net Smelter Return

•

Payable on 70% of net smelter returns from the sale of ferro-nickel products.

•

Net smelter return is the amount receivable from Inco under the sales agency agreement less the Inco sales commission, costs of transportation and any royalties paid to the Government of Guatemala.

•

Payments are made quarterly 60 days after the end of the quarter for which payment is being made.

•

Liability accrues from the first quarter after quarter in which commencement of commercial production is achieved.

•

Rate of royalty is 2% of NSR when average LME is US$ 4.00 or more for the quarter.

•

After payback, there is an additional 4% royalty when the average price is more than US$4.00 and a further 2% when the average price is more than US$ 6.00.

•

Payback occurs when the cumulative operating cash flow exceeds the allowable costs accumulated from the Initial Closing which occurred on December 15, 2004.

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Allowable costs are all costs incurred to explore, develop, and construct the project, including sustaining capital, plus 8% of these costs to cover unallocable overhead and head office expenses.

18.10.3

Government Royalty

Pursuant to Guatemalan Mining Law a royalty of 1% of gross revenues is payable, 0.5% to the Guatemalan federal government and 0.5% to the municipal government.

In addition, CGN has agreed to pay an additional royalty to the Guatemalan government to resolve the question of the government’s ownership interest in CGN. The additional royalty prior to payback, defined as $1.7 billion in cumulative revenues, is 1.0% of gross revenues in quarters when average LME nickel prices exceed $7.00/lb. Following payback, the additional royalty will be paid as follows.

Table 18-25: Additional Royalty Schedule (Post-Payback)


Average Quarterly LME Price at Least ($/lb Ni)		Additional Royalty Rate
$7.00		1.5%
$10.00		2.0%
$13.00		2.5%
$16.00		3.0%
$19.00		3.5%
$22.00		4.0%

18.11

Capital and Operating Cost Estimates

18.11.1

Capital Cost Estimate

The capital cost estimates were developed by a team of engineers, designers, procurement specialists and cost estimators from Hatch (process plant, power plant and site utilities/services), Snowden (mine and mine infrastructure) and Skye (Owner’s Costs). The capital cost estimate, before owner’s costs, for Phase 1 of the project, as described within this study, is US$577 million in July 2007 US Dollars (US$640 million including Owner’s costs). The Phase 2 capital cost, before Owner’s costs, is US$319 million (US$344 million including Owner’s costs). The estimates are subject to certain qualifications, assumptions and exclusions, all of which are detailed below in Sections 18.11.1.4 and 18.11.1.5.

Table 18-26 below compares the current capital cost estimate with that of the feasibility study.

Table 18-26: Comparison of Feasibility Study and Updated Capital Costs


$ Millions		Feasibility Study	Update
DIRECT	Mine	15	19
	Power Supply	187	15
	Process Plant	177	230
Infrastructure and Auxiliaries		76	91
INDIRECT		157	157
OWNER’S COST		62	63
CONTINGENCY		80	65
TOTAL		754	640

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18.11.1.1

Estimate Terminology

The capital cost estimate consists of four main parts: direct costs, indirect costs, contingency and Owner’s Costs, as described below. Owner’s costs were estimated separately by Skye.

18.11.1.1.1

Direct Costs

Direct costs are the costs of all equipment and materials, together with construction and installation costs for all permanent facilities. This includes offsite infrastructure such as trans-shipment facilities and temporary upgrades to public roads (WBS Area 1000), the mine (WBS 1100), the process plant (WBS 1200 and 1300), site services (WBS 1400), the power plant (WBS 1500), coal storage and handling (WBS 1600), plant utilities (WBS 1700) and electrical distribution, process controls and communications (WBS 1800). The direct costs include the costs associated with the following:

•

Refurbishment of existing equipment and facilities.

•

Procurement and installation of new equipment.

•

Procurement, fabrication and installation of bulk materials.

•

Supplemental resources for equipment and bulk material installation on site, such as labour and construction equipment.

•

Site preparations (bulk earthworks) and the construction/refurbishment of all in plant roads and storm water ditching.

•

Procurement, fabrication, erection of buildings and associated services.

•

Procurement, fabrication, erection and/or refurbishment of all utilities and distribution systems.

18.11.1.1.2

Indirect Costs

Indirect costs include the following:

•

Temporary construction facilities including worker lodgings/services, secure lay-down areas, warehouses, etc.

•

Temporary construction services.

•

Concrete batch plant, and aggregate crushing plant.

•

Freight.

•

Vendor representatives.

•

Capital spares.

•

Engineering, procurement and construction management services (including travel expenses).

•

Third party engineering.

•

Pre-operational testing services including associated materials.

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18.11.1.1.3

Contingency

General

Contingency included in the capital cost estimate is an allowance for normal and expected items of work which have to be performed within the defined scope of work covered by the estimate, but which could not be explicitly foreseen or described at the time the estimate was completed. The contingency amount is an integral part of the cost estimate. It does not cover potential scope changes, price escalation, currency fluctuations, allowances for force majeure or other project risk factors or any of the other items that are excluded from the capital cost estimate (see Section 18.11.1.5). It should be assumed that the contingency amount will be spent.

Methodology

It is recognized that there is a degree of uncertainty in the estimates that make up the project’s cost. To model and analyse this uncertainty in the capital cost, a facilitated workshop was held with the project team and selected stakeholders (i.e. the Skye project management team, the independent engineer) attending. The collective knowledge and experience of the combined team was used to develop a three-point estimate to replace the single point estimates taken from the project estimate.

The three point estimate consists of:

•

The likely value, i.e. the single point estimate in the project estimate.

•

The maximum or pessimistic value.

•

The minimum or optimistic value.

The Palisade program @RISK was then used to run the simulation of the risk model that had been developed. @RISK uses the technique of Monte Carlo simulation to do a contingency analysis.

Briefly, using this technique, the computer generates a distribution of possible outcomes by recalculating the worksheet many times each time using different, randomly-selected values from the three-point estimate distributions selected for each input variable in the estimate.

The output from this analysis is a distribution of likely outcomes for the total project cost in a histogram that displayed the shape of the distribution including the maximum, minimum and mean values of the distribution. From the histogram it was possible to directly determine the amount of contingency required to achieve the desired probability of under-running a certain target.

The amount of contingency assigned to a project depends upon the confidence level that is assigned to the estimate. The 85^th percentile confidence level has been selected as the appropriate value in accordance with good industry practice (i.e. there is a 85% probability of not exceeding the estimated capital cost). Contingency amounts included in the capital cost estimate are shown in Table 18-27.

18.11.1.1.4

Owner’s Costs

For the Phase 1 project an amount of US$63 million was estimated by Skye for Owner’s project related costs and is included in WBS 2200 of this estimate.

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18.11.1.2 Capital Cost Summary

The project is to be constructed in two phases as described below:

Phase 1 (July 2007 – October 2009):

•

Refurbishment of the existing facility including, the process plant, upgrading the electric furnace and converting the existing kiln to coal-firing.

•

Installation of the new process equipment including the new dryer, ladle refinery, coal processing plant and the second kiln.

•

Installing the 230 kV transmission line (from the Tactic substation to the Fenix plant site) and 230 kV/34.5 substation (capital costs by others).

•

Upgrades of public roads connecting the port of Santo Tomas to the Fenix site (capital costs by others).

The facility is scheduled to begin operation in October 2009 with electric power provided via the 230 kV transmission line.

Phase 2 (Post October 2009):

Construction of the new 150 MW on-site power plant will begin after the process plant begins production. The new power plant is scheduled to be ready for operation by October 2014 and will include:

•

A new 150 MW circulating fluid bed (CFB) boiler. The boiler will be designed to operate using petroleum coke (petroleum coke).

•

A new 90 MW steam turbine generator (STG).

•

Refurbishment of the existing 60 MW STG.

•

New ancillary facilities including cooling water system, petroleum coke handling system, limestone handling system, etc.

The capital cost for the Phase 1 project, summarized by WBS, is provided in Table 18-27. The capital cost for the Phase 2 project is summarized in Table 18-27.

Table 18-27: Phase 1 Capital Cost Summary by WBS


WBS	Description	Total (US$)	Responsible
	Direct Costs
1000	Off-Site Infrastructure	16,276,684	Hatch/Hatch Mott MacDonald
1100	Mine and Mine Facilities	15,090,087	Snowden
1200	Ore Preparation (Drying and Calcining)	97,308,632	Hatch
1300	Smelting (Refining and Product Handling)	115,715,224	Hatch
1400	Site Preparation and Site Services	15,773,977	Hatch
1500	Power Plant (Emergency generators, package boiler)	3,376,703	Hatch
1600	Coal Storage and Handling	22,135,965
1700	Plant Utilities	23,460,420	Hatch
1800	Electrical Distribution, Process Control and Communications	45,155,089	Hatch
	TOTAL DIRECT COSTS	354,292,781

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WBS	Description	Total (US$)	Responsible
	Indirect Costs
2010	Construction Indirects
	Construction Temporary Facilities and Services	6,360,815	Hatch
	Construction Camp	10,120,000	Hatch
	Catering Services	5,563,800	Hatch
	Construction Power and Fuel	5,023,000	Hatch
	Special Construction Equipment (Heavy Cranage)	3,056,000	Hatch
	Construction Communication (Equipment and Setup)	852,500	Hatch
	Power Required for Pre-Operational Testing	1,818,750	Hatch
	Construction Training	297,600	Hatch
2030	Construction Consumables, Spares, Vendor Reps, etc.
	Capital Spares	5,150,000	Hatch
	First Fills (does not include consumables)	1,000,000	Hatch
	Freight	25,250,000	Hatch
	Vendor’s Representatives	4,200,000	Hatch
	Taxes and Duties	In Owner’s Costs (Refer to Section 18.10)
	Insurance	In Owner’s Costs	Skye
	Permits	In Owner’s Costs	Skye
	Third Party Consultants	1,200,000	Hatch
2080	EPCM
	Engineering	33,416,454	Hatch
	Construction	18,183,256	Hatch
	Project Management, Procurement and Support	11,547,143	Hatch
	Expenses	11,646,580	Hatch
	Basic Engineering	12,544,000	Hatch
	TOTAL INDIRECT COSTS	157,230,000
	TOTAL DIRECT + INDIRECT COSTS	511,500,000
1970	Contingency	65,500,000	Hatch
	Total Before Owner’s Costs	577,000,000
1980	Owner’s Costs	63,000,000	Skye
	TOTAL CAPITAL COSTS	640,000,000

Table 18-28: Phase 2 Capital Cost Summary

WBS	Description	Total (US$)	Responsible
2000	Direct Costs	222,215,000	Hatch
2900	Indirect Costs	62,600,000	Hatch
	TOTAL INDIRECT COSTS	284,815,000
2970	Contingency	33,800,000	Hatch
	Total Before Owner’s Costs	318,600,000
2980	Owner’s Costs	25,000,000	Skye
	TOTAL CAPITAL COSTS	343,600,000

18.11.1.3

Basis of Estimate

The Updated Feasibility Study estimate has been developed in accordance with the Association for the Advancement of Cost Engineering (AACE) international recommended practice No. 18R-97 “Cost Estimate Classification System – As Applied in Engineering, Procurement, and Construction for the Process Industries”.

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18.11.1.3.1

Project Definition

The Phase 1 capital cost estimate was prepared using project information contained in but not limited to the following documents:

•

Process flow diagrams.

•

Design basis reports.

•

General arrangement drawings.

•

Equipment list.

