March 2010	Exhibit 99.73
	TECHNICAL REPORT ON AN UPDATE TO THE FENIX PROJECT, IZABAL, GUATEMALA Submitted to: Hudbay Minerals Inc. Dundee Place 2501-1 Adelaide Street East Toronto, Ontario M5C 2V9 Authors: Dr. Olivier Tavchandjian P.Geo Dr. Paul Golightly P.Geo Effective Date: March 31 2010 Report Number:        09-1117-6013

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

Table of Contents

1.0	IMPORTANT NOTICE		8
2.0	SUMMARY		9
	2.1	General	9
	2.2	Geology and Exploration	10
	2.3	Mineral Resource Estimates	10
	2.4	Prospects for Economic Extraction	13
	2.5	Qualified Persons and Site Visits	13
	2.6	Reliance on Other Experts	14
	2.7	Interpretation and Conclusions	14
	2.8	Recommendations	15
3.0	INTRODUCTION		16
4.0	PROPERTY DESCRIPTION AND LOCATION		17
	4.1	General	17
	4.2	Mineral Tenure Specifically Related to Fenix Project	18
	4.3	Location of Extracción Minera Fénix Exploitation Licence	18
	4.4	Environmental Aspects	19
5.0	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY		19
	5.1	Access	19
	5.2	Climate	20
	5.3	Topography, Elevation and Vegetation	20
	5.4	Infrastructure	21
	5.5	Additional information	21
6.0	HISTORY		22
	6.1	Ownership history	22
	6.2	Exploration history and evaluation	22
	6.3	Historical Mineral Resources Estimates	23
	6.4	Production	24
7.0	GEOLOGICAL SETTING		25

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	7.1	General	25
	7.2	Regional Geology	25
	7.3	Local geology	26
8.0	DEPOSIT TYPES		28
9.0	MINERALIZATION		29
	9.1	General	29
	9.2	Description of Mineralized Zones	32
10.0	EXPLORATION		33
	10.1	General	33
	10.2	Exploration Conducted by Exmibal from 1960 to 1981	33
	10.3	Exploration Conducted by CGN from April 2005 to November 2007	35
	10.4	Service Providers	35
	10.5	Interpretation of Exploration Information	36
	10.5.1	Material Type Classification	36
	10.5.2	Laterite Classification	36
	10.5.2.1	Bedrock Classification	38
	10.5.3	Interpreted Layers for Geological Model	38
	10.5.3.1	Prediction of Bedrock Position from Iron Chemistry	38
	10.5.3.2	Generalization Classification of Layers	38
11.0	DRILLING		39
	11.1	Exmibal Drilling and Pitting Prior to 1981	39
	11.1.1	Procedures	39
	11.2	CGN Drilling from 2005	40
	11.2.1	Twin Hole Drilling	41
	11.2.2	In-Fill Drilling	41
	11.2.3	Procedures	43
	11.3	Results (from 2006 2007 Campaign)	44
12.0	SAMPLING METHOD AND APPROACH		44
	12.1	Introduction	44
	12.2	Sampling Methods Prior to 1981	45

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	12.2.1	Sample Risk Factors	45
	12.2.1.1	Mechanical Bias of Auger Assays	45
	12.2.1.2	Correction for Sampling Bias:	47
	12.3	Sampling Methods from 2005	49
	12.3.1	Sample Risk Factors	49
	12.3.2	Density Determinations	49
13.0	SAMPLE PREPARATION, ANALYSES, AND SECURITY		50
	13.1	Introduction	50
	13.2	Sample Preparation prior to Dispatch of Samples	50
	13.2.1	Exmibal Sampling prior to 1981	50
	13.2.2	CGN Sampling from 2005	50
	13.3	Statement Regarding Sample Preparation	50
	13.4	Dispatch of Samples, Sample Preparation, Assaying and Analytical Procedures	50
	13.4.1	Exmibal Programs Prior to 1981	50
	13.4.2	Calculation of Nickel and Cobalt from Cyanidation Ni+Co Assays	51
	13.4.3	Calculation of MgO and SiO2 from Historical Data	51
	13.4.4	CGN Programs from 2005	52
	13.5	Quality Control Measures	53
	13.5.1	Exmibal Programs Prior to 1981	53
	13.5.2	CGN Programs from 2005	53
	13.5.2.1	Standards	53
	13.5.2.2	Blanks	53
	13.5.2.3	Monitoring of Assays	53
	13.5.2.4	Results and Corrective Actions taken	54
	13.6	Security	54
	13.6.1	Exmibal Programs prior to 1981	54
	13.6.2	CGN Programs from 2005	55
	13.7	Statement on the Adequacy of Sample Preparation, Security and Analytical Procedures	55
14.0	DATA VERIFICATION		55
	14.1	General	55
	14.2	Historic Database	56

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	14.2.1	Background	56
	14.2.2	Verification by AMEC, 2003	56
	14.2.3	Verification 2005-06, Golightly Geoscience	56
	14.2.3.1	Initial Checking of Historic Location Information	56
	14.2.3.2	Detailed Location Checking	56
	14.3	Calculated Grade Data	58
	14.3.1	Calculated MgO and SiO2 Values	58
	14.3.2	Calculated Ni and Co Values	58
	14.4	Bulk Density	59
	14.5	Twin Drilling	60
15.0	ADJACENT PROPERTIES		60
16.0	MINERAL RESOURCE ESTIMATES		60
	16.1	Extent of the resource models	60
	16.2	Drillhole data preparation and validation	61
	16.2.1	Diamond versus Auger drilling	62
	16.2.2	Geological coding	63
	16.3	Geological Layer Modelling	65
	16.4	Sample categorization and compositing	66
	16.5	Exploratory Data Analysis	66
	16.6	Data spatial Analysis	71
	16.6.1	Unfolding	71
	16.6.2	Variography	73
	16.7	Block Model Definition	75
	16.8	Estimation Methodology	76
	16.9	Model validation	77
	16.9.1	Global statistical validation	77
	16.9.2	Visual and local statistical validation	78
	16.9.3	Smoothing assessment and correction	79
	16.9.4	Independent peer reviews	79
	16.10	Mineral Resource Reporting and Classification	80
	16.10.1	Mineralization with potential for economic extraction	80

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	16.10.2	Mineral resource classification	80
	16.10.3	Mineral Resource statement	82
	16.11	Reconciliation with Historical Production Records	82
17.0	PROSPECTS FOR ECONOMIC EXTRACTION		84
18.0	INTERPRETATION AND CONCLUSIONS		84
19.0	RECOMMENDATIONS		85
20.0	REFERENCES		87

TABLES
Table 2.1:	2008 Mineral Resource Estimates of the Fenix Project	11
Table 2.2:	Mineral Resource Estimates for Areas 212, 213, 215, 216, 217 and 251	12
Table 2.3:	Historical Mineral Resource Estimates for the Other Areas	13
Table 2.4:	Author’s Responsibilities	14
Table 6.1:	Summary of Exploration and Evaluation	22
Table 6.2:	Historical Mineral Resource estimates outside of the Feasibility Study areas	23
Table 7.1:	Average Thickness and Extent of Stratigraphic Layers	28
Table 8.1:	Mineralogy of Laterite Profiles	28
Table 8.2:	Average Electron Microprobe (EMP) Analyses of Parent Rock Minerals	29
Table 9.1:	Summary of Landforms	29
Table 9.2:	Average Grade of Saprolite (Ni>0.4%) and Estimated Serpentine Content of Bedrock	32
Table 10.1:	Summary of CGN Drilling to November, 2007	35
Table 10.2:	Summary of Material Type Classification	36
Table 10.3:	Chemical Classification of Laterite Materials	37
Table 10.4:	Chemical Classification of Laterite Materials- Exmibal Samples	37
Table 10.5:	Chemical Classification of Laterite Materials- CGN Samples	37
Table 10.6:	Chemical Classification of Parent Rock Materials	38
Table 11.1:	Historic Drilling Methods and Extent	40
Table 11-2:	Diamond Drilling by CGN from April 2005 – November 2007	43
Table 12.1:	Comparison of Historic Exmibal and CGN Diamond Drilling	47
Table 13.1:	Formulae to Calculate SiO2 from Ni and Fe2O 3	52
Table 13.2:	Formulae to Calculate MgO from Ni and Fe2O3	52
Table 14.1:	Verification Stages	55
Table 14.2:	Large-scale Validation of SiO2 and MgO Estimates	58

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Table 14.3:	Sources of Density Data	59
Table 14.4:	Regression Formulae to Calculate DBD for Drill Samples	60
Table 16.1:	Assessment of the Reclassification of Low Ni Grade Saprolite Mineralization as Bedrock	64
Table 16.2:	Number and Type of Drillholes in the Resource Modelling Database	65
Table 16.3:	Linear Correlations – Example Area 217 Z 2	67
Table 16.4:	Impact of Density Weighting in the Interpolation Process – Example of the 217z2 area - SAP Layer	71
Table 16.5:	Experimental Variogram Parameters	74
Table 16.6:	Variogram models used for the Ni Service Variables in Saprolite	75
Table 16.7:	Block Model versus Wire Frame Volume Check	76
Table 16.8:	Search Parameters Used for OK Interpolation	76
Table 16.9:	Statistics on Interpolation Results per Search Ellipsoid	77
Table 16.10:	Smoothing Assessment and Correction per Area	79
Table 16.11:	Summary of Resource Classification Criteria by Area	81
Table 16.12:	Mineral Resource estimates for Areas 212, 213, 215, 216, 217 and 251	82
Table 16.13:	Comparison of Golder’s Mineral Resource Estimates and Exmibal Historical Production Records	84
Table 19.1:	Recommended Budget	86

FIGURES
Figure 4-1:	Location of Mineral Licences	17
Figure 4-2:	Location of Extracción Minera Fénix Exploitation Licence	19
Figure 5-1:	Monthly rainfall and average temperature for 2009 at the CGN plant site	20
Figure 7-1:	Regional tectonic setting of project area	26
Figure 7-2:	Bedrock geology of the Izabal region. Using arcview shapes from MEM	27
Figure 9-1:	Longitudinal Projection and Schematic Cross Section of the Mineralized Borehole Collars in the La Gloria area Deposits	31
Figure 9-2:	Relationship of Average Nickel Ggrade of Saprolite to Olivine	33
Figure 9-3:	Relationship of Average Serpentinization of Parent rock to Pyroxene	33
Figure 10-1:	Montufar: Topography and Location of Drillholes with >1% Ni	34
Figure 10-2:	Boreholes in the Montufar 221-1 Area	34
Figure 11-1:	Location of CGN Feasibility Study Drilling	41
Figure 11-2:	Location of Historic Drilling	42
Figure 11-3:	Location of most recent CGN Drilling inn the La Gloria area	44
Figure 12-1:	Comparison of Nickel Grades Between Co-located Boreholes	45
Figure 12-2:	Comparison of Iron Grades between Co-located Boreholes	46

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Figure 12-3:	Comparison of Mineralized Thicknesses between Co-located Boreholes	46
Figure 12-4:	Comparison of Historic Auger and Twin Core Assays	47
Figure 12-5:	Nickel Grade Profiles from Twinned Exmibal and CGN Holes Holes from Areas 212, 213 & 217	48
Figure 16-1:	Extent and drilling coverage of the 2009 resource models for the 212, 213, 215, 216, 217 and 251 areas	61
Figure 16-2:	Example of excessive bottom dilution in the coding of the saprolite layer	64
Figure 16-3:	Sample length histogram: example of 212 Area	66
Figure 16-4:	%Fe2O3-%SiO2 binary diagram in the SAP layer: example of Area 217Z2	68
Figure 16-5:	%Fe2O3-%MgO binary diagram in the SAP layer: Example of Area 217Z2	68
Figure 16-6:	%SiO2-%MgO binary diagram in the SAP layer: Example of Area 217Z2	69
Figure 16-7:	%Fe2O3-%MgO binary diagram in the SAP layer: Example of Area 212	69
Figure 16-8:	Dry density-%MgO binary diagram in the SAP layer: Example of Area 212	70
Figure 16-9:	Dry density-%Fe2O3 binary diagram in the SAP layer: example of Area 212	70
Figure 16-10:	Comparison of Cartesian and unfolded location of the SAP samples in the 212 area	73
Figure 16-11:	Example of variogram map of %Ni in the saprolite layer (Area 212)	73
Figure 16-12:	Example of OK Model Visual Validation – 212 Area	78
Figure 16-13:	Drilling coverage and perimeters of mined out areas	83
APPENDICES
Appendix A	Certificates of Qualifications and Consent to File

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1.0 IMPORTANT NOTICE

HudBay Minerals Inc. (“HudBay”) commissioned Golder Associates Ltd. (“Golder”) to peer review and evaluate the Fenix resource in mid-2009 and the subsequent recommendation was to re-calculate the resource in the proposed mine area where diamond drilling has been conducted. HudBay recognised that the new resource would impact the reserve statement and commissioned Golder to update the Fenix feasibility study with the new mineral resource and mine plan and Hatch to review and re-design the power strategy and update the capital and operating cost estimates proposed for the project. In early 2010, the changes due to the new resource calculation were deemed significant enough to warrant disclosure of the revised mineral resource estimate, in advance of completion of the updated feasibility study.

This report was prepared as a National Instrument 43-101 Technical Report in accordance with Form 43-101F1 for HudBay Minerals Inc. Each qualified person (QP) listed in Table 2.4 assumes responsibility for those sections or areas of this report that are referenced opposite their name. None of the QPs, however, accepts any responsibility or liability for the sections or areas of this report that were prepared by other QPs.

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.

The QPs have, in the preparation of this report, relied upon certain data, reports, opinions and statements provided to the QPs by CGN and certain other parties as referenced in Table 2.4 of this report. The QPs have conducted reasonable due diligence on these data and reports and have no reason to believe that they cannot be used for the estimation and the reporting of mineral resource estimates as presented in this technical report.

This technical report updates a technical report that was filed by HudBay on September 22, 2008.

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2.0 SUMMARY

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 Niquel S.A. (“CGN”). CGN is a Guatemalan company, the shares of which are 98.2% owned by HMI Nickel Inc., a subsidiary of HudBay and 1.8% by the Guatemalan Government. CGN was formerly named 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.

The initial Fenix feasibility study was completed on October 16, 2006 (the “2006 Feasibility Study”), and the related technical report was issued in September 2006 and filed at www.sedar.com on November 17, 2006 (the “2006 Technical Report”). The 2006 Technical Report was later updated to incorporate exploration results up to June, 2007, revised mineral resource and mineral reserve estimates and revised power supply and bulk material transportation strategies. Following the acquisition of Skye Resources by HudBay Minerals, the technical report that was prepared in 2007 was re-filed by HudBay in November 2008 (the “2008 Technical Report”) with the purpose of disclosing that the hydrometallurgical process contemplated for treatment of the limonite material was no longer considered to possess the immediate potential for economic extraction.

In late 2008, HudBay put the construction of the Fenix project on suspension due to the onset of the economic crisis at the time. The project was put on care and maintenance and some limited work continued on site including the upgrading of the road from the town of Rio Dulce to El Estor and the continuance of social programs in the region. HudBay in mid-2009 commissioned Golder to peer review and evaluate the resource methodology and the mine plan. Included in the findings was recognition that the mine plan and the resource methodology had not been optimized. Hatch was engaged in late 2009 to scope the potential possibilities for sourcing electrical power for the project in recognition that the methodology proposed in the 2006 Feasibility Study was not optimal for HudBay’s circumstances.