•

Preliminary engineering drawings and sketches.

•

Electrical single line diagrams.

•

Piping and instrumentation diagrams.

•

Topographic maps.

•

Existing documentation from the original (Exmibal) facility including flowsheets, general arrangement drawings, foundation designs, equipment lists, etc.

•

Geotechnical reports containing recommended geotechnical design bases for earthworks and foundations.

•

The Project Implementation Plan (PIP), including the project schedule. The PIP is based on an EPCM project execution methodology.

•

Vendor quotations for the design and supply of new equipment.

•

Vendor quotations for the refurbishment of existing equipment.

•

In-house data from similar projects.

•

Bulk material quantity take-offs.

•

Trip reports.

•

Freight/logistics quotes.

The labour component of the capital cost estimate is based on Hatch’s site visits and meetings with local and regional contractors who shared local labour practices, labour availability, labour rates and labour productivities in the general area of the site. No allowance has been included for labour rate increases attributable to future variations in the local and regional market conditions.

The Phase 2 cost estimate was developed using a data from similar sources and a similar methodology. The level of effort expended on the Phase 2 project was significantly lower than that for the Phase 1 project however. As a result, the level of accuracy of the Phase 2 capital cost estimate (-10%, +20%) is significantly lower than the level of accuracy for the Phase 1 capital cost estimate (-5%, +12%).

18.11.1.3.2

Estimate Basis - Summary

Table 18-29 summarizes the basis for estimating equipment prices, commodity quantities and unit rates.

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Table 18-29: Summary of Basis of Estimate


Item		Estimate Basis Description
Equipment

Major Equipment – New		Firm price quotations provided by vendors based on detailed specifications. Multiple vendors (where applicable) provided bids for each equipment package. Approximately 50% of the value of mechanical equipment is based on firm vendor quotations. (applies to Phase 1 project only)

Major Equipment – Refurbished		Budget price quotations provided by Vendors based on specifications and site inspections. In addition, information from a previous feasibility study commissioned by INCO (1993 study by DAVY) used as a reference.

Secondary Equipment – New		Costs developed from in-house databases, budget quotations and/or allowances based on similar projects and items.

Secondary Equipment – Refurbished		Refurbishment costs developed by a multidisciplinary team of Hatch engineers based on site inspections. Information from a previous feasibility study commissioned by INCO (1993 study by DAVY) used as a reference.

Bulk Materials & Site Works

Temporary Modifications to Public roads		An allowance of US$12.5 million has been included for the construction of an unpaved road and a new bridge, along a new (undefined) road alignment, parallel to road highway 7E, between Sumach and Fronteras (approximately 11 km). The new road is necessary for the transport of equipment and materials to the project site.

Site Preparation		Material Take-Offs (MTO’s) based on approved layouts and geotechnical investigation provided by Trow International Ltd.

Concrete		MTO’s based on approved layouts, design sketches, calculations and data from similar projects.

Structural Steel		MTO’s based on approved layouts, design sketches, calculations and data from similar projects.

Architectural		MTO’s based on approved layouts, design sketches, calculations and data from similar projects.

Site Services, Piping & Valves		• MTO’s based on preliminary piping material and valve specifications, approved layouts, design sketches, calculations and data for similar projects. • Minor utility piping (<65 mm) factored from equipment cost.

Process Piping & Valves		MTO’s based on preliminary piping material and valve specifications, approved layouts, design sketches, calculations and data for similar projects.

Electrical		MTO’s based on electrical design criteria, mechanical equipment list, layouts, single line diagrams (SLD), process flow diagrams (PFD) and data from similar projects.

Process Control and Instrumentation		MTO’s based on process control and instrumentation design criteria, mechanical equipment list, layouts, process and instrumentation diagrams (P&ID’s), process flow diagrams (PFD), data from similar projects and the assumptions listed in 18.2.4.1.3.

Installation

Installation Labour Productivities and Unit Costs		Installation manhours estimated based on Gulf Coast productivities. Local labour productivity multiplier and all-inclusive hourly costs developed based on information provided by local/regional construction contractors on a discipline-by-discipline basis.

Construction Equipment		Unit costs based on data obtained from local/regional construction companies

Vendor Reps/Supervision		Derived from vendor quotations for major equipment and in-house data from similar projects.

Contractor Overheads		Based on information provided by local construction contractors on a discipline-by-discipline basis.

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Item		Estimate Basis Description
Freight

Heavy Lift Items		A list of oversized/overweight items developed by Hatch. Costs to ship these items to site assuming in-land transport to site via truck along public highways. Costs based on transport studies and budget quotes by Kuehne+Nagel and Panalpina (freight forwarding companies).

Remaining Items		An estimate of freight requirements (tonnage and volume) developed by Hatch. Costs to ship these items to site assuming in-land transport to site by truck from Santo Tomas. Costs based on transport studies and budget quotes by Kuehne+Nagel and Panalpina (freight forwarding companies).

Critical Spare Parts		Percentage of equipment costs

EPCM		Estimated by Hatch as part of the Project Implementation Plan

Taxes and Import Duties		VAT and additional taxes applicable to foreign labour are excluded. Only applicable payroll taxes are included in the construction labour rates. Import duties are excluded.

Owner’s Costs		By Skye

18.11.1.4

Qualifications, Assumptions and Exclusions

18.11.1.4.1

Qualifications and Assumptions

General

Readers should be aware that any items excluded from the capital cost estimate (refer to section 18.11.1.5) have the potential to materially impact the actual costs of the project.

Some costs that have been excluded from the CAPEX are included in the economic analysis (e.g. working capital, sustaining capital, warehouse inventories, etc. – refer to Section 18.12).

Pricing

Firm purchase agreements are in place for approximately 5% of the major equipment. In addition, firm price vendor bids have been received for a total of $82 million in expenditures which represents 50% of the equipment purchases. None of the pricing for commodities or the design/supply of the remaining 50% of the equipment is based on binding quotations.

Project Currency, Estimate Base Date & Foreign Exchange

•

All costs are expressed in July 2007 US dollars.

•

Costs are based on July 2007 market conditions with no provision carried in the estimate for escalation beyond this date.

•

Costs submitted in other currencies have been converted to United States dollars (US$). Foreign currency exchange rates applied to the capital cost estimate relative to the US$ are set out in Table 18-30.

•

No provision has been made for variations in the currency exchange rates from those indicated in Table 18-30.

•

No provision has been made for any taxes or fees applicable to currency exchanges.

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Table 18-30: Exchange Rates


Country	Currency	Code	Per US$
Canada	Dollar	CAD$	1.06
USA	Dollar	US$	1.0
European Union	Euro		0.73
Guatemala	Quetzal	QTZ	7.61

Taxes

All tax related issues, such as VAT and additional taxes on foreign labour, are excluded from the capital cost estimate but included in the economic analysis. Local construction labour rates include applicable payroll taxes.

Accuracy

The Phase 1 capital cost estimate, including contingency, for the mine, process plant and infrastructure has been prepared to a level of definition appropriate for an intended level of accuracy of -5%, +12%.

The Phase 2 capital cost estimate, including contingency, for power plant has been prepared to a level of definition appropriate for an intended level of accuracy of -10%, +20%.

Project Implementation Plan

The capital cost estimate is based on the assumption that Skye will follow the project implementation plan described in the Hatch Project Implementation Plan. Any failure to follow this plan (e.g., use of an EPC or LSTK implementation strategy instead of an EPCM implementation strategy) may have a material impact on both project schedule and costs.

18.11.1.5

Exclusions

The following items are excluded from the capital cost estimate:

•

All costs associated with the design, supply and installation of the 230 kV transmission line including any upgrades to the Tactic substation of Guatemalan national grid.

•

All costs associated with the permanent upgrading of public roads (including pavement structures, road alignment and bridges) between Santo Tomas and the Fenix plant site (CA9, CA13 and 7E) to meet plant operations requirements beyond the $12.5 million allowance included in the CAPEX.

•

Escalation beyond the estimate base date, including escalation due to volatility in local market conditions.

•

All impacts of foreign currency exchange rate variations.

•

Costs associated with the construction of staff housing facilities

•

Allowances for any changes to the scope of the project (as described in Sections 18.1, 18.2, 18.3, 18.4, 18.5 and 18.6).

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•

Allowances for either (a) general project risks that could affect any project such as this project (e.g., variations in market conditions that could affect equipment, commodities and/or labour costs, adverse weather conditions, labour unrest, disputes with local residents including local indigenous groups, geotechnical or process related design issues, delays due to the late receipt of equipment or materials, poor performance by contractors, force majeure or acts of God), or (b) risks that are specific to this project (as described in Section 18.14.2).

•

Allowance for the risks associated with the Guatemalan political, legal or regulatory environment, including (a) the risk of changes to any laws, regulations, rules or policies in Guatemala, or the governmental or judicial interpretation thereof, (b) the risk of Skye or CGN failing to comply with any such laws, regulations, rules or policies and the costs of any resulting penalties, fines, suits, etc., and (c) the risk of Skye or CGN not being able to obtain or maintain any permits, licenses and other authorizations required for the project.

•

Allowance for the risks associated with CGN’s acquisition and maintenance of (a) the title to the Fenix property, and (b) its current or future rights to extract, process and sell minerals from the Fenix property.

•

Any costs incurred to accelerate the work or to get the work back on schedule if it falls behind schedule (e.g., overtime charges, expediting charges, etc.).

•

Costs associated with royalties (see Section 18.10) and property taxes (these are included in the economic analysis).

•

Costs associated with warehouse inventory over and above critical spare parts (these are included in the economic analysis).

•

Working capital (including first fills of reagents and other consumables) and ongoing/sustaining capital (included in the economic analysis, refer to Sections 18.12.7 and 18.12.8).

•

Research and exploration drilling.

•

Any incentive/bonus schemes.

•

Costs incurred in connection with the Project prior to January 2007 (e.g., site acquisition costs, costs associated with the preparation of the Feasibility Study and any prior studies, licensing and royalty charges already incurred, etc.).

•

Any financing charges or costs, whether related to debt or equity financing, including any interest charges.

•

Sustaining capital (included in the economic analysis).

•

Salvage value of plant equipment.

•

Facility closure costs (included in the economic analysis).

•

Cost associated with the disposal of demolished equipment and material. All demolished items will be stored at a location within the plant battery limits.

•

Any additional work that is required as a result of conditions, both subsurface conditions and conditions in and around the project site, that were not known as of the base date of the estimate (e.g., latent conditions in existing facilities and unknown geotechnical conditions), including any costs incurred in establishing and confirming as-built information over and above that defined in the estimate.

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18.11.2

Operating Cost Estimate

18.11.2.1

Introduction

Operating costs were calculated for each of the 30 years of operation of the RKEF plant for the Fenix Project. The estimate covers all costs normally expected in the ordinary course of operations for a project such as this, including process and power plant operations, mining activities, as well as general and administrative operating costs. The estimate was prepared in US Dollars (US$). Costs associated to thermal coal, petroleum coke and oil products have been priced based on long-term price forecasts provided by Skye. Transportation costs for consumables were based on road transportation by truck from the port of Santo Tomas to the site. Other consumables unit costs and prices were obtained between October 2005 and August 2006. Labour rates were updated as per Skye’s revision on their salary and wage levels. An exchange rate of US$1.00 = Q$7.61 was applied where required.

Table 18-31 summarizes the feasibility study and updated operating costs.

Table 18-31 Summary of Operating Costs for Feasibility Study and Update


$/lb	Feasibility Study		Update Years 6-20		Update Years 1-5
Mining	0.28		0.31		0.24
Processing	0.86		1.07		0.87
Power	0.40		0.53		1.76
Shipping/G&A	0.55		0.63		0.80
Iron Credits	(0.20	)	(0.20	)	(0.20	)
Total	1.89		2.34		3.47

No allowance has been made for variations in exchange rates or escalation after the base date in the costs of any inputs (other than thermal coal, petroleum coke and oil products, which have been priced based on long-term price forecasts provided by Skye), including unit costs of consumables or labour rates.

The intended level of accuracy of the operating cost estimate, based on the unit costs assumed for the Project and current at the time of the study, is +/-10%. No contingency was included in the operating cost estimate.

Where applicable, the Guatemalan value added tax (IVA of 12%) was included.

18.11.2.2

Scope of Estimate

Operating costs were estimated starting in Year 1, in which the second kiln is scheduled to begin operation, 3 months after the start-up of the plant. For the first five years of operation a 230 kV connection to the Guatemalan power grid is expected to supply power based on a 5-year Purchase Power Agreement (PPA) being negotiated between Skye and an IPP. .

In Year 6, the new CFB boiler is scheduled to begin operation with the grid connection as backup during boiler shutdown. From this point on, power will be generated from the captive power plant for the life of the project.

Operating costs were estimated to Year 30 based on this power plant configuration and the mine plan developed by Snowden.

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Table 18-32 below summarizes the scope of the estimate by area. Note that freight insurance, marketing fees, and credits are not included in this estimate. These items are included in the Financial Analysis (Section 18.12) to follow Brook Hunt’s definitions for C1 - Net Direct Cash Cost. See Section 18.11.2.7 for other exclusions in the operating cost estimate as described in this section.

Table 18-32: Scope of Operating Cost Estimate by Area


Area Description		Consumable/Service
Mine		Diesel Fuel
		Consumables (tires, lubricants, oils)
		Maintenance
		Other
		Labour
Ore Preparation		Power
Drying		Power
		Coal
		Heavy Fuel Oil
Calcining		Power
		Coal
		Oxygen
		Heavy Fuel Oil
Agglomeration		Power
		Flocculant
Smelting		Power
		Consumables (Electrode Paste, Electrode Casings and
		Refractories)
		Miscellaneous
Slag Granulation		Power
Refining		Power
		Reagents
		Refractory
		Miscellaneous
Metal Recovery		Power
Shotting		Power
Product Shot Drying		Power
		Diesel
Coal Handling (Crushing / Milling)		Power
		Nitrogen
		Diesel
Product Shipping		Product Bags
		Truck/barge transportation to port
		Ocean Freight
General & Administrative		General Power
		Miscellaneous supplies
		Legal services
		MIS & Telecommunications
		Corporate Office (Guatemala City) Expenses
		Insurance
Yard Services		Power
		Chemicals
		Diesel

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Area Description		Consumable/Service
General		Site Electrical Losses
		Transmission Losses
		Allowance for Process Plant Mobile Equipment
		Maintenance Materials
		Environmental Program
		Personal Protection
		Travel Expenses
		Employee Transport
		Public Relations
		Transmission Tolling Agreement (TTA)
		Contracted Services

Labour		Senior Management
		General Administration
		Process Plant Operations
		Technical Services, Engineering & Maintenance

The cost of electrical power for the first 5 years of Fenix operation is based on the 5-year PPA being negotiated between CGN and an IPP. Such energy cost includes a forecast of heavy fuel oil prices which are a determinant of the energy cost under the PPA. Beyond Year 5, power plant operating costs were estimated separately and included in the equivalent unit power cost applied in the process plant. The scope of this estimate includes fuels, labour, maintenance materials, and an allowance for power purchased from the grid, among others. The Financial Analysis assumes, however, that CGN makes a lump-sum payment in Year 6 of $25 million to terminate its payment obligations under the TTA (as it has a right to do under the TTA), and accordingly no further TTA payments are included in the Financial Analysis after Year 6.

18.11.2.3

Annual Operating Costs over 30 Years

Figure 18-9 and Table 18-33 below present the operating costs from Year 1 to Year 30.

LOGO

Figure 18-9: Project Operating Costs and Ore Grade – Years 1-30

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Table 18-33: Project Operating Costs – Years 1-30


Year	Ore Grade (Ni%)	OPEX without VAT
Year	Ore Grade (Ni%)	Total Cost (k$US)	US$/ lb Ni	US$/dry tonne ore processed
1	2.08	109,543	4.53	188.44
2	1.87	158,126	3.57	132.00
3	1.95	179,317	3.20	124.05
4	2.00	179,665	3.08	122.94
5	1.68	177,316	3.68	122.13
6	1.88	114,163	2.09	78.30
7	1.73	119,027	2.37	81.63
8	1.61	109,920	2.37	75.72
9	1.67	103,050	2.56	84.94
10	1.68	111,927	2.31	76.94
11	1.62	120,945	2.59	83.17
12	1.69	111,445	2.28	76.53
13	1.73	116,074	2.32	80.19
14	1.62	112,164	2.39	77.09
15	1.76	115,948	2.27	79.61
16	1.85	91,995	2.57	95.09
17	1.69	115,251	2.36	79.25
18	1.69	110,850	2.27	76.19
19	1.65	113,529	2.38	78.12
20	1.70	110,575	2.25	76.16
21	1.69	112,280	2.30	77.08
22	1.58	107,923	2.38	74.20
23	1.61	115,788	2.49	79.30
24	1.30	105,537	2.90	72.79
25	1.36	99,566	3.09	81.85
26	1.34	107,530	2.83	73.52
27	1.30	118,661	3.22	80.80
28	1.29	105,904	2.91	72.31
29	1.28	108,305	3.00	74.11
30	1.27	97,740	3.24	79.54

18.11.2.4

Operating Costs for Year 4

Year 4 represents the first full production year for the RKEF plant with both kilns and the electric furnace operating at full capacity with power sourced from the Guatemalan power grid, based on the 5-year PPA, being negotiated.

18.11.2.4.1

Operating Cost by Area

Table 18-34 outlines the operating cost breakdown by area for Year 4 for total project costs.

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Table 18-34: Project Operating Cost Breakdown by Area – Year 4


Area	Total Cost (kUS$)	US$/lb Ni	US$/dry tonne ore
Mine (incl. Mine Labour)	11,473	0.20	7.85
Ore Receiving & Preparation	341	0.01	0.23
Drying	6,493	0.11	4.44
Calcining	25,223	0.43	17.26
Agglomeration	939	0.02	0.64
Smelting	80,084	1.37	54.80
Slag Granulation	985	0.02	0.67
Refining	7,860	0.13	5.38
Metal Recovery	25	0.00	0.02
Shotting & Product Handling	132	0.00	0.09
Coal Handling	1,965	0.03	1.34
Product Shipping	11,017	0.19	7.54
Accounting & Administrative	6,496	0.11	4.45
Yard Services	1,632	0.03	1.12
General	18,784	0.32	12.85
Labour	18,620	0.32	12.74
Total	192,069	3.30	131.43

Figure 18-10 presents the cash cost distribution by area for Year 4.

LOGO

Figure 18-10: Operating Cost Distribution by Area – Year 4

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Smelting costs as a proportion of total operating costs in year 4 are significantly higher than was the case in the October 2006 feasibility study primarily because the majority of the project’s power consumption is in smelting, and power costs in year 4 are significantly higher as a result of power supply under the PPA.

18.11.2.4.2

Operating Cost by Element

Table 18-35 outlines the project operating cost breakdown by element for Year 4. The main elements of the project operating costs are:

•

Power.

•

Coal.

•

Labour.

•

Mining, including material movement, ore haulage and mine services.

•

Heavy Fuel Oil.

•

Diesel.

•

Smelting and refining consumables.

•

Maintenance and administrative materials.

•

Shipping and product handling.

•

General and contracted services.

Table 18-35: Project Operating Costs by Element – Year 4


Element	Total Cost (kUS$)	US$/lb Ni	US$/dry tonne ore
Power	93,275	1.60	63.83
Minesite Consumables	1,282	0.02	0.88
Coal	24,595	0.42	16.83
Heavy Fuel Oil	236	0.00	0.16
Diesel (including Mine)	3,041	0.05	2.08
Agglomeration consumables	11	0.00	0.01
Smelting consumables	1,922	0.03	1.32
Refining consumables	5,183	0.09	3.55
Nitrogen	564	0.01	0.39
Oxygen	1,253	0.02	0.86
Product Handling and Shipping	11,017	0.19	7.54
General and Administrative	17,370	0.30	11.89
Maintenance Materials (including Mine)	7,932	0.14	5.43
Water treatment chemicals	120	0.00	0.08
Labour (including Mine)	24,269	0.42	16.61
TOTAL	192,069	3.30	131.43

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Figure 18-11 presents the cost distribution by element for Year 4, based on cash costs.

LOGO

Figure 18-11: Operating Cost Distribution by Element – Year 4

18.11.2.5

Operating Costs for Year 6

Year 6 represents the first full production year for the RKEF plant with the new 150-MW CFB boiler operating.

18.11.2.5.1

Operating Cost by Area

Table 18-34 outlines the operating cost breakdown by area for Year 6 for total project costs.

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Table 18-36: Project Operating Cost Breakdown by Area – Year 6


Area	Total Cost (kUS$)	US$/lb Ni	US$/dry tonne ore
Mine (incl. Mine Labour)	12,785	0.23	8.77
Ore Receiving & Preparation	89	0.00	0.06
Drying	5,306	0.10	3.64
Calcining	21,857	0.40	14.99
Agglomeration	254	0.00	0.17
Smelting	22,431	0.41	15.39
Slag Granulation	258	0.00	0.18
Refining	6,003	0.11	4.12
Metal Recovery	7	0.00	0.00
Shotting & Product Handling	91	0.00	0.06
Coal Handling	1,133	0.02	0.78
Product Shipping	10,347	0.19	7.10
Accounting & Administrative	5,934	0.11	4.07
Yard Services	911	0.02	0.62
General	11,809	0.22	8.10
Labour	17,533	0.32	12.03
TOTAL	116,747	2.13	80.08

As mentioned earlier, power plant costs were calculated separately to generate unit costs for electricity, which have then been included in the total cost for each of the areas. Total power plant cost for Year 6 is estimated to be approximately US$24.4 million, equivalent to US$0.45 per pound of nickel and US$27.93 per MWh.

Figure 18-10 presents the cash cost distribution by area for Year 6.

LOGO

Figure 18-12: Operating Cost Distribution by Area – Year 6

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18.11.2.5.2

Operating Cost by Element

Table 18-35 outlines the project operating cost breakdown by element for Year 6. The main elements of the project operating costs are:

•

Power

•

Coal

•

Labour

•

Mining (including material movement, ore haulage and mine services)

•

Heavy Fuel Oil

•

Diesel

•

Smelting and Refining Consumables

•

Maintenance and Administrative Materials

•

Shipping and Product Handling

•

General and Contracted Services

Table 18-37: Project Operating Costs by Element – Year 6


Element	Total Cost (kUS$)	US$/lb Ni	US$/dry tonne ore
Power	24,393	0.45	16.73
Minesite Consumables	1,743	0.03	1.20
Coal	24,562	0.45	16.85
Heavy Fuel Oil	236	0.00	0.16
Diesel (including Mine)	3,337	0.06	2.29
Agglomeration consumables	11	0.00	0.01
Smelting consumables	1,922	0.04	1.32
Refining consumables	4,833	0.09	3.31
Nitrogen	560	0.01	0.38
Oxygen	1,219	0.02	0.84
Product Handling and Shipping	10,347	0.19	7.10
General and Administrative	11,291	0.21	7.74
Maintenance Materials (including Mine)	8,762	0.16	6.01
Water treatment chemicals	120	0.00	0.08
Labour (including Mine)	23,411	0.43	16.06
Total	116,747	2.13	80.08

Figure 18-11 presents the cost distribution by element for Year 6, based on cash costs.