In late 2009 Golder was engaged to re-calculate the resource model and re-design the mine plan. Hatch was engaged in early 2010 to update the 2006 Feasibility Study with current operating and capital cost estimates, review and re-design the power strategy and account for costs incurred on the project to date.

The limonite resource stated in the 2008 Technical Report is absent from this report because HudBay does not consider that this material has the potential to be economically viable under conventional processing technology.

In early 2010, the changes due to the new resource calculation were deemed significant enough to warrant disclosure of the revised mineral resource estimate in advance of completion of the updated feasibility study.

The update of the 2006 Feasibility Study is underway.

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2.2 Geology and Exploration

The nickel deposits are lateritic weathering profiles that have formed on peridotites thrust into place during the early Tertiary. The most important deposits are located in the area subject to the Fenix exploitation license largely on 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. Additional important deposits are found on lower terraces and spurs on Sierra de las Minas in the Montufar licence at elevations of 300 – 450 metres elevation.

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.

CGN’s core drilling was conducted in three phases:

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

• A program of in-fill drilling in 215, 216, 217 and 251 to achieve a general drill spacing of 50 metres or less.

• A preliminary program of twin drilling and a geostatistical cross in the high grade 221-1 area of Montufar.

The cut-off date for the drilling used in the Hudbay mineral resource estimates is 31 December, 2009.

As of February 2010 HudBay is diamond drilling the proposed mine areas of 212 and 213 which are currently delineated using the auger drill method. The proposed program is 8,000-10,000 metres with four drills operating and is intended to be completed in May of 2010.

2.3 Mineral Resource Estimates

The Mineral Resource estimates reported in the 2008 Technical Report are summarised in Table 2.1 below. These estimates are presented separately for the areas covered by the 2006 Feasibility Study and for the other areas. While Mineral Resource estimates in the 2006 Feasibility Study areas were estimated based on 3D modeling of the geology and grade distribution and last updated by Snowden in 2007, the Mineral Resource estimates for the other areas were originally prepared for all areas using Inco’s 2-D LES (Laterite Estimation System) method. These historical estimates were audited and validated by AMEC (2003) and by Golightly Geoscience (2006) and were deemed to be realistic in their stated purpose of defining a resource for a Preliminary Assessment. These deposits will require re-estimation in 3-D format to enable mine planning. This remodelling will be completed within the scope of the present update of the Feasibility Study to be completed by mid-2010.

The saprolite Mineral Resource estimates presented in Table 2.1 correspond to the portion of the mineralization to be processed by pyrometallurgy to produce a traditional Ferro-Nickel product while the limonite Mineral Resource estimates were deemed in 2008 to be amenable to an acid leach type of mineral processing.

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Table 2.1: 2008 Mineral Resource Estimates of the Fenix Project

	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

In second half of 2009, Golder was engaged by HudBay to update the Life of Mine (“LOM”) Plan supporting the 2006 Feasibility Study previously done by Snowden consultants. While the LOM plan is being finalised, Golder has completed an update of the mineral resource estimates for the areas 212, 213, 215, 216, 217 and 251 prior to conversion to mineral reserves. These updated mineral resource estimates are summarised in Table 2.2 below.

In order to report the mineral resource estimate, the potential for economic extraction of the mineralization was defined as follows:

• Only the saprolite (including transition) portion of the mineralization with a suitable chemistry was included in the updated mineral resource (note that internal limonite dilution has already been incorporated in the geological model). Suitability of the chemistry was tested on large areas based on the assumption that small individual blocks with extreme chemistry could be successfully blended by operating different part of the deposits simultaneously and by an expansion of the existing stacking-reclaiming facility.

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• After compiling grade-tonnage curves, a cut-off grade of 1.6% Ni in areas 212, 213z1, 213z2, 217z1 and 251 and 1.5% Ni in areas 215, 215inf, 216, 217z2 and 217z3 was deemed to constitute the optimum compromise between resource utilization and minimum acceptable economic return on investment.

• A minimum of two consecutive vertical blocks above cut-off (2 metres) was applied as an additional constraint in order to remove thin and isolated layers. A maximum of 2 metres of vertical internal waste within the “ore layer” was also applied based on the assumption than any larger waste inclusion could be selectively mined out and disposed without entering the plant feed stream.

• On the measured and indicated mineral resource estimates, a Lerch-Grossman algorithm was applied in order to eliminate the blocks which would necessitate an unrealistic amount of incremental stripping or mine development to be successfully extracted. The constraints applied through the Lerch-Grossman and minimum ore thickness criteria result in a very small portion of the blocks above cut-off to be excluded from the mineral resource estimate. In the opinion of Golder, approximately 97% of the mineralization grading above cut-off presents potential for economic extraction.

Table 2.2: Mineral Resource Estimates for Areas 212, 213, 215, 216, 217 and 251

Classification	Area	% Ni cut-off	000’ Dry Tonnes	% Ni
Measured	212	1.6	4,793	2.09
	213z1	1.6	724	2.19
	217z1 East	1.6	2,218	2.09
	Total		7,735	2.09
Indicated	213z2	1.6	1.68	1.94
	215	1.5	3,346	1.85
	216	1.5	5,895	1.75
	217z1 West	1.6	537	1.82
	217z2	1.5	9,441	1.86
	251	1.6	9,067	1.97
	Total		28,454	1.87
Measured + Indicated	Total		36,190	1.92
Inferred	217z3	1.5	3,900	1.8
	Total		9,700	1.8

It is important to note that the 2010 “Other Areas Estimate” are excluded from the revised estimate presented in Table 2.2 and presented separately in Table 2.3. The restatement of the mineral resource estimates for these other areas is necessary in order to avoid duplication since some of these zones, i.e. 216 area and part of 217 area have been covered by recent core drilling and are now included in the areas considered for the update of the Feasibility Study.

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Table 2.3: Historical Mineral Resource Estimates for the Other Areas

Classification	000' Tonnes	% Ni
Measured	8,713	1.79
Indicated	26,074	1.82
Measured+lndicated	34,787	1.81
Inferred	26,118	1.75

Prior mineral resource estimates presented in Table 2.1 include estimates for the Limonite portion of the profile which was deemed by Skye Resources to be amenable to a pressure acid leaching type of mineral processing. HudBay has conducted further investigations without being able to confirm the economic potential of this processing alternative. As a result, no limonite mineral resource estimates are being reported in 2010.

The significant reduction in the updated saprolite mineral resources amenable to pyrometallurgical processing, as reported in Table 2.2, is largely related to the use of a higher cut-off in 2009. This increase in cut-off was deemed necessary in order to support the potential for economic extraction. Previous mineral resource estimates were based on the assumption of stockpiling very large quantity of low grade mineralization which in Golder’s opinion would have proven difficult to execute in a safe and economic manner. When compared to the tonnage and grade reported by Snowden for the same areas in the 2007 Technical Report at a 1.6% Ni cut-off, the Golder mineral resource estimates are actually showing more tonnage of high grade mineralization.

In the other areas (Table 2.3), a reduction is observed in the total inferred mineral resource estimates and is due to the inclusion of areas 216 and part of 217 in Table 2.2 and to the removal of the Mantos areas, for which mineral rights have been released by CGN in 2008.

2.4 Prospects for Economic Extraction

A test to justify the reasonable prospects for economic extraction was conducted. This test included updating the 2006 Feasibility Study cash flow model supporting the historic reserve number with the newly calculated resource number outlined in this technical report, and applying reasonable mining recovery and dilution numbers to the Measured and Indicated resources. As anticipated the project’s realised return and cash flow after capital and operating costs were positive, justifying the determination that the resource contained in Table 2.2 provides for reasonable prospects for economic extraction. Due to the ongoing work by Golder and Hatch, HudBay felt guidance on specific costs was best left for completion of the updated Feasibility Study at which time HudBay will commission the publication of a new Technical Report and declaration of reserve.

2.5 Qualified Persons and Site Visits

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

• Dr. Paul Golightly, P.Geo has had an association with the Fenix Project since mid 2004 and visited the site on many occasions.

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• Dr. Olivier Tavchandjian, P.Geo has had an association with the Fenix Project since late 2009 and visited the site in August 2009.

Table 2.4: Author’s Responsibilities

Report Section	Q.P.	Input from Others
1: Important Notice	Tavchandjian
2: Summary	Tavchandjian	Golightly, Aceituno, Meagher
3: Introduction	Tavchandjian	Golightly, Aceituno, Meagher
4: Property Description and Location	Golightly	Aceituno, Meagher
5: Accessibility, Climate, Local Resources, Infrastructure and Physiography	Golightly
6: History	Golightly	C.B.McKenzie, RC Osborne
7: Geological Setting	Golightly
8: Deposit Types	Golightly
9: Mineralization	Golightly
10: Exploration	Golightly
11: Drilling	Golightly
12: Sampling Method and Approach	Golightly
13: Sample Preparation, Analyses and Security	Golightly
14: Data Verification	Golightly	Snowden, Skye
15: Adjacent Properties	Golightly
16: Mineral Resource Estimates	Tavchandjian
17: Prospects for Economic Evaluation	Tavchandjian	Hatch, Meagher
18: Interpretations and Conclusions	Tavchandjian	Golightly
19: Recommendations	Tavchandjian	Golightly, Aceituno, Meagher
21: References	N/A
22: Illustrations	N/A
22: Dates and Signatures	N/A
23: Certificates	N/A

2.6 Reliance on Other Experts

In preparation of this report, a series of reports and opinions of third party experts (see references) have been reviewed and accepted by Golder as reasonable basis for some of the assumptions used in the reporting of the mineral resource estimates.

2.7 Interpretation and Conclusions

The Fenix nickel deposits are typical wet-tropical climate laterites and occur as surficial deposits on topographic 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 Golightly Geoscience’s opinion, the core sampling, sample preparation, analytical procedures and security for the CGN drilling 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.

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Golder 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 Golder’s opinion the compiled data are acceptable for mineral resource estimation.

Golder prepared saprolite Mineral Resource estimates for Ni, Co, Fe₂O₃ , MgO, SiO₂ and dry density for Areas 212, 213, 215, 216, 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, cut-off grade and potential production rate into account.

The Golder mineral resource estimates were peer reviewed and reconciled with mine production records.

2.8 Recommendations

The following program is recommended by Golightly Geoscience and Golder 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; and

• 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.

• Test of ground penetrating radar to interpolate the contacts of the saprolite orezone between drill holes.

Programs of additional exploration drilling, both twin and in-fill, followed by updated mineral resource estimates are recommended in the following areas to expand and delineate resource in areas not considered in the forthcoming Feasibility Study and would cost approximately US$8.6 million.

• Diamond drilling the auger supported resource areas of 212 and 213 that are to be incorporated into the mine plan in the updated Feasibility Study.

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• Area 218 deposits.

• Montufar property, the Cristina area 221 & 222 deposits.

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

• Reconnaissance of favourable landforms in the Chulac area west of Amate and in the north west part of area 255.

• Upgrading test comprising:

• Excavation (pits or trenches) of bulk saprolite and size distribution of nickel and in selected sites in pyroxene-rich harzburgite hosted areas of relatively low serpentinization; and

• Size analysis of selected archival diamond drill core.

3.0 INTRODUCTION

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 CGN. CGN is a Guatemalan company, the shares of which are owned approximately 98.2% by HudBay through subsidiaries, and 1.8% by the Guatemalan Government. 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.

The geology of the Fenix nickel laterite deposit is well understood. 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. Historic Exmibal and new sampling data obtained by CGN were used in the July 2006 resource estimates. CGN conducted further drilling and sampling activities in 2005-07 including new drilling in the 216 area. Dr. Paul Golightly of Golightly Geoscience has 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. The cut-off date for the drilling used by Golder to update the mineral resource estimates is 31 December, 2009.

The mineral resource estimates have been revised by Golder for the following areas included in the revision of the Feasibility Study presently under completion: 212, 213, 215, 216, 217 and 251. The resource modeling completed by Golder in second half of 2009 was based on best industry practices including data validation with emphasis on the comparison of auger and diamond drilling and on geological coding, 3D interpretation of geological domains, sample compositing and unfolding, statistical and geostatistical analysis by layer, grade interpolation, use of service variables, model validation involving smoothing assessment and correction, resource classification and reporting. The resource models were validated through reconciliation with production results and by peer reviews.

The mineral resource estimates for the other areas are prepared based on the historical Inco LES estimates verified by AMEC (2003) and Golightly (2006) with adjustments to reflect the release of mineral

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rights for the Mantos areas and the inclusion of the 216 and part of 217 areas into the on-going update to the Feasibility Study. These historical estimates will be updated by Golder using the same methodology used for the Feasibility Study areas when the 2010 update of the Feasibility Study.

In early 2010, the results of the new resource calculation was deemed significant enough to warrant disclosure of the revised mineral resource estimate in advance of completion of the updated feasibility study.

4.0 PROPERTY DESCRIPTION AND LOCATION

4.1 General

The Fenix Project is located in two areas near the northwest and southeast shores of Lake Izabal in the Department of Izabal in eastern Guatemala.

The project is approximately 160 kilometres northeast of Guatemala City, and 70 kilometres from the Caribbean coast. The larger of the two areas, which is covered by Extracción Minera Fénix, an Exploitation Licence (formerly part of a slightly larger exploration licence area called ‘Niquegua Norte’), is 4 kilometres west of the town of El Estor and contains the nickel laterite deposits. (Figure 4.1)

Figure 4-1: Location of Mineral Licenses

Note: Areas covered by the Niquegua Montufar II Exploration (granted 13 December, 2004) and Extracción Minera Fénix, Exploitation license (granted April, 2006) are shown in Figure 4.1. Licences currently held by CGN in relation to named places near the licences and to topography as shown on a composite Landsat 7 image obtained from USGS website, https://zulu.ssc.nasa.gov/mrsid.

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4.2 Mineral Tenure Specifically Related to Fenix Project

HudBay has represented that the following information related to mineral tenure is a current description. Until December 2004 the Fenix Project property covered the Niquegua mining exploitation concession (entitling mining for nickel, cobalt, iron and chromium) granted to CGN on 14 August, 1965 for a 40 year term beginning 26 August, 1965 and ending 25 August, 2005. This concession consisted of three sections: (i) Montufar (covering 35.99 square kilometres), (ii) Chulac-EI Estor (covering 234.40 square kilometres), and (iii) Sexa (covering 114.97 square kilometres). Chulac-EI Estor, and the smaller, adjacent Amate and Xaan exploitation concessions, were located north of the Process Plant in the foothills of the Sierra de Santa Cruz. Sexa was approximately 33 kilometres to the northeast, and Montúfar about 40 kilometres southeast across the lake near the town of Mariscos.

Effective 13 December, 2004, the Amate and Xaan exploitation concessions and the Chulac-EI Estor section of the Niquegua exploitation concession were converted, with some minor changes in boundaries, into an exploration licence, Niquegua Norte, covering approximately 259.2065 square kilometres located near the northwest shore of Lake Izabal. The remaining sections of the Niquegua exploitation concession (namely, the Montúfar and Sexa sections) were also reorganized, also with some changes in boundaries, whereby the Sexa section was surrendered and the Montúfar section was converted into a second exploration licence, Niquegua Montúfar II, covering approximately 64.5292 square kilometres located south of Lake Izabal. These new exploration licences entitle exploration for nickel, cobalt, iron, chromium and magnesium.

The Chichipate exploitation licence (permitting limestone production and covering 1.24 square kilometres) which expires in 2010 was retained.