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LOGO

Figure 18-13: Operating Cost Distribution by Element – Year 6⁶

18.11.2.6

Basis of Cost Estimate

Based on Year 4 and 6, the most important operating costs are:

•

Power

•

Coal

•

Labour Cost

•

General and Administrative Costs

•

Product Shipping

General operating cost includes an allowance for miscellaneous supplies and services. Between year 1 and the end of the first quarter of year 6, a monthly transmission line charge of US$603,100 (excluding VAT) is expected to be paid based on the PPA being negotiated with an IPP. This cost has been included under this category and does not include any potential recoveries from the incorporation of the Tactic-El Estor line into the Guatemalan primary transmission system.

A review of the estimate basis for the most important operating costs is presented below. Mine operating costs were developed by Snowden.

⁶	Power cost includes fuels (petroleum coke, HFO, diesel), maintenance materials, miscellaneous consumables , labour and allowance for power purchase from the Guatemalan Power Grid.

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18.11.2.6.1

Power

For the first 5 years of operation, power is expected to be supplied by 230kV transmission line connected to the Guatemalan power grid. A 5-year PPA being negotiated between Skye and an IPP defines the power cost for such period of time. Such agreement is based on power consumption. A 5 MW reciprocating engine will serve for emergency power.

Power to the Fenix Project will be supplied by a captive, on-site power plant expected to begin operation in Year 6. A new 150 MW (gross) boiler fired with petroleum coke will be commissioned and the Guatemalan power grid will be used as backup when the CFB boiler is down.

For the purpose of operating cost estimation and financial analysis of the Project, power will be ‘sold’ at cost to the Project with the operation of the 150MW CFB boiler. Operating costs associated with the power plant are calculated separately and the equivalent electricity cost is developed for each year. Table 18-38 shows power demand/production and the equivalent cost of electricity for each operating year.

Table 18-38: Power Unit Cost (per MWh)


Year	Power Consumption (MWh)	Without Tax
1	383,065	$	109.79
2	707,497	$	100.33
3	839,509	$	98.63
4	846,329	$	98.40
5	843,861	$	98.40
6	845,382	$	27.93
7	844,259	$	32.07
8	843,335	$	27.94
9	715,690	$	32.81
10	843,843	$	27.93
11	843,426	$	32.07
12	843,966	$	27.93
13	844,215	$	32.07
14	846,406	$	27.94
15	844,455	$	32.07
16	588,492	$	29.73
17	843,952	$	32.07
18	843,912	$	27.93
19	843,644	$	32.07
20	843,994	$	27.93
21	843,947	$	32.07
22	843,099	$	27.94
23	843,402	$	32.07
24	840,978	$	27.95
25	713,805	$	32.82
26	841,380	$	27.95
27	841,085	$	32.08
28	840,991	$	27.95
29	840,912	$	32.08
30	840,862	$	28.65

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Power plant operating costs between Years 6 and 30 are broken down into the following:

•

Fuels: petroleum coke (petroleum coke), heavy fuel oil, and diesel.

•

Other consumables (limestone, sand, plant chemicals, lubricants, etc.).

•

Emissions testing, and permits and certificates.

•

Maintenance cost (including labour).

•

Operating and administrative labour.

•

Other Costs (waste disposal and supplies).

•

Allowance for power purchased on the Guatemalan electricity spot market during boiler shutdown (estimated to be US$100/MWh).

Table 18-39 shows power cost structure for Year 6 of operation.

Table 18-39: Power Cost Breakdown for Year 6


	Consumption tpy	Without Taxes $ /tonne	Total $000’s
Fuels*
Petroleum coke	304,669	58.8	17,926
Limestone	122,252	4.65	568
Sand	6,401	3.50	22
HFO	2,033	265.9	541
Light Oil	566	495.9	281
Other Consumables			317
Emission Testing, Permits and Certificates			94
Maintenance Materials and Labour			2,174
Labour			1,101
Waste Disposal			7
Staff Material Costs			112
Power Purchased from the Grid	4,642 MWh		464
Total Power Plant Operating Cost			23,608
Equivalent Cost of Electricity (US$/MWh)			27.93

*	Long term prices were applied.

18.11.2.6.2

Coal

For the October 2006 Feasibility Study, it was assumed that coal is imported from Colombia. A long term FOB load port unit cost of coal was determined based on the Pace Global’s “Coal and Petcoke Strategy Supply Plan Report” – July 11, 2006 developed for Skye. An updated report was prepared by PACE to reflect July 2007 prices. Ocean freight, local port and transportation charges were added based on the estimates prepared by Trans Mar and updated to July 2007 prices. Costs from Santo Tomas to the Site were based on local quotes ($US 0.10/t - km) for road transportation by truck. The Guatemalan VAT was estimated for port and transportation services incurred in Guatemala.

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Table 18-40 outlines how ocean and local freight costs, port and other charges were applied to determine the unit cost of coal for the Project.

Table 18-40: Coal Cost Breakdown


Coal Cost Items	US$/tonne
Coal Unit Cost (FOB Colombia)	$	44.76
Ocean Freight	$	15.15
Local Port Charges	$	6.38
Truck Transfer	$	2.25
Sto. Tomas Trans-shipment Terminal	$	2.41
Road Transportation	$	15.50
TOTAL COST WITH TAXES	$	86.45
Recoverable VAT	$	2.34

A coal net heating value (as fired) of 26.3 MJ/kg was used as reported for typical Colombian coal, with a fixed carbon and sulphur content of 47 wt% and 0.67 wt% on a dry basis, respectively.

Coal is used in the drying and calcining areas in two forms: pulverized and crushed. Pulverized coal is used to fire the rotary dryer and rotary kilns. Crushed coal is added to the rotary kilns to act as a reductant.

Coal consumption is calculated from the METSIM^® mass and energy balances for each year. The expected coal consumption for Year 4 is presented below.


Coal User		Annual Consumption (dry t/y)
Pulverized coal to the dryer		54,602
Pulverized coal to the kilns:
- Kiln #1		75,734
- Kiln #2		92,667
Crushed coal to the kilns:
- Kiln #1		23,545
- Kiln #2		29,420

18.11.2.6.3

Labour Cost

Labour costs were estimated based on the staffing requirements and a salary structure provided by Skye (updated in July 2007). Table 18-41 shows the estimated labour cost for Year 4.

Table 18-41: Labour Cost for Year 4


Area	Total Positions	Total Cost (US$)
Senior Management	5	1,676,654
Administrative Staff	86	2,683,659
Plant Operations	356	7,246,142
Mine Operation	206	5,648,422
Engineering & Technical	265	7,014,031
TOTAL	918	24,268,908

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The total headcount for Year 4 is 918 excluding miscellaneous contracted services provided by third party suppliers. The total headcount increases by 50 in Year 6, including staff for the power plant. The power plant labour cost is included in the equivalent electricity cost beginning in Year 6. Labour related to transport operations is included in the transportation unit cost for coal and petroleum coke. Costs related to service staff (catering, janitorial, etc.) is included in the Contracted Services allowance under General costs.

18.11.2.6.4

Product Shipping

Dried ferro-nickel product will be packaged at the process plant in 2-tonne bulk bags and will be loaded onto returning trailers used for petroleum coke/coal. Containers will be transported by truck to the Guatemalan port to be sold to customers CIF port of import. Table 18-42 shows shipping and transportation costs supplied by Trans Mar. Ocean freight costs are based on equal shipments to North America, Europe and the Pacific Rim.

Total freight costs to be assumed by CGN are as follows:

Table 18-42: Shipping Cost Breakdown (US$/tonne)


Per Tonne	Asia		North America		Europe
Transportation costs to local Port*	$	48.51	$	21.01	$	21.01
Ocean Freight	$	117.96	$	112.55	$	96.56
TOTAL to CGN	$	166.47	$	133.26	$	117.57

*	Includes VAT

Therefore, the average shipping costs included in the project operating cost estimate are US$139.20 per tonne.

18.11.2.7

Exclusions

The following items are excluded from this project operating cost estimate:

•

Marketing and royalties (included in the economic analysis).

•

Corporate overhead incurred by Skye or Compañia Guatemalteca de Níquel (included in the Financial Analysis).

•

Research and exploration.

•

Any financing charges or costs, whether related to debt or equity financing, including any interest charges.

•

Depreciation and amortization (included in the economic analysis).

•

Sustaining capital (included in the economic analysis).

•

Corporate taxes (included in the economic analysis).

•

Any costs incurred outside of the ordinary course of business (as contemplated by this Study), including (a) costs related to expansions, major upgrades or efficiency improvement projects, (b) costs related to industrial incidents, labour unrest, protests by local residents or any events normally considered force majeure events or acts of God, and (c) any costs or charges incurred in connection with litigation matters or other legal proceedings.

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In general, all of the risks that are identified in capital cost estimate in Section 18.4.2 and 18.11.1.5 (e.g., legal, political and regulatory risks, risks associated to property title and permitting, etc.) are also risks that are applicable to the operating cost estimate and, if they occur, could materially increase operating costs.

The operating cost estimate is based on unit costs and prices that were obtained between October 2005 and August 2006 and updated in July 2007. Long-term price forecasts were used for the major consumables – thermal coal, petroleum coke and oil products. Escalation in such costs or prices after the date on which they were obtained, together with foreign exchange rate variations that may impact such costs and prices, is not taken into account in the estimate. The economic analysis was carried out assuming that long term, real prices of operating cost inputs and of the nickel product are constant.

18.12

Economic Analysis

18.12.1

Overview

The economic analysis has been developed based on the capital cost and operating cost estimates set out in Sections 18.11.1 and 18.11.2 and is therefore generally subject to the same qualifications, assumptions and exclusions. For example, the occurrence of any of the risks that have been excluded from the capital and operating cost estimates would likely have a material impact on the accuracy of the economic analysis set out in Section 18.12.

In addition, the capital cost and operating cost estimates have been prepared as of base dates with no allowance for escalation in prices or for variations in currency exchange rates, unless expressly stated. Obviously, there will be some variations in currency exchange rates and there is likely to be escalation in certain prices, and both will have some impact on the accuracy of the economic analysis set out in this Section 18.12. Certain key capital cost assumptions are referenced in Section 18.12.5 and certain key assumptions respecting revenues and taxation have also been made and are detailed in Sections 18.12.10 and 18.10, respectively.

The findings of the economic analysis completed by Hatch as part of the Feasibility Study are summarized in Table 18-43. The economic analysis of the Fenix Project was completed using an after tax cash flow model (CFM) in which the production, revenues and cost performance of the Project were modeled over a 31-year life of mine.