In April 2006, the Ministry of Energy & Mines issued an exploitation licence called the “Extracción Minera Fénix, Exploitation Licence” (the “Fenix Exploitation Licencer) which converted substantially all of the Niquegua Norte exploration licence into a 25 year exploitation licence renewable for a further 25 years, covering an area of 247.9978 square kilometres. The exploitation licence allows mining of nickel, cobalt, iron, chromium and magnesium to take place on the licensed area.

The three year term for the Niquegua Montúfar II exploration licence ended on December 12, 2007 but prior to its expiration, the first extension was requested and granted for a period of two years. The request, based on prior deposit evaluations, surrendered 50% from the original license, and 32,265 square kilometres were kept.

To maintain the Fénix Exploitation Licence and Niquegua Montúfar II exploration licence, CGN must pay annual surface taxes of approximately CND$61,000 in advance in January of each year, and carry out a work program on the licence areas. The surface rights to the site of the Process Plant, to most of La Gloria Areas 210-217 inclusive, including the pit and the plant site, and to Manto 4 Area 251, all of which are located within the Fenix Exploitation Licence area, are owned by CGN or its affiliates. Surface rights and houses in a disused mine town site on the west side of El Estor are also owned by CGN or its affiliates.

4.3 Location of Extracción Minera Fénix Exploitation Licence

The location of all known mineralized zones relative to the outside property boundaries is shown in Figure 4.2.

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Figure 4-2: Location of Extracción Minera Fénix Exploitation License

Note: Red dots & clouds in the Fenix Project area are historic borehole collars.

4.4 Environmental Aspects

The principal environmental legislation applicable to the Fenix Project is Guatemala’s (1) Environmental Law – Decree 68-86 (Ley de Protectión y Mejoramiento del Medio Ambiente), and (2) MARN Regulation 23-2003. The Fenix Project is also subject to a number of additional Guatemalan environmental and health and safety regulations concerning land and water transportation of equipment, supplies and products, fuel storage, water treatment and discharge, noise levels, air pollution, solid waste, stewardship of forestry resources, and archaeological sites.

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Access

The Fenix Exploitation Licence area is accessible by road from Puerto Barrios on the Gulf of Honduras. It is paved from Puerto Barrios to a few kilometres west of Rio Dulce (79 kilometres) and gravel surfaced on the remaining route from Rio Dulce to El Estor (a distance of 43 kilometres) and thence to Coban (160 kilometres). The road link between El Estor and Rio Dulce is relatively recent and is not shown on many small regional scale maps. There is regularly scheduled public bus service linking El Estor to Rio Dulce and to Coban. The Montúfar II Exploration Licence area is accessible by paved Highway CA-9 from Puerto Barrios and is crossed by the paved access road to Mariscos from Highway CA-9.

An airfield maintained by CGN which is suitable for small aircraft is located at Las Dantas about one kilometre east of the process plant. During the exploration and production stages of the Exmibal project, access to the El Estor area was supported by ferry across Lake Izabal from Mariscos.

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5.2 Climate

The climate of the region is a typical wet tropical environment with approximately 2.5 metres of annual rain falling regionally mostly between the months of May and October. A weather station established at the CGN plant site recorded a total rainfall 1.27 metre for 2009.

The total rainfall at the plant is approximately a metre less than at a station at Panzos, 50 kilometres west southwest of the CGN Plant site at an elevation of 520 metres. The dominant wind direction throughout the year is from the southeast. However, north by northeast trade winds are more frequent than average during the rainy season. The plant site is probably in the rain shadow of the Sierra de la Cruz and the plant site observations may underestimate the rainfall expected at higher elevations in the mine. It is also believed that 2008 and 2009 were drier than normal.

The site lies in a region where tropical storms are a possibility. The plant site has not experienced a major hurricane since it was established.

Figure 5-1: Monthly rainfall and average temperature for 2009 at the CGN plant site

5.3 Topography, Elevation and Vegetation

The deposits comprise portions of lateritic weathering profiles that are thickest on the ridges and plateaux of the Sierra de Santa Cruz north of Lake Izabal and the Sierra de Las Minas on the south side of the lake. The most important deposits are largely confined to NE or SE trending terraces and spurs on the south flanks of these hills at elevations between 330 and 800 metres. The total relief is 1,000 metres. The highest topographic points, the site lying in area 215 north of the Lake and area 223 to the south, are each sites of telephone communication towers with dirt access roads reaching them from the south.

Unlike many nickel laterites, the vegetation is diverse tropical forest. A surficial cover layer containing a significant amount of non-ultramafic derived material has sufficient phosphorus to support normal vegetation. As a result, undeveloped areas of laterite outside of areas 212 and 213 are largely covered by rainforest. Exceptions are the 218, 219 & 251 areas in Fenix and the Cristina area at Montufar which have been partially cleared for subsistence farming. Semi-open pine forests in the Montufar licence and in a limited southwest part of area 217 are restricted to the mid to lower south slopes and are probably, at least in part, the result of lower rainfall on the leeward slope rather than sterile laterite soil.

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5.4 Infrastructure

The metallurgical and infrastructure facilities, which were constructed starting in 1974 and commissioned in 1977, consist of:

• Run-of-Mine ore dump pocket and crusher.

• Crushed ore conveyor and stockpile.

• Crushed ore kiln dryer (45 metres long by 3.66 metres in diameter) and associated dust handling equipment.

• Dry ore storage building (capacity 40,000 tonnes).

• Dry ore kiln roaster/reducer (100 metres long by 5.5 metres in diameter) and associated off-gas handling equipment.

• Ore transfer system and associated off-gas handling equipment.

• A 45,000 kilovolt-ampere three electrode electric furnace (18 metres in diameter).

• Slag granulating equipment with a capacity of 1,650 tonnes per day.

• Nickel matte handling equipment.

• Converters, granulators, product handling and associated off-gas handling equipment.

• A 61.3 megawatt heavy oil fired steam turbine power plant.

• Maintenance shops for the maintenance of both the mine fleet and metallurgical plant.

• Metallurgical laboratories for the mine and process plant control.

• Town site, hospital, school, water and sewage, facilities to house and support nonlocal staff.

• Dock and barging facilities for receipt of Bunker C oil and the shipment of nickel product.

• Bulk oil storage facilities at both the plant site and at Livingston on the Caribbean coast.

5.5 Additional information

At the time of writing this technical report, Golder and other consultants are updating the 2008 Feasibility Study for the re-establishment of mining and treatment operations of saprolite. Aspects of sufficiency of surface rights for mining operations, the availability and sources of power, water, mining personnel, potential tailings storage areas, potential waste disposal areas, and processing plant sites will be addressed in the aforementioned study and are expected to be disclosed in subsequent reports.

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6.0 HISTORY

6.1 Ownership history

The owner of the Fenix property is CGN (formerly ‘Exmibal’). In August, 2008, HudBay acquired Skye Resources Inc., which had previously acquired Inco’s 70% share of Exmibal, the original owner of the exploration, mining and processing project. Further ownership history details are provided in Skye, 2005. Annual Information Form (“Skye AIF”), available at www.sedar.com.

6.2 Exploration history and evaluation

The major exploration efforts on the Exmibal property were conducted by Inco prior to Skye’s involvement (Harju 1979). Inco’s exploration activities consisted mainly of geological mapping, drilling and pitting on various prospects. Historical details are provided in Hatch 2005 and the Skye AIF. The history is summarized in Table 6.1.

Table 6.1: Summary of Exploration and Evaluation

YEAR	NICKEL LATERITE EXPLORATION AND EVALUATION
1955	Nickel was first recognized in samples from the farm of J. M. Montufar south of Lake Izabal.
1957-60	Preliminary exploration of the Lake Izabal Montufar area was carried out by Hanna Mining Company1.
1960	Formation of Exmibal (originally 80% Inco 20% Hanna and subsequently shareholders have changed). Exmibal’s name was changed to ‘CGN’ in January 2005.
1961-70	Exmibal carried out reconnaissance exploration through to advanced project development on numerous prospects. The prospects were sampled with manual and power drills, and test pits.
1965	Exmibal was granted a 40 year mining concession and exploration entered a “development” stage.
1973	In agreement with the Guatemalan government, Exmibal started to build and operate a mine and smelter near El Estor.
1977-80	Exmibal mined about 1.08 million dry metric tonnes of saprolite from Area 212 (La Gloria) and produced approximately 32.9 million pounds of nickel as nickel matte. The plant was shut down in Q4, 1980 as part of a worldwide cut-back in production by Inco in response to low nickel prices and high petroleum prices during a global recession. Since shut down, the plant has been kept on care and maintenance.
1980-2005	No further exploration or mining activity took place prior to April, 2005.
1993	Inco, Sumitomo Metal Mining Co. Ltd. and Tokyo Nickel Company Ltd. completed a feasibility study on restarting operations. Inco estimated a resource using their Laterite Estimation System (LES) 2.2 proprietary 2-D software. This estimation showed 14,297,000 dmt of “mineable” resource at a grade of 1.91% nickel in their “20 year reserve”.
1998	Inco completed a further in house re-evaluation using their LES.

¹ Hanna were the operators of the Nickel Mountain Laterite Mine in Riddle, Oregon and the developers of the Cerro Matoso Mine in Colombia,

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YEAR	NICKEL LATERITE EXPLORATION AND EVALUATION
2004	On 15 December, 2004, Skye Resources Inc. conditionally acquired Inco’s 70% share of Exmibal. AMEC examined the Inco saprolite resource estimates for their 20 year reserve in the La Gloria area of the Fenix licence and reclassified them in terms consistent with the requirements of NI 43-101. Golightly Geoscience re-classified the remaining saprolite and limonite resource estimates in terms consistent with the requirements of NI 43-101 and reported them in Hatch, 2005.
2005	The 40 year mining license was renewed. Exmibal was reorganized as Compania Guatemalteca de Niquel ‘CGN’.
2005 - 2007	CGN drilled 62,981 metres in 2,479 diamond drill holes to produce a mining reserve consistent with NI 43-101 and published a feasibility study.
2008	A further 2,174 metres were drilled until exploration activities were suspended in February in the early stages of the global financial crisis. In August, 2008, HudBay Minerals completed their acquistion of Skye Resources.

6.3 Historical Mineral Resources Estimates

Golightly Geoscience reported historical Mineral Resource estimates for the deposits that occur in CGN’s licences outside of the Feasibility Study area. The estimates were originally prepared for all areas using Inco’s 2-D LES method and are realistic in their stated purpose of defining a resource for a Preliminary Assessment. These deposits will require re-estimation in 3-D format to enable mine planning to proceed. Table 6.2 below summarises these historical mineral resource estimates for which classification was reviewed and validated by AMEC (2003) and Golightly (2006).

Table 6.2: Historical Mineral Resource estimates outside of the Feasibility Study areas.

Source of data	Classification	Area	000’ Tonnes	% Ni	~hole spacing (m)
Golightly 2006	Measured	221-1	8,713	1.79	25
TOTAL MEASURED			8,713	1.79
Golightly 2006	Indicated	221-2,3	3,173	1.79	100
Golightly 2006	Indicated	222-1	883	1.86	50
Golightly 2006	Indicated	222-2	5,096	1.87	100
Golightly 2006	Indicated	222-3	2,256	1.77	100
Golightly 2006	Indicated	222-4	2,161	1.74	75
Amec 2003	Indicated	218-1	2,620	1.95	100
Golightly 2006	Indicated	218-7	2,734	1.75	50
Golightly 2006	Indicated	219-NE	2,850	1.78	100
Golightly 2006	Indicated	219-NW	810	1.93	100
Golightly 2006	Indicated	210	142	1.81	50
Golightly 2006	Indicated	211	685	1.85	50

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Source of data	Classification	Area	000’ Tonnes	% Ni	~hole spacing (m)
Golightly 2006	Indicated	214	310	1.72	50
Golightly 2006	Indicated	245	800	1.79	70
Golightly 2006	Indicated	260	1,554	1.81	100
TOTAL INDICATED			26,074	1.82
Amec 2003	Inferred	218-2	6,860	1.82	Amec 2003
Amec 2003	Inferred	218-3	1,255	1.71	Amec 2003
Amec 2003	Inferred	218-4	402	1.56	Amec 2003
Amec 2003	Inferred	218-5	2,517	1.59	Amec 2003
Amec 2003	Inferred	218-5	950	1.68	Amec 2003
Amec 2003	Inferred	219	7,335	1.73	Amec 2003
Amec 2003	Inferred	260	6,019	1.81	Amec 2003
Amec 2003	Inferred	224-2	780	1.63	Amec 2003
TOTAL INFERRED			26,118	1.75

6.4 Production

Inco, through CGN (then known as ‘Exmibal’), constructed and operated the mine and process plant on the Fenix Project area from 1977 to 1980 (Sopko 1979). The operation was planned for a minimum 20 years production but a combination of high energy costs and low nickel prices forced the shut down of the operation. The production over a period of approximately 40 months is summarized as follows:

• Overburden and waste removal was carried out by bulldozers and scrapers, ore removal by two excavators and a smaller backhoe for bottom recovery, haulage was by eight 35 tonne ore trucks.

• Annual mine production was nominally 604,000 dry tonnes of saprolite.

• Mine production totaled approximately 1,079,300 tonnes grading 2.12% of nickel and which yielded approximately 14,940 tonnes of nickel.

The production came mostly from the 212 deposit. Pre-production stripping was also carried out on the western part of 212 and the 213 deposit. Ore grade material removed from 213 while constructing the main haul road was stockpiled on the stripped deposit.

Ore was dumped, crushed and stacked at the plant site prior to being fed into the process plant. A front end loader was used to feed the process plant from the wet ore stockpile. At the time of the operation’s shutdown about 50,000 tonnes remained on the wet ore stockpile as well as at least 20,000 tonnes of in-process material in dry ore storage. In addition, small stockpiles of run-of-mine material were stored in the mine area.

During the three years of operation, the plant processed approximately 1,079,300 tonnes of ore (less stockpiles) to produce approximately 14,940 tonnes of nickel as nickel matte. After an extended commissioning and start-up period, the process plant operated over a period of nine months at an

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annualized rate of ore processing of 575,000 dry tonnes per year to produce 9,832 tonnes of nickel per year. This was 87% of its design capacity of 11,300 tonnes per year. A 90 day performance test was undertaken from May to August, 1980. During this period the process plant met its design capacity. This extended operating period confirmed the grade and tonnage of ore against the mineral resource forecasts and demonstrated its amenability to smelting and nickel production.

7.0 GEOLOGICAL SETTING

7.1 General

The regional, local and property geological descriptions are provided by Golightly Geoscience in Hatch 2005. Extracts are provided below for completeness.

7.2 Regional Geology

The nickel deposits are lateritic weathering profiles that have formed on peridotites thrust into place during the early Tertiary.

The ultramafic rocks of the Lake Izabal region lie in the vicinity of a major transform fault zone stretching from Guatemala through Jamaica to the southern peninsula of Hispaniola. The project area (red rectangle in Figure 7.1) lies adjacent to two major faults that separate the North American and Caribbean plates. The transform fault, which is thought to have a left lateral displacement of about 1,300 kilometres, comprises two branches, the Polochic fault through Lake Izabal and the parallel Motagua fault in the large valley south of the Sierra de Las Minas. These faults curve from their trans-Caribbean ENE trend into a WNW trend across the highlands of Guatemala to parallel the Americas trench off the Pacific coast. These faults are active and the ENE trending 777 km x 70 km zone extending from west of the property off-shore adjacent to Honduras has experienced seven magnitude 6 earthquakes since 1973 including a magnitude 7.6 quake in 1976 centred very close to the Montufar property.