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Table 18-43: Fenix Project Base Case Economic Analysis Summary


Production / Economic Parameter	Units	Production Years 1 - 30 US$/lb Ni
Period of Operations	years	30
Total Ore Feed (dry)	kt	41,348
Total FeNi production (dry)	kt	1,716
Total Nickel production	Mlbs	1,332
Total Site Operating Cost	US$/lb Ni	2.49
Net Direct Cash Cost, including by-product credits (C1)	US$/lb Ni	2.66
Base Case Nickel Price	US$/lb Ni	6.50
Net After Tax Cash Flow	US$M	2,743
(Year 3 to Year 22)	US$/lb Ni	2.83
Net After Tax Cash Flow	US$M	3,550
(Year 1 to Year 31)	US$/lb Ni	2.67
Net Present Value @ 6 %	US$M	797.2
@ 8 %	US$M	490.6
@ 10 %	US$M	275.0
Internal Rate of Return (IRR)	%	14.3
Pay back in Production Year	years	7.3

18.12.2

Cash Flow Model Basis

The key model assumptions made in developing the CFM are presented in Table 18-44.

Table 18-44: Key Cash Flow Model Assumptions


Model Parameter	Assumption	Note
Discounted Cash Flow (DCF) total time frame	32 Years	The DCF time frame covers the construction period (2 years) and the identified life of mineral reserves (30 years), as defined by the mining plan and schedule

DCF Time Period	1 year	A single year is used as the DCF time period in which all production, costs and revenues are assumed as average for the year. All cash flows applied to a particular year are assumed to occur at the end of that time period.

Cash Flow Model start year	Year –2	The start year in the CFM is year –2, i.e. 2 years before production operations start in year 1. It was assumed that year –2 begins on 1 July 2007.

Real Date for Costs and Revenue	Project base date, 1 July 2007	All cost / revenue inputs are assumed as real at this date. The CFM analysis is in real July 2007 terms for all time periods considered.

Inflation	No inflation factors applied	US$ values are assumed constant over the 30 years, with no inflation applied.

Exchange Rates	CAD$ 1.06 /US$ GTQ7.61 / US $	Where applicable fixed exchange rates of CAD$ 1.06 / US$ and GTQ7.61/US $ are used

Cash Flow Model Currency	US$	All foreign currencies are converted to US$ at the above rates.

Cash Flow Discount Rate Range %	6 - 15 %	Range of discount rates applied to calculate the Project NPV

Closure and Reclamation Costs	US$33.6M	Closure of the Fenix Project at end of mineral reserves is assumed. A provision for Closure and Reclamation of project assets based on a discounted cost of US$40.50M estimated for Years 30, 31, 32 of the pyro project, brought back to Year 30.

Equity	100%	All equity funded (unleveraged cash flows)

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18.12.3

Production Cost per lb Nickel

A summary of the production costs derived from the CFM, in line with the Brook Hunt definition of production costs, is presented in Table 18-45. These profile the Fenix Project cash costs per lb nickel as the Net Direct Cash Cost (C1). The C1 costs are derived for production years 1 to 5 and for production years 6 to 20. In the first time-frame, production is ramping up and electrical power is purchased though the PPA. In the second time-frame, the Fenix plant is operating at full production using site generated power.

Table 18-45: Summary of Production Costs after Brook Hunt


Cash Cost Category (Brook–Hunt)	Cost / Credit Item	Production Years 1 – 5 US$/lb Ni		Production Years 6 – 20 (15 years) US$/lb Ni
Net Direct Cash Cost (C1)	Total Site Operating Costs	3.30		2.16
	Transport, Freight and Delivery of FeNi	0.18		0.18
	Freight Insurance of FeNi	0.01		0.01
	Marketing Fees	0.19		0.19
	By-product Credits	(0.20	)	(0.20	)
	Net Direct Cash Cost (C 1)	3.47		2.34

18.12.4

Capital Costs

The capital costs for the Fenix Project are summarized in Table 18-27.

18.12.5

Key Capital Cost Assumptions

The following key assumptions were made regarding capital costs in the CFM (additional qualifications, assumptions and exclusions are set out in Section 18.11.1.4 and 18.11.1.5 above):

•

Capital costs are as of July 2007, with no currency escalation, or inflation factors applied.

•

All capital costs are in US dollars and include only applicable payroll taxes on construction labour. VAT and other duties are not included. Skye expects to receive VAT and duties exoneration on all imported supplies and services. Likewise it is assumed that all VAT on locally supplied goods and services will be recovered.

•

All capital cost expenditures are made with full payment in the year indicated.

•

Capital cost expenditure starts in Year 2 and is complete in Year 1 for the Phase 1 Capital project. For the Phase 2 Capital Project, expenditure starts in Year 2 and is complete in Year 5.

18.12.6

Capital Expenditure Profiles

In order to define the project capital expenditure distribution profile for Year –2 and Year -1, the Project Schedule and Implementation Plan were used to estimate the capital expenditure in each time period for each WBS area.

Figure 18-14 shows the overall capital distribution of expenditure for the Fenix project. The profile covers a pre -production period of 8 quarters ( quarter - 8 to quarter -1) in Year -2 through to start Year 1.

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FENIX Project Forecast of Expenditures in US$

Cashflow (Incurred Cost) based on Feasibility Study Estimate

LOGO

Figure 18-14: Fenix Project Capital Cost Expenditure Distribution Profile

18.12.7

Working Capital

The working capital requirements for the Fenix Project were estimated and included in the CFM. Working capital of the Fenix Project at any time comprises the aggregate of closing values for cash, accounts receivable, work in progress, product inventory, warehouse supplies and spares inventory, less accounts payable.

18.12.8

Sustaining Capital

In the cash flow model, capital outlays to sustain the name-plate capacity of the assets was made to account for costs associated with capitalized rebuilds, refurbishment projects, replacement of equipment and major spares that do not enhance, but only maintain business operations. Such capital generally applies to maintenance shutdowns, refurbishment projects and does not typically include consumables and minor spares used on a daily basis to maintain the facilities that are included as maintenance materials in site operating costs. A sustaining capital estimated was done for every year of operation for the following items:

•

Kiln and Furnace Rebuilds.

•

Power Plant Shuts and Ash Disposal.

•

Mine Development and Equipment Replacement.

•

Process Plant Mobile Equipment Replacement.

•

Computer Hardware and Software Replacement.

•

General Mine, Plant and Site Capital Maintenance.

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18.12.9

Owner’s Cost

The Owner’s Costs included in the Financial Analysis were developed by Skye including the expenditure profile.

18.12.10

Revenue

Sales Revenue for CGN is determined from the amount of contained nickel and iron in refined FeNi on a dry basis. In the CFM, sales revenue is determined as:

•

Total Realized Metal Value which is calculated from the realized value of the contained Ni and Fe metal in FeNi based on the market metal prices.

•

Net Sales Revenue which is the net ‘at plant gate’ revenue calculated from Realized Metal Value less freight and insurance costs.

In calculating the realized value of contained Ni in FeNi, 100% of the market nickel price is assumed as the realized price. The base market nickel price is US$6.50 per pound of nickel contained (US$14,342/tonne Ni). Skye has concluded that an iron credit is payable to CGN as revenue based on the Fe contained in FeNi. An Iron Credit of US$0.20/lb Ni is assumed so the total assumed base nickel price is US$6.70.

From Figure 18-15 the following can be noted regarding Fenix Project revenue streams, at a base Ni price of US$6.50/lb Ni:

•

In the first 3 years Net Sales Revenues depend strongly on the production ramp up rates.

•

Fenix Project net sales revenues approximately reach US$320M pa at full production. In years 9, 16 and 25 the reduction in revenue is clearly seen due to the loss in production from planned outages of furnace and power plant for major maintenance, as described earlier.

•

In the last 10 years (years 21- 25, 26 - 30) falling ore nickel grades reduces revenue to US$275M and to US$230M, respectively.

LOGO

Figure 18-15: CGN Annual Net Sales Revenue (at $6.50/lb Ni and $0.20/lb Ni Fe Credit)

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18.12.11

Depreciation

In the CFM, depreciation of capital assets is applied to site and equipment capital assets, pre-operational capital and sustaining capital over the life of operations under the alternative tax regime. No deduction for depreciation is allowed under the “general tax regime”. For each year of production, a calculation is made of depreciation for income tax purposes based on the character of the outlays, its eligibility for depreciation and the prescribed rates under the tax law and regulations. The total of depreciation claims per year is then a deduction from gross revenue of the Fenix Project to determine the annual taxable income. Therefore, depreciation of capital assets each year for tax purposes only affects the CGN Project’s Net Cash Flow in those years when CGN has made an election before December 1 of the prior fiscal year to have the alternative tax regime applied to its revenue.

For all new capital assets (all categories except buildings), straight-line depreciation over 5 years is assumed to apply (20% per year). For buildings, including improvements to buildings, straight-line depreciation over 20 years is assumed to apply (5% per year). For sustaining capital investments, a 50% average annual claim for acquisitions in a year is assumed. For pre-operational expenses, 50% of sunk costs of CGN incurred prior to July 1, 2007 is assumed to be eligible for depreciation beginning with commercial production.

18.12.12

Interest Payments

This economic analysis presents an assessment of the project return and does not consider financing options. Consequently the CFM does not account for interest payments on CGN loans, revolving credit facilities or the financing costs associated with debt funding of the project. Hence, no application is made of such interest payments as a deductible expense from net earnings to determine taxable income.

Such an approach assuming 100% equity funded project (unleveraged) gives a reasonably accurate representation of the economics of the CGN operations and net returns of the Project. The funding arrangements and associated costs for the project, via debt and /or equity have been excluded from this economic analysis.

18.12.13

Taxable Income (Alternative Income Tax Regime)

CGN’s taxable income was calculated in the CFM as Total Net Sales Revenue (at plant) less:

•

Cost of Goods Sold (Operating Costs adjusted for inventories)

•

General and Administration

•

Inco Payments

•

Other Payments to Government

•

Depreciation

The income taxes payable by CGN, under either the general income tax regime or the alternative income tax regime, are shown in below in Figure 18-16.

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LOGO

Figure 18-16: Total Income Tax Payable by CGN (at US$6.50/lb Ni)

18.12.14

Project Cash Flows

Figure 18-17 presents the annual net after tax cash flows for CGN operations over 30 years of operations, using a constant base Ni price of US$6.50/lb Ni.

LOGO

Figure 18-17: Net Cash Flows for CGN (at US$6.50/lb Ni)

From Figure 18-17 the following may be observed:

•

In the first five years of operation, the higher power cost and the capital expenditure of Phase 2 is expected to significantly reduce cash flow.

•

Net Cash Flows (after tax) rise significantly to about US$175.3M and US$173.8 in production years 6 and 7 with net cash flows ranging from US$120-160M beyond Year 7.

•

In the major shut years 9, 16 and 25 significant reductions in net cash flow occurs due to the planned loss of production.

•

In the last 5 years on average a Net cash flow of about US$83M is expected to be realized.

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18.13

Project Payback

18.13.1

Project Returns: NPV, IRR and Payback

At a base nickel price of US$6.50/lb Ni the Fenix Project gives an IRR of 14.3%.NPV’s at various discount rates are given in Table 18-46. For this nickel price, the payback is estimated at 7.3 years, the time point at which the cumulative net cash flows become positive.

Table 18-46: NPV of Fenix Project Net Cash Flows (at US$6.50/lb Ni)


Nickel Price Discount Rate %	US$6.50/lb Ni NPV US$M
6	797.2
8	490.6
10	275.0
12.5	89.3
15	(62.6	)

Table 18-47 shows the effect of Ni price on simple payback year of the project and on the IRR.