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Figure 7-1: Regional tectonic setting of project area.

7.3 Local geology

The geological framework in relation to the Fenix properties is shown in the regional geology map, Figure 7.2.

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Figure 7-2: Bedrock geology of the Izabal region. Using arcview shapes from MEM.

PROPERTY GEOLOGY

The laterite deposits have accumulated on relatively gently sloping terrain on topographic plateaus and terraces.

Some of the terraces in the La Gloria area are clearly separated from the higher terrain to the north by marked topographic lineaments that probably represent faults. It has been suggested that the terraces hosting the deposits may represent small down-dropped blocks flanking major faults related to a graben in a pull-apart basin along the Polochic fault, now occupied by Lake Izabal. (See also Figure 9-1).

The laterite profile of economic interest is developed on ultramafic bedrock. The ultramafic rocks are predominantly harzburgite (approximately 80% olivine + 20% orthopyroxene). In the mine area there are minor (<5%) small, several metre-scale bodies of dunite, pyroxenite and gabbro. There are significant variations in the proportion of harzburgite and the pyroxene contents of the harzburgites between different deposits. Pyroxene contents, according to the chemistry of the laterite profiles, may be as high as 50%. The ultramafic rocks are strongly jointed, cut by zones of breccia with serpentinite matrix and show 70% to 100% serpentinization. Olivine has been selectively serpentinized so that lower degrees of serpentinization are seen in the harzburgites.

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Within the laterite profile above the bedrock there are four units or layers that can be interpreted from drill samples, and mapped from exposures in the mine at La Gloria.

These are (from top-down) referenced as cover (layer 1), limonite (layer 2), transition (layer 3) and saprolite (layer 4). A fifth unit, bedrock (layer 5), comprises unaltered to weakly altered peridotite and serpentinite.

The cover sequence consists of transported aluminous limonite in part derived from gabbro and from volcanic ash-fall tuffs, presumably derived from volcanoes in adjacent regions. Transported limonite is commonplace in lateritic terrains but the high degree of exotic contamination in the cover is unusual.

The limonite and cover zones may be absent due to erosion and are thick and continuous only on gently sloping terrain. Saprolite may be absent below limonite either due to incomplete drill penetration or due to transport of limonite onto unweathered rock. Average thicknesses for the layers are given in Table 7.1. The cover is more extensive than the limonite.

Table 7.1: Average Thickness and Extent of Stratigraphic Layers

	Fenix		Montufar
Unit	Thickness (m)	Holes Drilled	Thickness (m)	Holes Drilled
Cover	4.3	54%	0.4	57%
Limonite	4.7	39%	1.9	44%
Transition	2.6	47%	3.6	88%
Saprolite	6.1	83%	6.4	87%

Note: Based on historical Exmibal drill records. Holes drilled gives percentage that had at least one sample of that unit.

8.0 DEPOSIT TYPES

The nickel deposits of the Lake Izabal region are typical examples of nickel laterite deposits formed in seasonally wet tropical climate on partially serpentinized peridotites. The nickel in such deposits derives ultimately from the igneous olivine (0.3% Ni in olivine) and the serpentine that replaces it.

The deposits comprise a saprolite zone of partially decomposed peridotite corestones overlain by a zone of in situ limonite. With progressive weathering nickel is leached out of the limonite zone and transition zone and enriched in altered bedrock serpentine in the underlying saprolite. Typical features of the units (or layers) are given in Table 8.1. Typically nickel is present in solid-solution in serpentine, in goethite and in manganese-, cobalt- oxides.

Table 8.1: Mineralogy of Laterite Profiles

Unit or layer	Mineralogy	Structure
Cover	Goethite + kaolinite ± quartz ± gibbsite (± traces of zircon ± xenotime, YPO4)	Massive to bedded
Limonite	Goethite ± Mn, Co oxides ± gibbsite ± quartz ±chromite	Relict bedrock textures changing upward to massive fine-grained

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Unit or layer	Mineralogy	Structure
Transitional Saprolite	Ni in serpentine + Mn, Co oxides ± quartz +goethite + magnetite ±chromite	Fe rich Mg poor saprolite crusts surround Ni enriched saprolite corestones
Saprolite	Ni in serpentine+goethite+magnetite Mg-rich, Fe-poor	Saprolite crusts surround corestones of weakly mineralized or un-enriched peridotite.
Parent Rock	Forsteritic olivine + serpentine+magnetite (±chromite ± pyroxenes ± talc)	Joint blocks, serpentinite breccia

Limonite, transitional saprolite (transition) and saprolite are typically amenable to mining in shallow pits. Transition and saprolite are typically processed by pyrometallurgical methods to produce nickel and iron. Alternatively transition may be processed along with limonite by hydrometallurgical methods to produce nickel and cobalt.

Table 8.2: Average Electron Microprobe (EMP) Analyses of Parent Rock Minerals.

Mineral	Source	Al2O3		CaO	FeO	MgO	MnO	NiO	SiO2	TiO2	V2O5	ZnO
Ortho pyroxene	EMP	2.82	0.57	1.19	5.75	31.48	0.01	0.05	52.86	0.00	0.00	0.00
Clino pyroxene	EMP	2.50	0.49	22.38	5.28	14.67	0.02	0.13	47.44	0.10	0.01	0.00
Chromite	EMP	27.65	40.41	0.00	19.26	11.20	0.28	0.07	0.03	0.05	0.24	0.11
Chromite	Calculated	21.00	57.00	0.00	19.26	11.20	0.28	0.07	0.03	0.05	0.24	0.11
Olivine	Calculated				8.4	44 5	0.13	0.38	46.65

Note: Average electron microprobe (EMP) analyses of parent rock minerals in metallurgical test samples from the existing pit in area 212. Second chromite and Source: SGS Lakefield.

9.0 MINERALIZATION

9.1 General

The mineralized zones are best described in relation to the landforms in each area. The landform layer thicknesses and elevations are summarized in Table 9.1 for each area for which a resource was estimated by Golder. In the Fenix Exploitation License landforms at elevations below about 450 metres have almost no cover (layer 1) and little limonite (Layer 2) or transition (Layer 3). The exception is area 217, sectors 4, 5, and 7. Montufar has little or no cover.

Table 9.1: Summary of Landforms

Deposit	Land Form	Remarks Alias	Holes	Elevation (m)	Area (km2)	Saprolite (m)	Transition (m)	Limonite (m)	Cover (m)	Status
Fenix- La Gloria Area
210	Spur		50	264	0.19	8.24	1.60	0.42	0.74
211	Spur		37	278	0.12	11.32	3.32	0.76	0.46
212	Terrace	Stripped Partially Mined	1830	433	0.82	7.82	2.06	5.76	1.05	FS
213-1	Spur		66	105	0.06	7.00	1.38	0.21	0.00	FS
213-2	Spur		38	134	0.11	4.50	1.45	0.00	0.03	FS
213-3	Slope		140	123	0.09	6.84	2.30	0.69	0.06	FS
215-1	Terrace		181	779	1.23	6.36	2.30	8.36	2.48	FS
215-2	Hill		23	824	0.50	5.09	2.43	5.74	1.65	FS
215-3	Terrace		27	535	0.59	4.56	3.26	3.96	3.07	FS

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Deposit	Land Form	Remarks Alias	Holes	Elevation (m)	Area (km2)	Saprolite (m)	Transition (m)	Limonite (m)	Cover (m)	Status
216	Terrace		121	602	2.00	6.43	2.02	5.24	3.75	FS
217-1	Terrace		231	473	0.87	6.62	2.36	10.61	3.55	FS
217-2	Hill		145	471	0.32	9.69	2.69	4.61	0.64	FS
217-4	Spur		161	364	0.10	7.21	1.88	4.94	0.58	FS
217-5	Spur		252	386	0.28	6.42	1.91	9.26	1.62	FS
218-1	Plateau		68	733	0.55	5.66	2.68	4.68	1.84
218-2	Plateau		56	742	0.58	6.95	3.13	7.82	2.75
218-3	Plateau		14	793	0.23	4.50	2.07	7.21	3.36
218-S	Hill on Plateau		6	412	0.13	3.33	0.83	7.50	0.50
219-E	Plateau		16	799	0.25	4.69	1.69	7.81	2.44
219-M	Plateau		21	819	0.70	1.67	2.33	6.29	1.71
219-W	Plateau		18	699	0.37	2.72	1.72	5.78	2.94
South west Fenix
251-1,2a	Terrace	Manto 4	218	1066	2.95	6.94	2.12	6.56	1.09	FS
2512b	Ridge	Manto 4	38	691	3.2725	3.1579	2.3684	3.4211	2.1842
260	Terrace	Amate	48	282	1.10	8.58	3.50	1.67	0.88
Montufar
221-1E	Terraced spur	Cristina	195	321	0.28	9.1	4.1	0.5	0.6
221-1w	Terraced spur	Cristina	324	309	0.37	9.4	3.7	0.4	0.5	CGN
221-1U	Terrace	Cristina	391	437	0.67	7.7	4.3	2.2	1.5
221-2	Summit Ridge	Cristina	133	500	0.89	5.0	3.9	3.9	2.4
222-1	Spur	Cristina	130	492	0.88	6.9	3.4	3.0	1.7
222-2	Summit Ridge	Cristina	97	610	1.01	5.4	2.9	5.1	3.4
223-1L	Terrace	Cristina	6	315	0.07	11.8	1.2	1.5	1.0
223-1U	Spur	Cristina	91	464	0.72	4.9	4.7	0.7	1.0
223-2	Summit Ridge	Cristina	131	730	1.31	3.1	4.0	4.0	3.0
224-1	Slope	Trincheras	8	250	0.15	4.5	2.6	0.1	0.0
224-2	Summit Ridge	Trincheras	24	314	0.59	6.0	2.5	2.0	1.1
224-3	Plateau	Trincheras	96	361	0.78	2.9	3.8	3.5	1.6

Laterite characteristics and status of deposits. Based on historic boreholes and pits. Status: FS = Golder FS resource estimate. CGN = limited DDH testing.

The mineralized layers, saprolite through limonite, range in average thickness from 10.2 m at 213-1 to a maximum average thickness of 22.5 m at 217-1. The largest continuous laterite landform occurs at 251, which covers an area of approximately two million square metres. Three landforms, 215, 217-1 and 212 occupy areas of almost one million square metres.

Deposits not marked as FS status require confirmatory drilling and resource calculations.

Terraces and spurs in the La Gloria Area and to the Southwest at Manto 4 and Amate appear to be blocks down-dropped along faults. At La Gloria many of the terraces are inclined south west or north east as show in Figure 9-1. Evidence from core samples such as sand-filled crevices and local occurrence of

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saprolite overlying limonite and cover sediments indicate these faults have been active during laterite formation. Other deposits at Fenix northwest of the line of faults are formed on hills or edges of an extensive plateau landform. Locally on the 215-1 terrace the in situ laterite profile is overlain by saprolite, some of it derived from an unidentified felsic rock, which must have come into place by landsliding from the adjacent slope leading up to the highest hill to the northwest.

Terraces at Montufar are less well defined and generally horizontal features formed on topographic spurs emanating southeast from the north east trending ridgeline. Mineralization zones extend more continuously down the relatively shallow slopes between the terraces than they do at Fenix.

Figure 9-1: Longitudinal Projection and Schematic Cross Section

of the Mineralized Borehole Collars in the La Gloria area Deposits

Note: Black symbols are holes with >1% Ni. The terraces, T, and spurs, S, hosting the mineralization are inclined mostly to the west. This section is close to what would be seen from the distant south-east from the surface of Lake Izabal. The faults are schematic but supported by evidence in drill core and in the old Exmibal pit which is on the terrace at intermediate elevation. Z+Ni Column and Z+Fe Column give the elevation of surface required to provide the Ni and Fe in the laterite profile. The Ni and Fe required are roughly equal on the plateau, P. but z+Ni is progressively higher than z+Fe going to lower and lower elevations on the terraces and spurs due to limonite erosion and more intensive laterization. The cross-section is located along the white line in the location map.

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9.2 Description of Mineralized Zones

The nickel grade of the saprolite depends upon the topographic setting and the amount of rock that was decomposed to form the laterite profile and on the grade of nickel in the parent rock. The latter depends upon the olivine content of the parent rock. This makes a significant variation of the nickel grade of the saprolite zone. Al2O3 in the laterite derives largely from pyroxene, which has very low nickel (approximately 0.01%). As shown in Table 9.2 below, the ratio of the insoluble oxides Al2O3/(Fe2O3+Cr2O3+Al2O3) gives a good measure of the pyroxene content of the parent rock and through this to the average Ni grade and serpentine content of the deposit.

Table 9.2: Average Grade of Saprolite (Ni>0.4%) and Estimated Serpentine Content of Bedrock

Area	Ni	Al2O3	Cr2O3	Fe2O3	Serpentine	Al2O3/ (Fe2O3+Al2O3+Cr2O3)	Ni/ Fe2O3
	Weight Percent
212	2.05	1.22	0.66	13.23	101%	8.1%	0.155
217.2	1.46	1.97	0.74	15.17	97%	11.0%	0.096
213.3	2.27	2.30	0.77	16.99	100%	11.5%	0.134
221.1	1.61	3.95	1.49	28.28	82%	11.7%	0.057
251	1.52	2.05	0.70	14.24	96%	12.1%	0.107
216	1.20	2.19	0.76	15.12	93%	12.1%	0.080
217.1	1.43	2.11	0.73	14.37	93%	12.3%	0.099
215.2	1.18	2.09	0.72	14.09	88%	12.3%	0.084
217.3	1.50	2.37	0.77	15.78	97%	12.5%	0.095
213.1	2.13	2.88	0.81	18.27	94%	13.1%	0.117
215.1	1.46	2.48	0.77	15.19	81%	13.5%	0.096
213.2	0.86	2.87	0.75	16.65	94%	14.1%	0.052
215	1.06	2.67	0.76	15.15	79%	14.4%	0.070
215.3	1.12	2.91	0.77	15.56	82%	15.1%	0.072
Orthopyroxene	0.04	2.82	0.57	6.39	0%	28.8%	0.006

Note: Average grade of saprolite (Ni>0.4%) and estimated serpentine content of bedrock in the areas drilled by CGN. Serpentine estimated very approximately as 100 * LOI/13%. Values > 100% may reflect presence of brucite Mg(OH)2 in dunitic bedrocks.

As shown in Figures 9-2 & 9-3, the average Ni grade of saprolite varies inversely with the proportion of pyroxene as anticipated.

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Figure 9-2: Relationship of Average Nickel Ggrade of Saprolite to Olivine

Note: Relationship of average nickel grade of saprolite to olivine content of parent rock is indicated by the negative correlation of nickel with aluminum provided by pyroxene.

Figure 9-3: Relationship of Average Serpentinization of Parent rock to Pyroxene

Note: Relationship of average serpentinization of parent rock to pyroxene content. Low serpentine in saprolite ores is related to mechanical sampling bias and to potential upgradability of ore by barren corestone rejection.

10.0 EXPLORATION

10.1 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 CGN on behalf of Skye, is identified in Sections 10.3 below and areas explored for the 2006 Feasibility Study are shown in Figure 11.1.

10.2 Exploration Conducted by Exmibal from 1960 to 1981

Areas explored by Exmibal are shown in Figures 10-1, 10-2 and 11-2. 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 Montufar 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 Montufar 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.