Table 18-47: Impact of Ni Price on Payback Period and IRR


Nickel Price	Simple Payback After Production Year	IRR %
US$5.50/lb Ni	10.1	9.8
US$6.50/lb Ni	7.3	14.3
US$7.50/lb Ni	6.1	17.8
US$8.00/lb Ni	5.6	19.8

18.13.2

Sensitivity Analysis

Using the CFM, a sensitivity analysis was performed on the Fenix project by variation of the key economic variables of:

•

Nickel metal price.

•

Site operating costs.

•

Project capital costs.

•

Petroleum coke price.

•

Oil based supplies.

Nickel price was varied from US$5.50/lb Ni to US$8.00/lb Ni with a Ni price of US$6.50 as the base case price. Nickel price variation shows the following impact on the project economics, presented in Table 18-48. A 5% change in Nickel price equates to a NPV(8%) change of about US$88M.

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Table 18-48: Fenix Project Sensitivity to Ni Metal Price


Economic Variable	Variable Change %	Ni Price US$/lb	NPV (8%) US$ million	NPV (10%) US$ million	IRR %	Payback Years
Nickel Price	-15	5.50	132.5	-16.7	9.8	10.1
	-12	5.75	227.6	60.4	11.0	9.3
	-8	6.00	307.6	127.0	12.1	8.5
	-4	6.25	400.3	202.1	13.2	7.8
	0	6.50	490.6	275.0	14.3	7.3
	4	6.75	579.0	346.1	15.3	6.9
	8	7.00	637.3	394.6	16.1	6.6
	12	7.25	714.0	455.2	16.9	6.4
	15	7.50	800.1	524.3	17.8	6.1
	23	8.00	977.2	666.7	19.8	5.6

Site Operating Costs are varied from –20 to +20% with 5% increments and shows the following impact on the project economics, presented in Table 18-49. Each 5% variation equates to about US$58M of NPV (8%).

Table 18-49: Fenix Project Sensitivity to Site Operating Costs


Economic Variable	Variable Change %	Value of Change US$M		NPV (8%) US$M	NPV (10%) US$M	IRR %	Payback Years
Site Operating Cost	-20	(659.5	)	722.9	466.1	17.2	6.2
	-15	(494.6	)	664.8	418.4	16.5	6.5
	-10	(329.7	)	606.8	370.6	15.8	6.7
	-5	(164.9	)	548.7	322.8	15.0	7.0
	0	0.0		490.6	275.0	14.3	7.3
	5	164.9		432.5	227.2	13.6	7.6
	10	326.8		375.5	180.3	12.8	8.0
	15	494.6		316.3	131.6	12.1	8.5
	20	659.5		258.3	83.8	11.4	9.1

The Project Capital Costs were varied with 5% increments from+20% to –20%, as shown in Table 18-50. This 5% variation equates to about US$39M NPV (8%).

Table 18-50: Fenix Project Sensitivity to Project Capital Cost


Economic Variable	Variable Change %	Value of Change US$M		NPV (8%) US$M	NPV (10%) US$M	IRR %	Payback Years
Project Capital Cost	-20	(196.5	)	648.1	424.7	17.8	6.1
	-15	(147.4	)	608.7	387.3	16.8	6.4
	-10	(98.3	)	569.4	349.9	15.9	6.7
	-5	(49.1	)	530.0	312.4	15.1	7.0
	0	0.0		490.6	275.0	14.3	7.3
	5	49.1		451.2	237.6	13.6	7.6
	10	98.3		411.8	200.1	12.9	7.9
	15	147.4		372.4	162.7	12.3	8.4
	20	196.5		333.0	125.2	11.7	8.8

The Petroleum coke price, which is the key cost driver of the electrical power cost from Years 6 to 30, was selected for sensitivity analysis given the fact that energy represents the single largest component of operating costs. The sensitivity of the Fenix Project to petroleum coke price is shown in Table 18-51.

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Table 18-51: Fenix Project Sensitivity to Petroleum coke Price


Economic Variable	Variable Change %	Value of Change		NPV (8%) US$M	NPV (10%) US$M	IRR %	Payback Years
Petroleum Coke Price	-20	(12.2	)	513.4	292.2	14.5	7.2
	-15	(9.2	)	507.7	287.9	14.4	7.3
	-10	(6.1	)	502.0	283.6	14.4	7.3
	-5	(3.1	)	496.3	279.3	14.3	7.3
	0	0.0		490.6	275.0	14.3	7.3
	5	3.1		484.9	270.7	14.2	7.3
	10	6.1		479.2	266.4	14.2	7.3
	15	9.2		473.5	262.1	14.1	7.3
	20	12.2		467.7	257.8	14.1	7.4

From the curves in Figure 18-18 following may be noted regarding the relative sensitivities of the Project to changes in Ni price, petroleum coke price, site operating costs and capital costs:

•

The project is most sensitive to Ni price, with approximately a +1.0% gain in IRR for an increase of 5% (US$0.25/lb Ni). This Ni price increase results in an estimated NPV (10%) gain of US$71M and an estimated NPV (8%) gain of US$88M.

•

The project is less sensitive to changes in project capital costs and operating costs.

LOGO

Figure 18-18: Fenix Project Relative Sensitivity to Variable Changes

From Figure 18-19 the Fenix Project is sensitive to the prices of the major energy sources, such as petroleum coke, coal and oil based supplies (like diesel and HFO). The HFO is a key cost driver of the electrical power cost from Years 1 to 5 based on the PPA under negotiation between Skye and an IPP.

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LOGO

Figure 18-19: Fenix Project Relative Sensitivity to Fuel Supplies

18.14

Project Risks and Opportunities

18.14.1

Introduction

The Fenix project is large, relatively complex, and located in a developing country, and as such, involves some risks and challenges. These risks and challenges have the potential to affect:

•

The performance of the facility in terms of production.

•

The cost of construction and operation.

•

The implementation schedule.

•

The time required to reach full capacity.

•

The environmental performance and impact of the mine and plant.

Any of the above items could affect the financial performance of the project.

18.14.2

Project Risks

Major capital projects, such as the Fenix project, are subject to risk. Generally, project risks can be categorized as:

•

Resource/Reserve Estimation and Mine Planning

•

Regulatory/Approvals/Permits

•

Design/Process

•

Financial/Commercial/Contractual

•

Logistics/Access

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•

Construction, Testing & Commissioning

•

Political

•

Environmental

•

Health, Safety & Security

•

Labour

•

Social

A comprehensive list of all of the risks that apply to the Fenix ferro-nickel project has not been developed and is beyond the scope of expertise of the authors of this report. Project risks (specific technical risks that apply to the Fenix project, together with certain other risks that Hatch has actual knowledge of due to our involvement in the project) are described below.

18.14.2.1

Resource/Reserve Estimation and Mine Planning

The estimates of the mineral resources and mineral reserves have been prepared in accordance with relevant CIM guidelines; this should not be construed as a guarantee that such mineral resources actually exist in the estimated quantities or are mineable.

18.14.2.2

Regulatory/Approvals/Permits

18.14.2.2.1

Land Acquisition

To date the following land/property has not been acquired by Skye/CGN: the Santo Tomas trans-shipment terminal site (Skye has, however, purchase a lease option for a suitable tract of land adjacent to the port) the Chichipate and Sepur Limite road bypasses. Failure to obtain these properties in a timely manner could impact the project implementation and start up.

18.14.2.2.2

Approval for Interconnection to Guatemalan Electrical Power Grid

An application has been submitted to the Guatemalan electrical power grid operator (CNEE) for the interconnection of a 230 kV power transmission line from the Fenix plant site to the Guatemalan national grid. Failure to obtain permission for the interconnection is a material risk to the project.

18.14.2.3

Design/Process

From1979-1981, Inco mined the Fenix deposits and smelted the ore to matte in the existing smelter. Because the ore has been smelted on a commercial scale, it is subject to less technical uncertainty than a comparable greenfield project. The Fenix project will incorporate several modifications to the existing facility to increase capacity, reduce operating costs and to produce a ferro-nickel product. While these measures will improve the competitiveness of the business, they introduce some technical uncertainty. The key technical risks associated with this are described below.

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18.14.2.3.1

Coal Firing of Kilns

The kilns are to be fired with pulverized coal, and not heavy fuel oil as was done in the original Exmibal smelter. This is being done to reduce the consumption of high cost liquid fuel. There is technical and commercial precedent for doing this at the Pamco and Hyuga smelters in Japan. The main uncertainty is the proportion of heavy fuel oil needed to ensure flame stability in the kiln. Pamco normally uses 100% coal to fire their kilns (except during start-up), while Hyuga augments the coal fed to the burner with fuel oil.

The operating costs estimated in this Feasibility Study are based on 100% coal-fired kilns (except during start-up). The risk to the project is that some amount of supplemental fuel oil may be required, resulting in somewhat higher operating costs.

18.14.2.3.2

Single Furnace Operation at 90 MW

The proposed Fenix process plant consists of two rotary kilns feeding calcine to a single electric furnace. As such, the plant’s capacity to process ore will be limited by the capacity of the electric furnace and its ancillary systems (e.g., power supply, feed system, off-gas system, slag granulation and metal tapping).

The time required to ramp-up the furnace and its ancillary systems to operate at 90 MW on a sustained basis may exceed schedule expectations.

At the present time, there are no nickel laterite furnaces operating at 90 MW on a sustained basis. PT Inco is approaching this power level and Cerro Matoso’s line 2 furnace, while designed to operate at up to 90 MW, is limited to 75-80 MW due to limitations in kiln throughput.

Therefore the Fenix flowsheet represents an incremental increase in smelting capacity.

18.14.2.3.3

Captive Power Plant

After 2014 the Fenix project will have a power plant captive to the smelter as in the original Exmibal configuration. A key difference is the change from immersed electrodes to shielded arc furnace operation which will introduce larger load swings and potential load fluctuations to the power plant. To reduce impact of these power fluctuations on the performance of the power plant, the Fenix scope includes furnace power compensation equipment (SPLC) that has been commercially proven at Falconbridge Dominicana in the Dominican Republic. Given the established grid connection, however, this risk is largely mitigated.

18.14.2.3.4

Restart of Mothballed Equipment

Fenix will utilize a great deal of equipment that has been idle for over 25 years. This equipment has been visually inspected by vendors (major equipment) and discipline engineers (secondary equipment), in 1993 (Davy) and in 2005 (Hatch), and is generally considered to be in good condition. Equipment inspections have been limited, however, the equipment has not been disassembled, electrical power has not been connected, etc. Until the equipment is restarted there will be uncertainty in its ability to perform as required. The consequence of this could be additional costs, if further refurbishment or replacement is necessitated, and a longer ramp up period could result from the associated delays.

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18.14.2.3.5

Geotechnical

Preliminary geotechnical investigations have been completed for the mine, stormwater containment dams process/power plant, and slag disposal areas. Further investigation has been initiated and should confirm the assumed geotechnical conditions. Variations from assumed geotechnical conditions would have an impact on capital costs.

18.14.2.4

Financial/Commercial/Contractual

18.14.2.4.1

Economic Conditions

The cost estimates (capital and operating) and economic analysis have been developed based on January 2006 market conditions. Variations in market conditions (affecting commodity and/or labour costs), or in the assumed foreign exchange rates, could effect the project economics.

18.14.2.4.2

Securing Interim Power Supply

CGN is negotiating with an Independent Power Producer (IPP) an interim supply of power through a Power Purchase Agreement (PPA) and a Transmission Tolling Agreement (TTA). Failure to conclude the PPA/TTA negotiations on the terms described in Sections 18.4.2 and 18.4.3, is a material risk.

18.14.2.4.3

Community Support

The Fenix project will have an impact on the communities situated in proximity to the mine, process/power plant and transportation routes. The success of the project will depend on the support of those communities.

18.14.2.5

Overland Transport of Equipment and Materials to Site

The project has assumed that all construction equipment and materials, including heavy and oversized equipment (i.e. kiln sections, dryer sections, transformers, etc.) will be transported from the port of Santo Tomas to site by truck. Special permits and equipment will be required for the transport of the oversized and/or overweight (greater than 55 tonnes) loads, temporary supporting of bridges may be required, a bypass of the 11 km section of highway 7E west of Fronteras constructed, etc. A cost allowance has been included in the Phase 1 capital cost estimate for the construction of an unpaved bypass of the referenced section of highway 7E. Further work is required to confirm the location of the new road alignment, the geotechnical conditions, etc., and to develop a +/-15% cost estimate and an execution schedule. The failure of obtaining all necessary permits for the upgrades to bridges and roads or to complete the upgrades in accordance with the project schedule and budget is a material risk.

18.14.2.6

Construction & Commissioning

18.14.2.6.1

Weather Conditions

The project implementation schedule has been developed based on typical weather conditions for the region. Adverse weather conditions could prolong the schedule and impact the project.

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18.14.2.7

Political and Regulatory

The Fenix project is located in Guatemala and, as such, is susceptible to certain risks including currency fluctuations, political or financial instability, exchange controls, aboriginal issues, changes in mining taxation and regulations, export controls, changes in permit and licensing requirements, delays in the issuance of permits, embargos, and environmental issues, all of which may materially and adversely impact the project.

18.14.2.8

Labour

Of particular importance for Fenix is labour cost. There is little precedent in Guatemala for a project of this type so the cost and productivity of construction labour has some degree of uncertainty. Variations in labour rates and/or productivities from those assumed for the project (refer to section 18.11.1.3) could impact the capital cost of the project.

18.14.2.9

Social

On September 16, 2006, three parcels of land owned or leased by CGN were illegally occupied. Subsequently, on October 3 and 4, two additional parcels of land were illegally occupied. The apparent objective of these occupations was to obtain land for housing and subsistence farming and, because land occupations are not uncommon in Guatemala, to bring attention to unresolved land tenure issues. These occupations have not caused any material interruptions of the Project, although some remediation of exploratory drill sites was delayed pending resolution of the invasion of one site.

On January 8 and 9, 2007, the local Guatemalan District Attorney, assisted by a special unarmed team of the national police, directed the removal of the groups from CGN’s property, carrying out a Guatemalan judge’s order. While some families returned to four of the parcels in contempt of court orders, an agreed upon solution has been reached with respect to two of the groups and positive discussions continue with all of the groups involved. CGN has initiated a comprehensive strategy to develop long-term solutions to the occupations as well as preventative measures for potential future occupations, including hiring international experts to work alongside our local community relations team and developing partnerships with, and solutions involving, the Guatemalan government and international and local housing and land rights organizations. CGN’s efforts also benefit from the participation and commitment of local church and rights organizations as CGN continues to work towards a satisfactory, long-term resolution with the families involved. The CGN community relations team is working with the full support of Skye management to resolve the remaining occupations as quickly as possible, and CGN remains committed to building and maintaining good relationships with the local communities.

The project will be vulnerable to road blockages arising from protests.

18.14.3

Opportunities

Some opportunities for the Fenix project are described below.

18.14.3.1

Power Plant

As Fenix will be connected to the Guatemalan grid. CGN will likely have the opportunity to enter into a long-term PPA with a third party based on a solid-fuel power plant or another power source not dependent on petroleum prices. A substantial reduction in Phase 2 capital and the avoidance of a significant increase in project transportation requirements could be achieved if electric power is supplied through a third party.

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18.14.3.2

Mineral Resources

Five resource areas were evaluated in detail for the feasibility study and scheduled for future mining. Other Fenix mineral resources, both within the exploitation licence and to the south of Lake Izabal at Montúfar, represent a significant opportunity to achieve environmental, operational and financial gains for the project. The properties contain additional saprolite mineral resources of:

•

Measured + indicated: 35.8 Mt grading 1.81% Ni

•

Inferred: 55.4 Mt grading 1.64% Ni

Some of the other resource areas within the exploitation license may be able to provide better grade material to the plant during at least part of the proposed operating period. For example, Area 218-1 contains estimated indicated mineral resources of 2.6 Mt grading 1.95% which could partially replace lower grade ore from Area 215, which has an average resource grade of 1.77% Ni. A smaller deposit within 2 km of the plant, Area 211, also contains 0.7 Mt grading 1.85%.

The alternate mineral resources also offer the opportunity to partially replace or defer higher cost ore from Area 251. Operational costs related to this more distant area add $1.35/tonne of plant feed. In addition, potential environmental and social impacts of development of the area and transporting large quantities of material over 20 km of public road present management challenges which could be mitigated by exploiting more proximal deposits using established infrastructure.

The additional mineral resources may also significantly extend the overall life of the operation. It is expected that future evaluation of the mineral resources not considered in the feasibility study will demonstrate saprolite of suitable grade and chemistry to support operations beyond the 30 year project.

18.15

Mine Life and Exploration Potential

Mineral reserves estimated by Snowden support maximum annual production of 1,464,000 tonnes of saprolite ore over a project mine life of 30 years. Mining will be conducted in five deposit areas, 212, 213, 215, 217 and 251. All mineral reserves are based upon measured and indicated mineral resources.

The best defined saprolite exploration potential are the other measured, indicated and inferred mineral resources listed in Table 1.1. This potential is largely related to mineral resources in Area 218 7 km west of Area 217, Montufar which is south of Lake Izabal and in more distant areas where inferred mineral resources have been defined by wide-spaced, hand auger reconnaissance drilling at 200 – 500 m centres. With the exception of Area 218 the other mineral resources are lower in Ni grade than the measured and indicated. However, if substantiated by further exploration, they could add in the order of an additional 30 years of saprolite production.

More hypothetically, there remain several untested landforms at the most favourable elevations in the Fenix license in the Amate-Chulac area and some small untested landforms between the low elevation areas 214 and 245 and the terraces of 217-5 and 218 in La Gloria. The possibility of transported limonite ores in relatively low lying valleys in the Sierra Santa Cruz has never been tested.

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18.16

Hyrdometallurgical Expansion Preliminary Assessment

19.

Interpretation and Conclusions

19.1

Geology and Mineral Resources

The Fenix nickel deposits are typical wet-tropical climate laterites and occur as surficial deposits on spurs and terraces in the Lake Izabal region of north-eastern Guatemala. Nickel laterites typically include limonite and saprolite layers and these are both represented in the Fenix deposits.

In 2005 CGN undertook a two phase diamond core drilling program to evaluate known laterite mineral resources in the 212, 213, 215, 217 and 251 areas; and to validate the historic sample database in support of the Fenix feasibility study. Dr. Paul Golightly of Golightly Geoscience verified the historic data and compiled a located sample database of nickel, cobalt, iron, magnesia, silica and dry density. CGN’s core drilling assays were used to establish factors to correct an iron bias in the historic sample database, and to derive calculated values for magnesia, silica and dry density as these variables were not present in the entire historic database.

In Golightly Geoscience’s opinion, the core sampling, sample preparation, analytical procedures and security for the CGN programs are industry standard. The procedures for sampling, compiling and security of the historic data are acceptable and there has been sufficient checking from original sources to confirm that the historic database is acceptable for Mineral Resource estimation.

Snowden prepared limonite, transition and saprolite Mineral Resource estimates for Ni, Co, Fe₂O₃, MgO, SiO₂ and dry density for Areas 212, 213, 215, 217 and 251 by first constructing 3-D wireframes of the laterite layers and then interpolating the grade variables by ordinary kriging into conventional 3-D block models in accordance with CIM guidelines. The Mineral Resource estimates were validated by alternative interpolation methods and classified as Measured, Indicated and Inferred categories consistent with the requirements of CIM and NI 43-101. The classification scheme took borehole spacing, mineralization continuity and relative error into account (Table 17-1 and Table 17-2).

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The Snowden estimates for two blocks in Area 212 were compared with mine production and reconciled well on a global basis. The within-pit portion of the resource model reconciled in terms of overburden volume and mineralized material grade and tonnes to within 8% of the historical production data on a global basis. Bench-by-bench reconciliation, although poor, was as expected given the nature of the mining process and the uncertainties in local estimates derived from wide-spaced exploration data.

Information from historic mining and the latest core drilling indicates that the full depth extent of the laterite profile and short scale variation can be misinterpreted from drilling data alone. This uncertainty is addressed in the Mineral Resource classification scheme, whereby Measured category in saprolite is achieved by drilling at spacings of 35 m or less, and spacings of 35 m to 75 m are required for Indicated saprolite mineral resources.

Golightly Geoscience reported Mineral Resource estimates for the other deposits that occur in CGN’s licences. The estimates were originally prepared for all areas using Inco’s 2-D LES method. The LES estimates in five of the areas however are superseded by the Snowden 3-D estimates (Table 17-1 and Table 17-2). The LES estimates for saprolite and 2-D estimates of limonite by Snowden for the other deposits are provided in Table 17-3 and Table 17-4. The LES resource estimates were realistic in their stated purpose of defining a resource for a Preliminary Assessment and these deposits will require re-estimation in 3-D format to enable mine planning to proceed.

19.2

Mineral Reserve Estimates and Mining

In accordance with the CIM (2000) definitions in NI43-101, Snowden derived proven mineral reserves from measured mineral resources and probable mineral reserves from indicated (and in a few instances, measured) mineral resources. No inferred mineral resources were used in the estimates or in mine planning. Mineral reserves yield a positive cash flow based on project economic assumptions and detailed technical evaluations of mine production and ore processing. Total proven and probable mineral reserves for the project are 41.3 million tonnes grading 1.63% nickel, 18.2% Fe with SiO₂/MgO of 1.54.

Snowden utilized the mineral resource block models with appropriate dilution and ore loss in conjunction with financial criteria in Whittle Four-X software to generate a series of nested pits shell shells. A two stage approach designed to maximize recovered grade in the first 24 years of the project, followed by lower grade in the final 6 years. Within the selected pit shells, the mine schedule is designed to maximize grade to the extent possible by utilizing high grade and low grade stockpiles. The schedule is also constrained by the ore chemistry and the requirement for to meet criteria for the SiO₂/MgO and Fe/Ni ratios. Scheduling also provides a rational approach to exploitation by working progressively through each area.

Mine equipment has been sized according to the anticipated mine conditions, production demands and work schedule. Mining will be carried out with large hydraulic excavators loading 40 tonne mine trucks. Overburden and limonite are stripped sequentially to expose the saprolite ore. Excavation proceeds from the uphill part of the deposit to the base; no blasting is required. Progressive reclamation for mined out areas is planned over the life of the operation and ultimately all pits will be backfilled and re-vegetated.

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20.

Recommendations

20.1

Mineral Resource Evaluation

The following program is recommended by Golightly Geoscience and Snowden as useful to the continuing planning process:

•

Potential for upgrading saprolite by corestone rejection. The observations by Exmibal during mining shows that this opportunity is unlikely to occur in highly serpentinized saprolite in 212 and 213. However the degree of serpentinization in other areas subject to in-fill drilling appears to be lower and there is potential for upgrading by rejecting fresh corestones during mining. Bulk sampling or large diameter core drilling tests of a few sites in each of 215, 217 and 251 would provide:

•

Size distribution of the nickel grade.