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Figure 10-1: Montufar: Topography and Location of Drillholes with >1% Ni

Figure 10-2: Boreholes in the Montufar 221-1 Area.

Note: 53 CGN diamond drill holes are located here on one of the topographical lowest terraces in Montufar. Legend as in previous figure.

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10.3 Exploration Conducted by CGN from April 2005 to November 2007

From April 2005 to November 2007, CGN’s program involved 2,620 HQ diamond core boreholes totalling 65,034 metres and providing 60,286 assay samples². The areas drilled are shown in Figure 11-1. Drill data up to January 2006 were included in the feasibility study but additional drilling continued to November 2007 and is compiled with the feasibility study drilling. 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 metre grid spacing in deposits 215-1, 217-1 & 2 and 251. Six cross patterns of 12.5 -18 metre spaced holes were drilled in four of these deposits.

• 53 holes including additional twins and a cross pattern were drilled in area 221-1 in the Montufar exploration license.

The holes include 8 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 November, 2007

	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	8551	In-fill to 75m	2007
221-1	53	1525	Twin & Crosses	2007
Chichipate	4	121	Limestone	2006
Total	2,624	65,155

10.4 Service Providers

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.

² Excluding core duplicates.

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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.

• Geografia y Foresteria, S.A. (GYFSA)

• Century Systems Technologies Inc.

Drilling on the Fenix Exploitation License was performed using track mounted diamond drill rigs by the Guatemalan subsidiary of the Boart Longyear Group. Drilling at Montufar using a portable diamond drill was done by Sococo / Palo Verde.

10.5 Interpretation of Exploration Information

10.5.1 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

10.5.2 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.

The chemical formulae used are provided in Table 10.3, Table 10.4 and Table 10.5.

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Table 10.3: Chemical Classification of Laterite Materials

Major Unit	Layer	Formula
Limonitic		(Ni-0.1) <0.0005x(Fe2O3-10)2
	Cover	Ni < 5.42 - 0.08 Fe2O3
	Limonite	Ni > 5.42 - 0.08 Fe2O 3
Saprolite		(Ni-0.1 )>0.0005x(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) <0.0005x(Fe2O3-10)2
	Cover	Ni < 5.42 - 0.08 Fe2O3
	Limonite	Ni > 5.42 - 0.08 Fe2O 3
Saprolitic		(Ni-0.1)>0.0005x(Fe2O3-10)2
	Transition	Ni< -2.4+0.14 Fe2O 3
	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*Fe 2O3) and Quartz>10
	Bedrock	((Ni<.4)and SiO2<(40÷35)*MgO
	Chromitic	(Cr2O3+0.31 )>1.5+0.0452*Fe2O3
	Aluminous	Al2O3>0.3*Fe 2O3
	Gabbroic	(TiO2÷Fe2O 3) > 0.045*(Al2O3÷Fe2O3 )
	Fresh Gabbro	CaO>1
Limonitic:		(-0.14+Ni) < 0.00052*(Fe2O3-10)2
	Limonite	(Ni> 5.416666667+-0.119149747*Fe
	Cover:	(Ni<5.416666667+-0.119149747*Fe
	Gabbroic Cover	(TiO2÷Fe2O3)>0.045*(Al 2O3-Fe2O3 )
	Anomalous Cover:	MgO>2
	Calcic Cover:	CaO>1
	Sand Cover:	Cr2O3> 0.01720121613*Fe2O31.3075

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-AxMgOⁿ)x(SiO₂-100)÷((SiO₂-100)-MgOxnxAxMgO^(n-1))

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Where A = 2.044417877 and n=0.8348996748

10.5.2.1 Bedrock Classification

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

Fe₂ O₃, Cr₂O ₃, 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 Parent Rock Materials

Rock Type	Discriminant Criteria
Gabbro	(Al2O3>0.3*Fe 2O3) and (Cr2O3 <0.047*Fe2O3)
Anomalous	(Al2O3>0.3*Fe 2O3) and (Cr2O3 >0.047*Fe2O3)
Clinopyroxenite	(CaO>2)
Harzburgite 1	(Al2O3 < 0.3*Fe2O3) and Al2O3 > 0.18*Fe2O3
Harzburgite 2	(Al2O3< 0.18*Fe2O3) and Al2O3 > 0.12*Fe2O3
Dunitic Harzburgite	(Al2O3 < 0.12*Fe2O3) and Al2O3 > 0.03*Fe2O3
Dunite	Al2O3 <0.03*Fe2O3

10.5.3 Interpreted Layers for Geological Model

10.5.3.1 Prediction of Bedrock Position from Iron Chemistry

Twin core holes on average penetrated nearly 10 metres 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 classified as ‘Inferred’ Resources.

10.5.3.2 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.

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 limonite or transition material. The cover usually lies in 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 metre 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 5 metre compositing it was classified manually, with particular attention to mis-classification near layer boundaries.

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

Golder reviewed the layer boundaries and revised the layer 4/5 boundary upwards to include material with Ni<0.7% in layer 5.

11.0 DRILLING

11.1 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 Montufar Cristina area 221.

All boreholes and pits were vertical. The pit openings generally were at least one metre square. Research pits centred on boreholes showed boreholes to deviate from vertical by as much as 10% of depth.

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.

11.1.1 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.

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 metre by 1.5 metre in area and up to approximately 40 metres in depth. All pits were normally 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. Some pits at Montufar appear to not have been back-filled.

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Table 11.1: Historic Drilling Methods and Extent

Area	Holes	Length	Samples	Hole Type	Hole Numbering 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

11.2 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

11.2.1 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 14.0 – Data Verification.

11.2.2 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.

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Since January 2006, 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 metres in order to upgrade the classification of the resource. To identify new mineral resources, historic drilling on a lower terrace and an upper ridge in 215 was filled into a 50 metre grid. Historic drilling on area 216 was filled in to a 75 metre grid. At Montufar, to begin confirming the historical resource, some twin holes and a “geostatistical” cross for a total of 53 holes were drilled on the lowest terrace of the 221 Area shown in Figure 11-2.

Figure 11-2: Location of Historic Drilling.

Drilling statistics are listed in Table 11-2.

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Table 11-2: Diamond Drilling by CGN from April 2005 – November 2007

Project # AREA		HOLES	Length Assayed	Hole Depth	Assays	FROM*	UNTIL
212	La Gloria	72	2460.69	2532.5	2405	25-May-05	15-Feb-06
213	La Gloria	13	294.86	299.59	312	25-May-05	17-Jan-06
217.1	La Gloria	609	17466.97	17665.4	16033	27-Jun-05	26-Sep-06
251	Manto 4	704	13504.65	13613.3	12945	15-Jul-05	6-Jun-06
215.1	La Gloria	333	7356.35	7369.25	7150	24-Oct-05	13-Feb-07
217.2	La Gloria	143	3318	3319.8	3083	5-Dec-05	11-Oct-06
217.3	La Gloria	121	3214.45	3244.45	2973	22-Jan-06	22-Jan-07
215.3	La Gloria	97	2723.1	2730.6	2434	26-Jan-06	6-Jun-07
216	La Gloria	290	8544.2	8550.7	7763	26-Jan-06	24-Apr-07
215.2	La Gloria	98	2482.8	2487.3	2293	11-Oct-06	13-Feb-07
221.1	Montufar	53	1523.85	1525.35	1364	1-Jun-07	1-Sep-07
215	La Gloria	87	1694.65	1695.95	1531	13-Oct-07	27-Nov-07
Total		2620	64584.6	65034.2	60286

Note: Dates are of assays received including QC-re-assays. Assay count excludes duplicate core samples. Table does not include limestone drilling for sample blanks at Chichipate.

The program was suspended in November, 2007.

Currently since January 2010, a program of 254 diamond drill holes on a 50 metre grid is in progress to re-confirm the historical and twin drilling, 68 in area 213 and the remainder in area 212 for an expected 7,000 metres. Five (5) holes are being drilled to provide geotechnical data.

11.2.3 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.

• 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.

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• 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.

11.3 Results (from 2006 2007 Campaign)

Diamond drilling in 2006 and 2007 post the 2006 Feasibility Study, and included condemnation drilling in areas 215, 217, 216 and 251, contributed to increasing the mineral resources at Fenix. The principal result for this campaign was the inclusion of area 216 in the resource estimate (Skye news release February 13, 2008). Snowden estimated the resource at a 1.25%Ni cut-off 8.1 million tonnes of 1.52 %Ni in the indicated category and 1.6 million tonnes at 1.46%Ni in the inferred category for the saprolite material. Snowden also calculated the resource contained in limonite at a 1%Ni cut-off with 5.1 million tones at 1.03%Ni in the measured category, 0.7 million tonnes at 0.96%Ni in the indicated category and 1.1 million tonnes at 0.99%Ni in the inferred category.

Figure 11-3: Location of most recent CGN Drilling inn the La Gloria area.

12.0 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.

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12.2 Sampling Methods Prior to 1981

Most samples were obtained by track mounted auger drills, with selected sites re-sampled by pits.

12.2.1 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.

12.2.1.1 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 metre grid auger drilling patterns were “twinned” by a new CGN diamond borehole located about 1 to 2 metres 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 Figures 12-1 to 12-3.

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

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

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

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 core stones 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.

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A high nickel bias in the historic holes as a function of iron is a maximum at 30% Fe₂O ₃ and decreases upwards and downwards from there. As seen in Figure 12-4 a high bias in the lower part of the saprolite is reversed in the upper part and in transition zone material.

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

12.2.1.2 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

	Depth Relative to Top of Saprolite		CGN Fe2O3	CGN Ni	Exmibal Fe2O3	Exmibal Ni	DFe2O3 Exmibal - CGN	DNi Exmibal - CGN
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

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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
2171&2	-10	0	57.34	1.49	55.83	1.52	-1.50	0.03
2171&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
221-1	-10	0	49.49	0.85	50.73	0.81	1.24	-0.04
221-1	0	10	25.05	1.74	30.28	1.89	5.23	0.16
221-1	10	20	9.50	0.52	13.50	0.76	3.99	0.24

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.

Comparisons of nickel and iron grades in equivalent parts of the laterite profile are illustrated in Figures 12-1 to 12-3. Average grades from the twin holes and corresponding historic holes are plotted by depth above and below the limonite boundary in Figures 12-4 and 12-5.

Figure 12-5: Nickel Grade Profiles from Twinned Exmibal

and CGN Holes Holes from Areas 212, 213 & 217

The variation in nickel was found to be localized and inconsistent and no corrections were applied to the historic nickel assays.

At Montufar area 221-1, as shown in the last 3 line entries of Table 12-1 there is a larger difference in nickel, than (Exmibal about 0.3% high), seen elsewhere. The area tested at Montufar is very localized and the apparent bias can not be confidently extended to the deposit as a whole.

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The bias in iron was found to be significant and consistent as is illustrated in Figures 12-4 and 12-5, thus supporting a decision to adjust some of the historic iron assays.

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[(Fe₂O₃ -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 2005-2007 for the resource estimates was conducted by diamond core drilling.

12.3.1 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.

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 misidentification of the bedrock and thus the full extent of the laterite profile may remain untested.

12.3.2 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 14.4.

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13.0 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 kilogram 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.

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.

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).

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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.

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 SiO_z 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-Fe₂O₃ 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).

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Table 13.1: Formulae to Calculate SiO₂ from Ni and Fe₂O₃

Equations for SiO2
Saprolitic: SiO2 = a0 + a1Fe2O3 + a2 Fe2O32 + a3 Fe2O3 3 + a4 ln(Ni/Fe2O3)
Limonite & Cover: = a0 + a1 Fe2O 3
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

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 Fe₂O ₃

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 105° 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, K₂O, TiO₂, MnO, P₂O₅ , V₂O₅ and LOI.

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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.

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. The final 190 external check assays were performed by ALS-Chemex in Vancouver using the same method during the autumn of 2009.

13.5.2.1 Standards

Six 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. In December 2005, a second pair of standards and in mid 2006, a third pair of standards, each pair comprising 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 third set of additional standards has 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 DH Logger database. Golightly Geoscience’s program searched for and identified individual standards.

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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.

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,

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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.

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.

14.0 DATA VERIFICATION

14.1 General

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.

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 - 2007	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.

14.2 Historic Database

14.2.1 Background

Assay data were routinely digitized by key punch entry with verification in the 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 from earlier digital compilations and from paper logs during the early 1990’s.

14.2.2 Verification by AMEC, 2003

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

14.2.3 Verification 2005-06, Golightly Geoscience

14.2.3.1 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 9. 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.

14.2.3.2 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. A small number could not be re-located even though the old bulldozer-cut drill-pads were generally recognizable.

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In rare cases the holes may have been filled by rain washed material and in rare instances, it is possible that the wooden marker post that Exmibal had inserted into the collars, sprouted and grew into a tree. 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; the holes 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 metre in-fill drilling program was adjusted and reclassified to Inferred in the current study.

Field verification of historic borehole locations in the 221-1 area at Montufar was started in 2007. This was complicated by two factors. 1) The database collar co-ordinates were in local grid co-ordinates only; an accurate location of the historic grid had never been done 2) More holes, both pits and boreholes, were found in the field than were present in the database; investigation of microfiche records revealed that boreholes drilled by Hanna in the late 1950’s prior to the formation of Exmibal had not been digitized by INCO although they can be seen on old paper collar-maps. Eventually the holes found were matched as closely as possible to the database using a computerized location matching program.

A 1961 Hunting Survey photogrammetric contour map was available for identifying elevation anomalies and the location of a farm track on the historic collar maps also appeared to correspond to the available access road.

The agreement is more approximate than at La Gloria. It is thought that this may be the result of the use of more traditional surveying methods in the 1960’s as compared to the radar theodolite method used in the 1970’s.

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14.3 Calculated Grade Data

14.3.1 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.

Table 14.2 below gives a comparison of overall SiO₂ and MgO between diamond drilling data and calculated values based on the corrected Fe₂O₃ 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.

14.3.2 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 .

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14.4 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).

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	12,7483
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).

³ 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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Table 14.4: Regression Formulae to Calculate DBD for Drill Samples

Regression for Bulk Specific Volume (SV) m³/t

	a0	a1	a2
Saprolitic Fe2O3 < 30%	0	0.0645838	-0.0009304
Saprolitic Fe2O3 > 30% (includes transition)	1.086
Limonite	0.910
Cover	0.780
Specific Volume = a0 + a1 Fe2O 3 + a2 Fe2O3 2

Note: DBD = 1/SV.

14.5 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 10.5.3.1).

• 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.

15.0 ADJACENT PROPERTIES

Most of Anfield Ventures Ltd.’s Maya Niquel Project lies adjacent to the west side of the Fenix Exploitation Licence area. Using a cut-off grade of 1% Ni they report combined measured & indicated resources of 15.2 DMT at 1.53% Ni and 20.9 DMT at 1.43% Ni are classed as inferred. The property is currently under exploration.

16.0 MINERAL RESOURCE ESTIMATES

16.1 Extent of the resource models

The extent and drill-hole coverage of the resource models constructed in 2009 by Golder for the Fenix project are illustrated on Figure 16.1.