•

Additional large scale bulk density determinations.

•

A direct measure of the lithological and structural controls on those deposits.

•

Additional check assays. External check analyses suggested a small low nickel bias in the first 3,000 samples of the 2005 diamond drilling and a high alumina bias, in the second half of the program. The external check samples should be re-analyzed.

•

Geological mapping. There is considerable outcrop on drill roads in and near areas 217 and 251 as well as excellent exposure of 213. These should be mapped with particular emphasis on structural features.

•

Additional drilling, both twin and in-fill, followed by updated mineral resource and reserve estimates, where warranted, in the following areas:

•

Area 218 deposits.

•

Montúfar property, the Cristina area 221 & 222 deposits.

•

The lower terraces of area 215, south of terrace 215-1. In-fill drilling at 50 m spacing is recommended for 215-3, 215-2 (North) and 216 (north), 260 (Amate) and 219.

•

Depth extensions. Much of the historic drilling did not penetrate to the bottom of the weathering profile and despite the relatively effective penetration of the bedrock by diamond drilling, many diamond holes in area 215-1 terminated in nickel grades greater than 1.5%. Skye’s mining and grade control plan needs to address troughs and pinnacles in the bedrock.

An estimate of costs is provided in Table 20-1.

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Table 20-1: Recommended Budget


Area	Holes	Length	Cost US$		Purpose	Status
217-1	350	8,255	$	1,816,000	37 m	Complete
216	100	2,359	$	518,800	75 m in-fill	Complete
217-2&3	442	10,425	$	2,293,400	37 m	Complete
215-2	135	3,184	$	700,400	50 m in-fill	Complete
215-1	38	896	$	197,100	West Extension	Complete
215-3	103	2,429	$	534,400	50 m in-fill	Complete
216	284	6,698	$	1,473,600	37 m in-fill	Planned contingent
Subtotal	1,452	34,246	$	7,533,700
217	20	472	$	414,800	20 cm core holes or	Recommended
215	20	472	$	414,800	bulk sampling, for density	Recommended
251	20	472	$	414,800	and upgrading potential	Recommended
Subtotal	60	1,415	$	1,244,400

Additional Drilling						Priority
218-2&2	510	12,028	$	2,646,200	50 m in-fill	1
218	355	8,373	$	1,842,000	50 m in-fill	1
221	100	2,359	$	518,800	Initial Assessment Twins	2
222	20	472	$	103,700	Initial Assessment Twins	3
260	50	1,179	$	259,400	50 m	4
219	56	1,321	$	290,500	50 m	5
212	153	3,609	$	793,800	Depth Extension	6
217-4&5	31	731	$	160,800	Depth Extension	6
215-1	87	165	$	36,300	Depth Extension	7
217-1&2&3	94	2,217	$	487,700	Depth Extension	7
Subtotal	1,456	32,453	$	7,139,200
Database			$	85,000
verification
Resource			$	70,000
estimates
Total	2,968	68,114	$	16,072,300

Purpose of the 50 metre infill drilling is to upgrade the area to Indicated and the 37 metre infill drilling is to upgrade the area to Measured

20.2

Mining

Snowden recommends that subsequent mining studies:

•

Explore the opportunity for additional stockpiling to possibly reduce the amount of sterilized ore.

•

Evaluate the use of highway trucks for off-highway ore haulage.

•

Investigate the aerial tramway or alternative methods for transport of ore from area 251 to the transfer station.

•

Investigate detailed access within the selected pit shells and for the bypass roads.

•

Evaluate contractor hauling from area 251 to the plant.

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20.3

Process and Power Plant

During subsequent phases of the project Hatch recommends the following:

•

A detailed assessment of upgrade requirements to public roads and bridges between Santo Tomas and the Fenix plant site be completed. The study should include a +/- 15% capital cost estimate, an implementation schedule and a proposed alignment between Sumach and Fronteras.

•

Additional geotechnical studies be conducted to confirm foundation design parameters.

•

Core samples of existing concrete and structural steel specimens will be obtained and tested to confirm material properties.

•

To complete a more comprehensive inspection (including disassembly) of major equipment (i.e. steam turbine generator, electrode columns) to confirm refurbishment requirements and costs.

•

Visits to existing nickel laterite smelters that utilize pulverized coal to fire their kilns to confirm design and operating practices.

•

That more detailed Seismic Zoning studies be completed to confirm building design requirements.

21.

References

AMEC, 2003. Technical Report: Exmibal Nickel Project AMEC: by Dominique François Bongarçon, Ph.D. & Brian Montpellier, P.Eng. December 2003.

Barnard, 1972a. Guatemala Pitting and Upgrading Program, Status Report for 1971. Inco – Exmibal Internal Report, 1 January, 1972.

Barnard, 1972b. Guatemala Special Pitting Program, Status Report for 1972. Inco – Exmibal Internal Report, 28 December, 1972.

CIM, 2000. CIM Definition Standards – for mineral resources and mineral reserves. Adopted by CIM Council on 20 August, 2000.

CIM, 2003. Estimation of mineral resources and mineral reserves. Best practice guidelines. 30 May, 2003 adopted by CIM Council on 23 November, 2003.

CIM, 2005. CIM Definition Standards – for mineral resources and mineral reserves. Adopted by CIM Council on 11 December, 2005.

Exmibal, 1980. Producción a la fecha, Diciembre 1980.

Golightly, 1977. Guatemala La Gloria: Statistical inference of cobalt from nickel+cobalt and iron data. Inco - Exmibal Internal Memo from J.P. Golightly to R.A. Alcock. 17 November, 1977.

Harju, 1979. Exploration of Exmibal’s Nickel Laterite Deposits in Guatemala. In International Laterite Symposium, AIME, New Orleans, pp245-252.

Hatch, 2005. Fenix Preliminary Assessment: Prepared by: T. Armstrong, P.Eng., B. Krysa, P.Eng, J. Sajer, P.Eng, J.P. Golightly P.Geo. for Skye Resources Inc. 5 August, 2005.

MacKenzie, 1980. Grade control procedures, Exmibal mine. 20 November, 1980.

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MacKenzie, 2006. Bulk Density Pitting Project, Fenix Project, El Estor; Internal CGN report. February 2006.

Skye, 2005. Phase 5 Pilot Plant Ore Samples. Memo from D. Neudorf. 19 June, 2005.

Skye, 2005 AIF. Annual Information Form for the Financial Year ended 31 December, 2005. 6 April, 2006.

Snowden, 2005 Fenix deposit, drill spacing and resource classification. Letter report from M. Murphy, Snowden to C. McKenzie, Skye. 19 October, 2005.

Snowden, 2006 Mineral Resource Estimate, Fenix Project, Izabal, Guatemala. NI43-101 Technical Report: Prepared by A. F. Ross P.Geo., J.P. Golightly P.Geo., F. Poretta P.Eng., B.D. Krysa P.Eng. for Skye Resources Inc. 4 July, 2006.

Snowden, 2006a Interim reconciliation report – Areas 212-2 and 212-5. Internal Memo from W. Board to C. McKenzie, Skye. 1 June, 2006.

Snowden, 2000b Fenix project, Model selectivity. Internal Memo from M. Murphy to A. Ross. 23 June, 2006.

Snowden, 2006c Fenix drillhole spacing study. Internal Memo from A. Trueman to C. McKenzie, Skye. 13 July, 2006.

Sopko, 1979. The Exmibal Nickel Project. In International Laterite Symposium, AIME, New Orleans, pp273-291.

Toomver, 1979. Development of Inco’s Selective Reduction Smelting Process for Nickel Laterite Ores. In International Laterite Symposium pp255-257.

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22.

Dates and Signatures

Technical Report

Technical Report on an Update to the Fenix Project, Izabal, Guatemala

19 November, 2008

LOGO

Date 19 November 2008

LOGO

Date 19 November 2008

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23.

Certificates

Page 177

CERTIFICATE of QUALIFIED PERSON

I, John Scott, of Mississauga, Ontario, do hereby certify that:

a)	I am currently employed as Fenix Engineering Manager with HMI Nickel Inc., Suite 2500, 1 Adelaide St. East, Toronto, Ontario.

b)	I am a co-author of the report titled Technical Report on an Update to the Fenix Project, Izabal Guatemala and dated November 19, 2008 (the “Technical Report”).

c)	I hold a B Sc (Eng) degree in civil engineering from London University (UK). I am a member in good standing of Professional Engineers Ontario.

I have practiced my profession continuously for over 44 years and have been involved in the mining industry for more than thirty years, primarily as project manager for consulting engineers on major projects.

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

d)	My most recent visit to the Fenix Property occurred on October 21^st. 2008 for a total of 2 days.

e)	I am responsible for the preparation of the sections of the Technical Report as detailed in Table 2.1 of the Technical Report.

f)	I am not independent of HudBay Minerals Inc. as defined in section 1.4 of the Instrument, as HMI Nickel Inc. is a wholly-owned subsidiary of HudBay Minerals Inc.

g)	I have had prior involvement with the property that is the subject of the Technical Report as Engineering Manager of the ferro-nickel project that is referenced in this Technical Report.

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

i)	As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 19th Day of November 2008.




John Scott

HudBay Minerals Inc.

Dundee Place

1 Adelaide Street East, #2501

Toronto, ON M5C 2V9

CERTIFICATE of QUALIFIED PERSON

I, Colin B. McKenzie, of 942 Maramis Court, Mississauga, ON, L5H 4G2 do hereby certify that:

(a) I am currently employed as a consultant to HudBay Minerals Inc. and was formerly Vice President, Exploration of :

Skye Resources Inc.

Ste 700 – 1111 Melville St.

Vancouver, BC V6E 3V6

now known as HMI Nickel Inc., a wholly-owned subsidiary of HudBay Minerals Inc.

(b) This certificate is in support of the report titled Technical Report on an Update to the Fenix Project, Izabal Guatemala and dated November 19, 2008 (the “Technical Report”).

respectively. I am a member (#0385) of the Association of Professional Geoscientists of Ontario, the Canadian Institute of Mining and Metallurgy and the Prospectors and Developers Association of Canada.

I have practiced my profession continuously for over 33 years and have been involved in mineral exploration and development projects in Canada and Latin America. I have had senior management roles in Voisey’s Bay Nickel Co. Ltd. and various sulphide and laterite nickel projects for Inco Ltd.

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

(d) My most recent visit to the Fenix Property occurred on December 17, 2007 for a total of 2 days, prior which I conducted approximately monthly visits of 3 to 5 days duration dating back to Skye Resources’ initiation of activities on the property in January, 2005.

(e) I am responsible for the supervision of the preparation of the sections of the Technical Report as detailed in Table 2.1 of the report.

(f) I am not independent of HudBay Minerals Inc. as defined in section 1.4 of the Instrument, as I was employed by HMI Nickel Inc. as Vice President, Exploration for a period of more than four years until March 31, 2008. HMI Nickel Inc. is a wholly-owned subsidiary of HudBay Minerals Inc.

(g) I have had prior involvement with the property that is the subject of the Technical Report as Vice-President, Exploration of HMI Nickel Inc. from February 2004 to March 2008 and as Director of Exploration, Inco Ltd from March 2002 to December 2003.

(h) I have read the Instrument and Form 43-101F1, and the Technical Report has been prepared in compliance with the Instrument and Form.

(i) As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 19th Day of November 2008.

LOGO

Colin McKenzie

HudBay Minerals Inc.

Dundee Place

1 Adelaide Street East, #2501

Toronto, ON M5C 2V9