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Figure 16-1: Extent and drilling coverage of the 2009 resource models

for the 212, 213, 215, 216, 217 and 251 areas

The Fenix mineral resource modeling covered 10 different areas separated on the basis of landforms, drilling methods and drilling density, extending from 224.970E to 247,945E, from 1,712,175N to 1,721,050N and from elevation 50 metres to 853 metres. The coordinates used for modeling were reduced by 200,000 in Eastings and by 1,700,000 in Northings for ease of data processing purposes. The resource modeling work was completed by Olivier Tavchandjian, P.Geo and was peer reviewed by Greg Greenough, P.Geo and Kevin Palmer, P.Geo. The 3D resource block models were constructed utilizing Datamine Studio 2, version 2.1.1547.0 in single precision with the exception of area 215inf which was modelled utilizing Datamine Studio 2, version 2.1.1547.0 in extended (double) precision.

The mineral resource estimates for the other deposits within the HudBay Minerals license including areas 210, 211, 214, 218, 219, 221-224 (Montufar), 245 and 260 (Amate) were not updated in 2009 as they were not included in the scope of the resource modeling work conducted by Golder. The mineral resource estimates for these areas remain based on historical estimates presented in Table 6.2.

16.2 Drillhole data preparation and validation

The Fenix project drillhole database was provided to Golder by CGN, a Guatemalan company owned 98.2% by HudBay. With the exception of the 216 area, the checks conducted on the drillhole database used by Golder had been documented extensively by Golightly Geosciences and Snowden in the 2007 technical and Feasibility study reports. The drillhole data for the 216 area has been more recently validated by Golightly Geoscience in 2009.

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In addition, Golder has conducted a systematic series of data validation checks including:

• A validation through a comparison of 100% of the assays and lithological coding of the excel databases provided by CGN to Golder and the Datamine databases used for resource modeling

• A systematic validation that missing assay or density data occur either in the cover or bedrock zones or represent lost core samples. An exemption was 34 5 metre sections of production holes in saprolite which were not assayed. The chemistry and density of these samples was left as unknown. These intervals all located in Auger holes of the 212 area.

• A verification that all the drill-hole samples start at a zero distance from the collar of the hole.

• Absence of collocated holes (collar located within 1m). Eight pairs of collocated holes were found and in this instance the oldest was removed. This usually occurs where hand Auger holes would have been replaced by power Auger holes or when a diamond drill holes were re-drilled due to poor recoveries in the initial hole.

• No gaps exist in the drill-holes.

• No layer coding was missing.

• No layers were duplicated.

The resource modeling process used by Golder uses a series of Datamine scripts and macro functions which prevent the user from completing the model if all these checks have not all been successfully completed.

Overall the geology of the Fenix deposits is well understood and a lot of effort was devoted by Golightly Geosciences to validate and maximise the use of all the historical and existing information in addition to validating the most recent diamond drilling exploration programs. Further investigations were conducted by Golder on specific issues which were identified as being potentially of concern early in the modelling process. These investigations consisted in:

• a comparison of the diamond and Auger drilling data; and

• some adjustments to the geological layer coding.

16.2.1 Diamond versus Auger drilling

Several sections of the reports produced by Snowden and Golightly Geosciences between 2005 and 2007 confirm the expected bias in the sampling obtained from Auger drilling compared to sampling from core drilling where both techniques have been used. Biases occur both in terms of profile penetration and with respect to chemistry and density measurements. Golightly Geosciences developed a series of correction factors in order to allow the combination of core drilling and Auger drilling data to support previous mineral resource and mineral reserve estimates.

Golder has conducted follow-up investigation in order to confirm that biases observed in the Auger drilling had been adequately and reliability corrected (Tavchandjian, 2009a). This work indicated a high likelihood that biases still exist both in the coding of the geological layers and the chemistry of these layers even after the corrections have been applied to the Auger data. Overall all sources of bias are interpreted by Golder as under-estimating either the tonnage or the grade of the saprolite mineralization and are deemed to produce conservative mineral resource estimates.

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Based on the results of this comparative study, it was decided by Golder to remove all the Auger data where diamond drilling spacing is 100 metres or less including areas 215, 216 and 217z2. The density of diamond drilling in areas 212, 213 and 217z1 is insufficient to remove the Auger data and the resource estimates that will be generated from the Golder resource models for these areas can be seen as prudent. For the areas 217z3 and 215inf, only sparse Auger data (drill spacing is usually 200 metres or higher) is available and these areas can only support inferred mineral resource estimates. For area 251, core drilling constitutes approximately 75% of the drilling but removing the Auger holes would give a drill spacing in excess of 100 metres in these areas and as a result it was decided by Golder to maintain these holes in this area, acknowledging that they may introduce a slight conservative bias in the resource estimates.

Golightly Geosciences developed an extension formula in order to combine Auger and diamond drilling data. This extension zone with simulated samples and unknown chemistry has not been retained by Golder for resource modeling purposes. In areas where Auger drilling is still used for resource modeling, it was deemed that using the chemistry from real assay data into the deeper “extended areas” was not desirable since this approach ignores the close correlation between chemical composition and vertical position within the saprolite layer. This decision results in a reduction of tonnage of the inferred mineral resource estimates in several areas where “simulated data” had been used by Snowden in 2007.

16.2.2 Geological coding

Inconsistencies were noted by Golder in the resource modeling approach used in 2007 in the handling of the boundary between the transition and saprolite layers. For some of the steps in the modeling process the boundary was treated as a hard one (e.g. interpolation) while it was considered as a soft boundary for other steps (e.g. variogram calculation, resource reporting). In order to maintain a more consistent modeling approach and to better reflect the difficulty to separate the very thin transition layer from the saprolitic material, Golder has grouped the two layers in all the steps of the modeling process.

In the visual validation of the geological coding assigned by Golighlty Geosciences, it was found that in many instances, a large portion of the saprolite layer consisted of very low Ni grading mineralization with very low Fe2O3 and high MgO content. Although this material can still be considered saprolitic, its chemistry and Ni grade indicate a very low degree of alteration typical of hard saprolite and can often be separated by a hard contact in drillhole profiles (Figure 16.2). The inclusion of this large quantity of low grade mineralization was deemed to have an undue significant diluting impact on the resource model and a methodology was developed to efficiently remove the vast majority of this low grade bottom dilution. The saprolitic material reclassified as bedrock (or hard saprolite) can represent in some zones up to 15% of the total volume previously coded as saprolite with an average composition showing usually 0.5% or less in Ni, 10% Fe2O3 and in excess of 40% in SiO2 and 30% in MgO (O. Tavchandjian, 2009a).

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Figure 16-2: Example of excessive bottom dilution in the coding of the saprolite layer

Any change to the geological coding from the Golightly Geosciences coding was validated by Marco Aceituno, Chief Geologist for CGN who had extensive direct field and logging supervision experience over the past 5 years with the Fenix Project.

Table 16.1 summarises for each area, the impact on the average thickness and average grade of the saprolite layer of the reclassification implemented by Golder.

Table 16.1: Assessment of the Reclassification of Low Ni Grade Saprolite

Mineralization as Bedrock

Area	% Reduction in Saprolite Thickness	% Increase in Ni Grade
215	8%	5%
215 inf	8%	5%
212	8%	7%
213z1	15%	12%
213z2	15%	12%
216	10%	6%
217z1	13%	10%
217z2	12%	9%
217z3	11%	7%
251	9%	6%
Average	10%	7%

The Fenix project drillhole databases validated and utilised for the resource modeling of the 10 selected areas consisted of a total of 5,273 holes. The number and type of drilling is detailed for each area in Table 16.2. The drill-hole data was provided to Golder by CGN on August 11, 2009 in multiple comma-delimited text files, with separate files for the collars, lithologies and assays. All data was in the UTEM grid. The comma-delimited text files were imported into Datamine Studio using a customised procedure which combines the various individual files into a single de-surveyed 3D drill-hole file.

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Table 16.2: Number and Type of Drillholes in the Resource Modelling Database

Area	Number of Holes	Type
215	310	Core only
215 inf	299	Auger only
212	1749	Dominantly Auger
213z1	203	Dominantly Auger
213z2	38	Dominantly Auger
216	290	Core only
217z1	505	Dominantly Auger
217z2	845	Core only
217z3	110	Dominantly Auger
251	924	Dominantly Core
Total	5273

16.3 Geological Layer Modelling

The contacts between the following geological layers (from top to bottom of the profile) were interpreted as digital terrain models (DTMs) by direct triangulation between actual drilled intersections only:

• Cover (COV)

• Limonite (LIM)

• Transition + Saprolite (SAP)

• Bedrock (BRK)

This method is useful in layered deposits for minimizing interpretational bias that could occur if the 3D model of the layers was to be generated through linking of sectional or plan interpretations. These DTMs were extended beyond drilling coverage by a distance corresponding usually to one third or one half of the average drill-hole spacing over each area. The outer limits of the mineralization were ‘squared off’ rather than tapered. This extrapolated volume beyond the last mineralized drill holes represents a minor portion of the total mineralization volume.

The top of the COV layer represents the topography as interpreted from drill-hole collar. The LIDAR topographic surface provided by CGN has indicated significant local discrepancies between the LIDAR and the collar surfaces. The resource modeling method used by Golder for Ni laterites uses an “unfolding algorithm” which interpolates grade between samples based on their relative vertical position within the profile of each layer. As a result, these topographic discrepancies have no impact on the resource estimates. After the creation of the resource models, the columns of blocks in the models were moved up and down to match the LIDAR topography which is deemed to be more accurate and more reliable.

No attempt was made to model separately inclusion of material locally present in another layer. These inclusions of LIM, SAP or BRK in another layer do not have sufficient lateral continuity to be modelled separately with confidence and as a result are included as “internal dilution”. This approach is consistent with the definition developed by Golightly Geosciences and used by Snowden in the construction of previous resource models. The 3D interpretation of these contacts was visually checked for each zone in order to ensure that there were no intersections between layers.

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16.4 Sample categorization and compositing

The top and bottom DTMs were used to select the samples belonging to each geological layer Only those sample intervals with their centroid lying between the two surfaces were captured.

Histograms of the selected sample lengths were used to help determine an optimum composite length for grade estimation. The vast majority of the samples were originally assayed over a 1 metre length in all the areas and for all the layers as illustrated on Figure 16.3 for the 212 Area. As a result, a composite length of 1 metre was chosen as most appropriate in all cases. Compositing was conducted from the top to the bottom of the profile and separately for each layer. Since density is calculated and strongly related to chemistry, it was deemed essential to weight by density during the compositing process

Figure 16-3: Sample length histogram: example of 212 Area

A verification was conducted to ensure that the total sample length before and after compositing remained the same by layer. These composite files constitute the basis for the exploratory and spatial data analysis and for the interpolation process.

16.5 Exploratory Data Analysis

Primary univariate statistics as well as histograms of the composited data for each area were plotted and reviewed. All the statistics were computed for %Ni, %Co, %Fe₂O₃, %SiO₂ and %MgO. For zones 213z1 and 213z2, no meaningful histograms could be constructed for the COV and LIM layers due to the small number of samples available for these geological units. Generally all the histograms constructed for the Fenix areas display typical characteristics of Ni laterrite deposits for each layer. The %Ni histogram is usually close to a normal distribution while the other elements are skewed. The occurrence of outliers in %Fe₂O₃, %SiO₂ and %MgO in each layer is usually related to the inclusion of heterogeous material in the layer profile, e.g. a small inclusion of saprolite material in the limonite layer. For this reason these outliers were not capped as they represent true measurements. Grade capping on the major oxides would result in biased estimates of chemistry.

The matrices of linear correlations confirm the expected correlations between the major oxides %Ni and %Co within each layer as illustrated in Table 16.3 as an example for the 217Z2 Area. All the other areas present very similar matrices of correlations between the various elements. The 217Z2 area was selected since the modeling database in this zone is entirely constituted of diamond drill holes with actual

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assaying for all the elements. In areas where Auger drilling constitute a significant portion of the modeling database, linear correlations between Ni, Co and Fe2O3 with SiO2, and MgO would be artificially enhanced due to the fact that Si02 and MgO which were not assayed have been calculated using a regression from the other elements by Golightly Geosciences. Also for older Auger data, the %Co and %Ni could have been estimated from their combined assayed value.

Table 16.3: Linear Correlations – Example Area 217 Z 2

217z2	Element	%Ni	%Co	%Fe2O3	%MgO	%SiO2
COV	%Ni	1.00
	%Co	0.53	1.00
	%Fe2O3	0.73	0.37	1.00
	%MgO	0.33	0.14	0.06	1.00
	%SiO2	-0.49	-0.20	-0.84	0.04	1.00
LIM	%Ni	1.00
	%Co	0.43	1.00
	%Fe2O3	0.25	0.05	1.00
	%MgO	0.21	0.08	-0.25	1.00
	%SiO2	-0.21	-0.02	-0.59	0.23	1.00
SAP	%Ni	1.00
	%Co	0.06	1.00
	%Fe2O3	0.08	0.78	1.00
	%MgO	-0.06	-0.72	-0.93	1.00

%Fe₂O₃ which is dominant in the fines is positively correlated with %Ni and %Co in the COV and LIM layers and decreases from the top to the bottom of the profile. On the other hand %SiO₂ and %MgO are dominant in the coarser fractions and as a result show strong negative correlations with %Fe₂O₃ in all the layers. %SiO₂ and %MgO increase towards the bottom of the profile. Binary plots between the major oxides and relative depth within each layer can be used to illustrate the correlation between these major components and their relative position within the profile. Figures 16.4 to 16.6 constitute examples of binary diagrams constructed between their major elements in the SAP layer for area 217Z2. These binary diagrams can also be used to show the presence of inclusions of exogenous material, e.g. limonite within the saprolite layer as illustrated on Figure 16.7.

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Figure 16-4: %Fe₂O₃ -%SiO₂ binary diagram in the SAP layer: example of Area 217Z2

Figure 16-5: %Fe₂O₃ -%MgO binary diagram in the SAP layer: Example of Area 217Z2

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Figure 16-6: %SiO₂-%MgO binary diagram in the SAP layer: Example of Area 217Z2

Figure 16-7: %Fe₂O₃ -%MgO binary diagram in the SAP layer: Example of Area 212

Golightly Geosciences has established that dry density can be confidently correlated with chemistry and as a result also increases with depth as well as with the %MgO content and inversely decreases with an increasing %Fe₂O₃ (Figures 16.8 and 16.9). The strong correlations observed between dry density and chemistry in the SAP layer emphasizes the need to weight by this factor during the compositing and interpolation processes.

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Figure 16-8: Dry density-%MgO binary diagram in the SAP layer: Example of Area 212

Figure 16-9: Dry density-%Fe₂O₃ binary diagram in the SAP layer: example of Area 212

A global comparison was conducted in area 217z2 in order to assess the materiality of weighting by density on the interpolation results in the SAP layer (0. Tavchandjian, 2009b). The results of this comparative study are summarised in table 16.4 and clearly show that the impact of weighing by density during compositing and interpolation is significant for all the elements including %Ni and is as expected decreasing grade for elements more common in the fines and increasing grade for elements occurring in the coarser fraction of the saprolite layer.

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Table 16.4: Impact of Density Weighting in the Interpolation Process –

Example of the 217z2 area – SAP Layer

Element	Average Grade (Not weight by Density)	Average Grade (Density weighted)	Relative Difference
% Ni	1.36	1.26	-7%
% Co	0.041	0.032	-22%
% Fe2O3	24.0	20.4	-15%
% MgO	22.3	25.1	+13%
% SiO2	35.0	36.8	+5%

Based on the results of the exploratory data analysis, the variables selected for interpolation are service variables defined as the product of grade and dry density. Although non-linear such as indicator transformation may prove useful when variables present a high degree of skewness as observed at Fenix for %Co and the major oxides, since these elements are of moderate concern with respect to their local estimates, this level of sophistication was not deemed to add value in this particular modeling exercise.

This “metal quantity” is then used for variogram calculation and interpolation. Grade is obtained by dividing for each block in the model the interpolated quantity of metal (e.g. QNi = Ni x density) by the interpolated density. This approach was used for the modelling of all the elements in the SAP layer.

Because the validity of the dry density measurements remains subject to a lower confidence and also considering that no resource will be reported from the LIM and COV layer, service variables were not used for these layers and interpolation was conducted directly on grade.

16.6 Data spatial Analysis

16.6.1 Unfolding

The unfold process within Datamine Studio 2 was used to transform the composite sample data coordinates into an unfolded coordinate system, as defined by the layer’s top and bottom surfaces. This transformation essentially provides a very effective way to consider “geological distances” between samples rather than traditional Cartesian distances, allowing for more robust variogram calculations and grade estimations. This approach also greatly helps in honouring the observed correlations between chemistry and vertical position in the layer profile.

In this process, a very important decision in the unfolding of each layer consists in deciding if vertical coordinates should assume that the metal distribution is following one or both of the surfaces and if only one of the surfaces should be used as a reference if it should be the top or bottom surface. A consensus was reached between Golder and CGN’s Chief Geologist that the bottom surface only should be used as a reference for the COV and LIM layers while the top surface should be used as the reference surface for the SAP layer. This process implies for example that grade intersected in the bottom of a trough in the SAP layer would not be extended laterally when surrounded by rock pinnacles.

The top and bottom DTMs of each surface were sliced on South-North sections on every Easting corresponding to an actual drill-hole location with a minimum separation distance of 1 metre between sections. A maximum separation distance of 20 metres was also used between sections. In order to

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ensure a maximum efficiency in the unfolding of all the samples, the top and bottom surfaces were translated respectively up and down by 3 to 5 meters depending on the area. This process does not change their relative vertical position and simply ensures that both the samples (within 1 metre) and the block discretization points (within 10 metres) that do not exactly fit on one of the unfolding sections will still be unfolded. All samples were successfully unfolded in all the zones.

After unfolding, the samples were visualised both in XYZ and in unfolded space in order to check that the unfolding process worked as intended. No significant issues were found in the unfolding process. Figure 16.10 shows as an example the SAP samples before and after unfolding in the 212 area.

Location of samples in Cartesian coordinates (vertical exageration = x2)

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Location of samples in Unfolded coordinates (vertical exageration = x2)

Figure 16-10: Comparison of Cartesian and unfolded location of the

SAP samples in the 212 area

16.6.2 Variography

Experimental grade variograms were calculated for %Ni, %Co, %Fe₂O₃ , %SiO₂ and %MgO in unfolded space. No noticeable non-orthogonal grade trends were noticed in the unfolded samples so no rotations were applied to the variogram calculations. An example of variogram map is presented on Figure 16.11 for Area 212.

Figure 16-11: Example of variogram map of %Ni in the saprolite layer (Area 212)

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The parameters used in the calculation of the experimental variograms are provided in Table 16.5. Because of the large number of drill-holes and the significant aspect ratio between the vertical and horizontal sample density, the variograms were estimated using a very tight tolerance on orientation. For the same reason, the experimental variograms were filtered on a minimum of 300 pairs of samples for modeling purposes. For zones 213z1, 213z2, 215inf and 217z3 no meaningful variograms could be constructed due to either of their short extent, the small number of samples available or the large spacing between drill-holes. Following discussions with CGN geologists, it was deemed that the 212 area was most similar and variogram models for this area were used for modeling in the 213z1 and 213z2 areas while the 217z2 variogram models were applied to the 215inf and 217z3 areas. The variograms were all divided by the sample variance so that the sill would tend to a value of 1.0 for all the elements. This was used as an effective tool to rapidly derive variogram models for variables with strong linear correlations.

Table 16.5: Experimental Variogram Parameters

Lag Distance	10
Number of Lags	25
Sub-lag Distance	10
Number Lags to be Sub-lagged	2
Rotation	none
Angular tolerance in horizontal plane	20	°
Angular tolerance in vertical plane	5	°

Interactive fitting of models to the experimental variograms was carried out using enhanced geo-statistical tools in Datamine Studio 2. Working with the experimental variogram of grade in COV and LIM and of quantity of metal and dry density in SAP, variogram modelling was obtained using a two structure anisotropic model with the addition in most cases of a vertical zonal anisotropy in order to account for the expected additional vertical variability due to the laterization process. The second structure ranges were used to help define the search distances in the three directions for both the Nearest Neighbour (NN) and Ordinary Kriged (OK) estimates.

In general, two series of variogram models can be identified corresponding to two different types of zones both in terms of mineralization and in terms of drilling coverage. The areas 212, 213 and 217z1 present in general a higher degree of serpentinization in SAP usually reflected by a higher Ni content. These zones have been for the most part drilled by Auger drilling only during Inco’s ownership in the 1970’s. The variogram models for this zone were based on the dense Auger drilling data over the 212 area. It was then checked that the same models could reasonably fit experiemental variograms with sometimes some rescaling of the sill for areas 213 and 217z1. The validation of these models focused on the quantity of Ni in the SAP layer. Small discrepancies for the other elements were not considered material.

Variogram models were then adjusted in area 217z2 where dense core drilling information is available. This second set of models was then found to reasonably fit the data in areas 215, 251 and 217z3. In area 216, a small adjustment was made to the quantity of Ni variogram to improve the fit to the actual

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experimental variogram obtained in this area. This second group of areas was recently redrilled by Skye Resources using diamond drilling. It is not entirely clear at this stage if the differences observed between the variograms are due to mineralization or to the drilling method. This issue will be further investigated by CGN with a dedicated diamond drilling program which will aim at replace the existing Auger information over the 212, 213 and 217z1 areas. Since Ni is the primary element of interest the search strategy developed to construct the interpolated resource models was based on the variogram models adjusted to the Ni service variables. These variogram models are summarised in Table 16.6.

Table 16.6: Variogram models used for the Ni Service Variables in Saprolite

	212, 213z1, 213z2, 217z1	217z2, 217z3, 215, 215inf, 251	216
nugget	0.14	0.05	0.05
sill 1	0.61	0.40	0.40
range 1 X (m)	30	20	20
range 1 Y (m)	15	20	20
range 1 Z (m)	7.5	5	5
sill 2	0.25	0.55	0.55
range 2 X (m)	230	130	130
range 2 Y (m)	135	130	130
range 2 Z (m)	19	20	15
sill 3	0.22	—	—
range 3 X (m)	10000	—	—
range 3 Y (m)	10000	—	—

16.7 Block Model Definition

A block model was created to cover the volume of the mineralized zone in the coordinate system of the drillhole data for each area. The block size was set to 12.5 metres x 12.5 metres x 1 metre in all areas Based on the assumption (common in Ni laterites) that the mining will be conducted with a combination of dozers and backhoes and avoiding blasting, the blocks were entirely assigned to one of the layer (no sub-blocking) based on the location of their centroid.

The use of a 12.5 metres x 12.5 metres x 1 metre block size was selected as a reasonable approximation of the selective mining unit (smu) that will be achieved during mining. This is an approximation only as mining in laterites does not proceed by extraction of blocks of uniform dimensions but rather as a combination of stripping of the COV and LIM layers to expose the mineralised saprolite followed by very selective mining with the use of backhoes. The 12.5 metres x 12.5 metres x 1 metre block size was based on the author’s experience with other similar Ni laterite operations and was ultimately validated with a reconciliation exercise conducted over the former La Gloria mine (O. Tavchandjian, 2009c).

A volume check of the regularized block model vs. the wire frames used to generate it revealed a very good representation of the volume overall for each area with the exception of the very small volume of COV in the 213 Area which are of no consequences for resource modeling or mine planning (Table 16.7).

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Table 16.7: Block Model versus Wire Frame Volume Check

		DTM Method	Block model volume			DTM Method	Block model volume
212	COV	1,309,412	1,317,968	216	COV	7,348,711	7,348,281
	LIM	4,201,551	4,215,625		LIM	13,481,199	13,454,844
	SAP	7,705,629	7,717,500		SAP	16,531,950	16,521,875
213z1	COV	3,387	4,688	217z1	COV	474,000	477,813
	LIM	58,308	59,375		LIM	2,266,755	2,260,000
	SAP	895,253	896,562		SAP	3,686,490	3,704,063
213z2	COV	4,000	6,250	217z2	COV	7,838,135	7,828,281
	LIM	—	—		LIM	16,220,586	16,194,375
	SAP	309,201	312,812		SAP	17,264,008	17,246,250
215	COV	3,583,694	3,598,906	217z3	COV	6,066,563	6,037,965
	LIM	8,287,708	8,308,906		LIM	6,624,531	6,588,910
	SAP	6,881,256	6,894,687		SAP	19,106,094	19,101,505
215inf	COV	7,947,976	8,035,000	251	COV	3,265,834	3,287,969
	COV	11,744,548	11,813,906		LIM	15,914,435	15,906,875
	LIM	15,768,814	15,817,343		SAP	15,870,299	15,883,125

16.8 Estimation Methodology

Block model grades for %Ni, %Co, %Fe₂O₃ , %SiO₂ and %MgO and dry density were estimated using Ordinary Kriging (OK). The interpolation was conducted on the service variables defined as the product of grade and density for each metal listed above for the SAP layer. For the LIM and COV layers, grade was estimated without the use of service variables since dry density is set to a constant value for these layers due to the poor confidence in density measurements.

Nearest Neighbour estimates of the same elements provided declustered sample grades for block model validation. Anisotropic searches were performed, using the %Ni variogram model ranges in saprolite as a guide for each of the 3 axes orthogonal to the plane of the deposit in unfolded space. Because the quantity of nickel in saprolite is the main element of interest and also because the variogram models were not vastly different between elements and between areas, the same search strategy was used for all the elements, for all the layers and for all the areas. The search parameters used for each layer are summarized in Table 16.8.

Table 16.8: Search Parameters Used for OK Interpolation

Parameters	1st Search	2nd Search	3rd Search
Search radius in metres in unfolded coordinates	70x70x5	130x130x10	390x390x30
Rotation	None	None	None
Minimum number of octants with samples	5	5	1
Maximum number of samples per octant	4	4	4
Minimum number of samples	20	20	1
Maximum number of samples	32	32	32

Three nested searches were utilized to grade as many blocks in the model as possible. The first search reflected the ranges determined in the variogram modeling, the second search factored these ranges by

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2, and the third search factored them by 4. Octant restrictions were used in all cases, with a minimum of 1 sample and a maximum of 4 samples in each one of at least 5 octants required for the first and second ellipsoids. Overall a minimum of 20 samples is required to complete the interpolation in the first and second search ellipsoid with a maximum of 32 samples used. The minimum sample and octant constraints were removed in the third search ellipsoid.

As part of the interpolation validation process, the search volume used in the interpolation of each block was recorded for each cell in the model. The statistics on the number of blocks interpolated in each search is provided for the SAP layer by area in Table 16.9. This data is considered for resource classification purposes. With the exception of the zones 215inf, 217z3 and 213z2, over 75% of the blocks were estimated with the first two search ellipsoids with tight octant and minimum number of samples constraints. In area 213z2 the large number of blocks interpolated in the 3^rd search ellipsoid is deemed related to the small extent of the zone and the fewer number of holes. This area is still deemed to provide good local interpolated estimates.

It should be noted that usually only a few percent of the blocks were not graded during the interpolation process. Factors contributing to this were; lack of block unfolding during the interpolation process and octant searching at the edges of the deposit. Since these blocks were predominantly along the edges of the model they are not considered to have any detrimental effect on the resource estimate.

Table 16.9: Statistics on Interpolation Results per Search Ellipsoid

Area	Proportion of Blocks Estimated in the SAP Layer
	1st Search		2nd Search		3rd Search
212	85.0	%	8.0	%	7.0	%
213z1	75.0	%	10.0	%	14.0	%
213z2	9.0	%	33.0	%	57.0	%
215	32.0	%	46.0	%	21.0	%
215inf	1.0	%	5.0	%	89.0	%
216	4.0	%	73.0	%	23.0	%
217z1	46.0	%	29.0	%	24.0	%
217z2	52.0	%	26.0	%	21.0	%
217z3	0.0	%	6.0	%	94.0	%
251	42.0	%	42.0	%	15.0	%

16.9 Model validation

16.9.1 Global statistical validation

Global statistics were systematically compared between the sample composites, nearest neighbour and ordinary kriging models. In the majority of the zones, there is a decrease Ni metal content between the composite sample stats and all the modelling techniques. This result was expected as the drilling is usually focused in the high grade areas.

Comparisons show in all cases very good consistency between the nearest neighbour (NN) and ordinary kriged (OK) average estimates confirming the absence of global bias. These systematic comparisons also confirm as expected the smoothing effect of OK reflected in the lower variance between blocks.

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16.9.2 Visual and local statistical validation

Visual validation of the block model did not reveal any apparent problem in grade distribution between the MIK model and the drillhole data. In addition, Figure 16.12 illustrates the visual validation process and some examples of the benefits of using an unfolding algorithm in the interpolation process.

Figure 16-12: Example of OK Model Visual Validation - 212 Area

Swath plots validations were also conducted to check for the absence of bias in the OK model These swath plots were conducted systematically for %Ni in the saprolite layer and randomly for the other elements and for %Ni in the other layers. Visual examination of these swath plots did not indicate any problem.

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16.9.3 Smoothing assessment and correction

An assessment of the level of smoothing was conducted in order to compare the variance between the 12.5 metres x 12.5 metres x 1 metre blocks obtained with each modeling technique to the theoretical variance of the particular block geometry according to the variogram models. Based on the %Ni variogram models, the expected variance between blocks 12.5 metres x 12.5 metres x 1 metre is provided for each area in Table 16.9. The expected variance is also provided for comparison purposes for a selective mining unit (smu) size of 10 metres x 10 metres x 3 metres which is deemed to be closer to the practical smu to be utilised during the mining operation. This assessment of the smoothing was only conducted on %Ni and for the SAP layer since a cut-off grade will only be considered for this element and in this particular layer. Where deemed necessary, the areas were further divided into sub-domains for smoothing assessment and correction purposes. This further sub-division was deemed necessary in areas with variable drill-spacing and/or marked differences in the %Ni histogram.

Table 16.10: Smoothing Assessment and Correction per Area

Area	NN Model		OK Model				Corrected OK Model
Parameter	Mean	Variance	Mean	Variance	Theoretical 10mx10mx3m block variance	Smoothing ratio	Mean	Variance
212	1.79	0.510	1.76	0.160	0.280	1.73	1.76	0.280
213z1	1.92	0.594	1.87	0.179	0.327	1.83	1.87	0.330
213z2	1.65	0.339	1.56	0.064	0.186	2.91	1.56	0.187
215	1.54	0.254	1.52	0.078	0.165	2.12	1.52	0.163
215inf	1.43	0.193	1.4	0.076	0.125	1.65	1.39	0.124
216West	1.42	0.212	1.41	0.063	0.144	2.29	1.40	0.143
216East	1.15	0.200	1.15	0.062	0.136	2.19	1.15	0.135
217z1East	1.53	0.278	1.49	0.042	0.153	3.64	1.49	0.153
217z1West	1.70	0.410	1.65	0.092	0.226	2.45	1.64	0.227
217z2	1.47	0.305	1.45	0.098	0.218	2.22	1.45	0.218
217z3	1.10	0.357	1.10	0.155	0.250	1.61	1.10	0.229
251	1.59	0.334	1.57	0.097	0.224	2.31	1.56	0.225

Table 16.10 indicates that the variances between blocks obtained by OK interpolation are as expected lower than their theoretical values. Smoothing is a common and desirable feature of a proper interpolation and reflects the level of uncertainty on the estimates. This over-smoothing in the OK %Ni estimates was corrected using an affine correction (Journel and Hjuibregts, 1983) as the histogram of the Ni grade is generally close to a normal distribution for the SAP layer in all the areas. The %Ni mean and variance before and after the affine correction are presented in Table 16.10 for each area and are in very good agreement with the expected values calculated from the variogram models.

16.9.4 Independent peer reviews

As per Golder standard best practices for resource modeling and estimation, independent peer reviews were conducted by Greg Greenough, P.Geo and Kevin Palmer, P.Geo for three selected areas: 212, 215 and 217. No material issues were found and the resource modeling methodology was independently validated.

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16.10 Mineral Resource Reporting and Classification

16.10.1 Mineralization with potential for economic extraction

In order to report mineral resource estimate, the potential for economic extraction of the mineralization was defined as follow:

• Only the SAP portion of the mineralization can be economically extracted and has a suitable chemistry (note that internal limonite dilution has already been incorporated in the geological model).

• After compiling grade-tonnage curves, a cut-off grade of 1.6% Ni in areas 212, 213z1, 213z2, 217z1 and 251 and 1.5% Ni in areas 215, 215inf, 216, 217z2 and 217z3 was deemed to constitute the optimum compromise between resource utilization and minimum acceptable economic return on investment.

• The %Ni cut-off grade was applied after correcting for over-smoothing of the kriged estimates as described in section 16.9.3.

• A minimum of two consecutive vertical blocks above cut-off (2 metres) was applied as an additional constraint in order to remove thin and isolated layers.

• A maximum of 2 metres of vertical internal waste within the “ore layer” was also applied based on the assumption than any larger waste inclusion could be selectively mined out and disposed without entering the plant feed stream.

• No blocks grading above cut-off were discarded as waste based on their chemical composition. It was assumed by Golder that blocks with extreme chemistry could be successfully blended by operating different part of the deposits simultaneously and by an expansion of the existing stacking-reclaiming facility.

• Finally a Lerchs-Grossman algorithm (Garcia, 2009) was applied in order to eliminate the blocks which would necessitate an unrealistic amount of incremental stripping or mine development to be successfully extracted.

The constraints applied through the Lerch-Grossman and minimum ore thickness criteria result in a very small portion of the blocks above cut-off to be excluded from the mineral resource estimate. In the opinion of Golder, approximately 97% of the mineralization grading above cut-off presents potential for economic extraction.

16.10.2 Mineral resource classification

The confidence in mineral resource estimates usually increases with a higher drilling density, better continuity of the mineralization at the anticipated cut-off grade and higher annual production rates. The relationship between drilling density and the continuity in the mineralization is usually captured through the variogram, which is used to define the search strategy during the interpolation process. Table 16.11 provides a useful summary in that respect and clearly indicate that areas 215inf and 217z3 were based on very sparse information and should be classified based on this criteria as inferred mineral resources. However, Table 16.11 does not capture the impact of cut-off grade (confidence in the estimates usually decreases with a higher cut-off grade) and the impact of the annual production rate (it is easier to predict the average recoverable tonnage and grade above cut-off over a large area than over a small area). In addition, a specific challenge with respect to the classification of mineral resource estimates for the Fenix

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project is the significant contribution of Auger data in areas 212, 213z1, 213z2 and 217z1. Tavchandjian (2009b) has established however, that the Auger estimates tend to under-estimate the tonnage of economic mineralization and as such should not constitute a penalizing factor on resource classification. The approach of trying to correct for biases in Auger drilling to support additional inferred mineral resource estimates has not been considered by Golder.

Based on experience in other similar deposits, the resource classification was based on the following factors:

• Corresponding percentile in the SAP composite database of the %Ni cut-off

• Proportion of blocks estimated in the various search ellipsoids (Table 16.8)

• Number of holes to cover the area supporting one year of feed to the plant

• Number of holes to cover the area supporting one year of feed to the plant

The parameters obtained for each area to support the recommended resource classification scheme are summarised in Table 16.11.

Table 16.11: Summary of Resource Classification Criteria by Area

Area	Proportion of blocks estimated in the SAP layer						Drill spacing (Type of drilling)	%Ni cut- off equivalent percentile	Number  of holes supporting annual production estimates	Number  of holes supporting quarterly production estimates	Recommended classification
	1rst search		2nd search		3rd search
212	85	%	8	%	7	%	25m (Auger)	P35	256	64	Measured
213z1	75	%	10	%	14	%	25m (Auger)	P38	256	64	Measured
213z2	9	%	33	%	57	%	50m (Auger)	P45	64	16	Indicated
215	32	%	46	%	21	%	50m (Core)	P43	64	16	Indicated
215inf	1	%	5	%	89	%	200m (Auger)	P51	4	1	Inferred
216	4	%	73	%	23	%	75m (Core)	P63	28	7	Indicated
217z1 West	68	%	18	%	14	%	25m (Auger)	P43	256	64	Measured
217Z1 East	1	%	55	%	44	%	50m (Auger)	P52	64	16	Indicated
217z2	52	%	26	%	21	%	50m (Core)	P49	64	16	Indicated
217z3	0	%	6	%	94	%	200m (Auger)	P70	4	1	Inferred
251	42	%	42	%	15	%	50m (Core)	P47	64	16	Indicated
212	85	%	8	%	7	%	25m (Auger)	P35	256	64	Measured

General Guidelines:

Measured: If >50% of blocks estimated in 1st search and >80% of blocks estimated in first two searches, >50% of samples above cut-off, more than 50 holes per quarter of anticipated production

Indicated: If ~50% of blocks estimated first two searches, ~50% of samples above cut-off, more than 25 holes per year of anticipated production

Inferred: If >75% of blocks estimated in third search, <50% of samples above cut-off, less than 25 holes per year of anticipated production

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16.10.3 Mineral Resource statement

The mineral resource estimates for the areas 212, 213, 215, 216, 217 and 251 prior to conversion to mineral reserves are summarised in Table 16.12 below. The inferred mineral resource estimates can only be used in preliminary economic assessment exclusive of the mineral estimates in order to investigate the potential for project expansion and cannot be converted to mineral reserve estimates.

Table 16.12: Mineral Resource estimates for Areas 212, 213, 215, 216, 217 and 251

Classification	Area	%Ni cut-off	000’ Dry Tonnes	%Ni
Measured	212	1.6	4,793	2.09
	213z1	1.6	724	2.19
	217Z1 East	1.6	2,218	2.09
	Total		7,735	2.09
Indicated	213z2	1.6	168	1.94
	215	1.5	3,346	1.85
	216	1.5	5,895	1.75
	217z1 West	1.6	537	1.82
	217z2	1.5	9,441	1.86
	251	1.6	9,067	1.97
	Total		28,454	1.87
Measured + Indicated	Total		36,190	1.92
Inferred	215inf	1.5	5,800	1.8
	217z3	1.5	3,900	1.8
	Total		9,700	1.8

Note: Given the superficial nature of the deposits and the large drill spacing supporting this category, the inferred mineral resource estimates are reported without application of a Lerchs-Grossman algorithm.

16.11 Reconciliation with Historical Production Records

A reconciliation study was conducted between the block model constructed for the 212 area and the production records available for the three zones excavated as part of the former La Gloria mine area which is located within the 212 area (Figure 16.13).

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Figure 16-13: Drilling coverage and perimeters of mined out areas

Details of the reconciliation study are summarised in Tavchandjian (2009b). The excavated volumes reported by Snowden in the 2007 technical report were compared with the mining perimeters, the recent 2009 pit survey and the historical Exmibal production records (RC Osborne, Pers. Com.). Table 16.13 shows very close agreement between the global excavated volumes reported from the 2007 technical report and the historical production records. Comparison of the excavation records with the volume of economic saprolite mineralization available in the resource model however indicates that production was abandoned before the excavation reached the bottom of the “ore layer” for two of the three zones. This was confirmed during a site visit by Golder personnel. This finding was later confirmed from historical records of Exmibal and through in depth comparisons of the excavation survey and LIDAR topographic data. In order to provide a meaningful reconciliation, an elevation cut-off was applied in the areas 212-1 and 212-5. It was empirically found that reporting volumes for each type of material above elevation 297 metres in area 212-5 and above elevation 298 metres in area 212-1 was giving the best global fit on volume between the Golder model and both the technical report tables and RC Osborne’s production report.

After applying the elevation adjustment described above, the tonnes and grade reported from the Golder resource model at a 1.6% Ni cut-off for the mined out areas were compared to the production tables of RC Osborne for each area. The tonnage and grade were reported from the resource model applying a 1.6% cut-off on the Ni grade corrected for over-smoothing and using a 2m minimum ore width and a 2 metre maximum waste allowance. Results obtained are summarised in Table 16.13 and demonstrate that the Golder resource model requires minimal dilution to match the production records.

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Table 16.13: Comparison of Golder’s Mineral Resource Estimates

and Exmibal Historical Production Records

	212-1	212-5	400	TOTAL
Production Records
Dry Tones	526,628	298,746	105,497	930,871
%Ni	2.17	2.13	2.03	2.14
Golder Resource Model
Dry Tonnes	523,491	297,300	109,238	930,029
%Ni	2.19	2.16	2.07	2.17

Based on the reconciliation results, it is deemed that only ballast dilution estimated at 3% (Pullman 2009) and a mining recovery factor of 95% based on operational results at other similar operations need to be applied in order to convert the mineral resource estimates to mineral reserve estimates.

17.0 PROSPECTS FOR ECONOMIC EXTRACTION

A test to justify the reasonable prospects for economic extraction was conducted. This test included updating the 2006 Feasibility Study cash flow model supporting the historic reserve number with the newly calculated resource number outlined in this technical report, and applying reasonable mining recovery and dilution numbers to the measured and indicated resources. As anticipated, the project’s realised return and cash flow after capital and operating costs were positive, justifying the determination that the contained resource provides for reasonable prospects for economic extraction. Due to the ongoing work by Golder and Hatch, HudBay felt guidance on specific costs were best left for completion of the updated Feasibility Study at which time HudBay will commission the publication of a new Technical Report and declaration of reserves.

18.0 INTERPRETATION AND CONCLUSIONS

The Fenix nickel deposits are typical wet-tropical climate laterites and occur as surficial deposits on topographic 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.

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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.

Golder 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 Golder’s opinion the compiled data are acceptable for Mineral Resource estimation.

Golder prepared saprolite Mineral Resource estimates for Ni, Co, Fe₂O₃ , MgO, SiO₂ and dry density for Areas 212, 213, 215, 216, 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, cut-off grade and potential production rate into account.

The Golder Mineral Resource estimates were peer reviewed and reconciled with mine production records.

Golightly Geoscience (2006) and AMEC (2003) audited and validated 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 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.

19.0 RECOMMENDATIONS

The following program is recommended by Golightly Geoscience and Golder 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.

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• Additional drilling, both twin and in-fill, followed by updated mineral resource and reserve estimates, where warranted, in the following areas:

• Diamond drilling the auger supported resource areas of 212 and 213 that are to be incorporated into the mine plan in the updated Feasibility Study.

• 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 metre spacing is recommended for 215-3, 215-2 (North) and 216 (north), 260 (Amate) and 219.

• Reconnaissance of favourable landforms in the Chulac area west of Amate.

• 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%. HudBay’s mining and grade control plan needs to address troughs and pinnacles in the bedrock.

• Three ground penetrating radar profiles were done the La Gloria area but have never been used. There have been considerable advances in recent years by Ground Probe in Ni laterite interpretation. It is recommended that this method should be tried again.

An estimate of costs is provided in Table 19.1.

Table 19.1: Recommended Budget

Area	Holes	Length	Cost US$		Purpose	Status
212-213	254	10,000	$	1,500,000	Diamond drill supported resource vs. auger	In progress
217	20	480	$	415,000	20 cm core holes or bulk	Recommended
215	20	480	$	415,000	sampling, for density and	Recommended
251	20	480	$	415,000	upgrading potential	Recommended
Subtotal	314	11,440	$	2,745,000
	Additional Exploration Drilling Recommended					Priority
218-1	510	12,000	$	2,700,000	50 m in-fill	1
218-2	355	8,400	$	1,800,000	50 m in-fill	1
221	100	2,400	$	520,000	Initial Assessment Twins	2
222	20	470	$	105,000	Initial Assessment Twins	3
260	50	1,200	$	260,000	50 m	4
219	56	1,300	$	290,000	50 m	5
Subtotal	1,091	25,770	$	5,675,000
Ground penetrating radar test	3 line km. @ $3000 per processed line km + mobilization		$	9,000	To assess usefulness for mapping saprolite ore-zone thickness details.	Suggest Ground Probe due to laterite experience
Database verification			$	85,000		Golightly Geoscience & Golder
Resource estimates			$	70,000		Golder
Total	1,405	37,210	$	8,584,000

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.

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20.0 REFERENCES

AMEC, 2003. Technical Report: Exmibal Nickel Project AMEC: by Dominique François Bongarçon, Ph. D. & Brian Montpellier, P.Eng. December 2003.

Anfield Nickel, 2009. Mineral Resource Estimate For The Mayaniquel Project, Guatemala. NI 43-101 Technical Report

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.

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.

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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.

Tavchandjian, O., 2009a, 09-1117-0037_HudBay_Minerals_Fenix_AugervsCore comparison, Golder Associates Confidential Technical Memorandum, 11p.

Tavchandjian, O., 2009b, 09-1117-0037 Tech Memo05Oct09 Reconciliation La Gloria Mine, Golder Associates Confidential Technical Memorandum, 7p.

Toomver, 1979. Development of Inco’s Selective Reduction Smelting Process for Nickel Laterite Ores. In International Laterite Symposium pp255-257.

More Recent Reports:

Golightly, 2006. Montufar: Possible Sites for Twin Holes, 17 August 2006

Golightly, 2007. Variation of Moisture Content with Depth, Report, 28 February

Golightly, 2007. Application of Mass Balance Concept to the Distribution of Nickel Laterite in the Fenix Project, Report 12 March 2010

Golightly, 2007. Montufar: Identifying Found Pits & Boreholes, Powerpoint 17 June

Golightly, 2007, Montufar: Assessment of Multielement Chemistry, DRAFT report 13 November 2007

Golightly, 2007. Comparison of New Montufar Drilling with Historic Exmibal, Report 20 November.

Golightly, 2010. Fenix Project: An Overview of QA/QC and Procedures, 18 March 2009

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Report Signature Page

Effective date of March 31, 2010

GOLDER ASSOCIATES LTD.

Original Signed and Stamped By:	Original Signed and Stamped By:
Olivier Tavchandjian, P.Geo	Paul Golightly, P.Geo
Senior Project Manager	President
Golder Associates Ltd.	Golightly Geoscience Ltd.
Original Signed By:
Greg Warren
Associate
Golder Associates Ltd. OT/GW/ms/dh

Effective date of March 31, 2010

Golder, Golder Associates and the GA globe design are trademarks of Golder Associates Corporation.

March 2010 Report No. 09-1117-6013

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

APPENDIX A

Certificates of Qualifications and Consent to File

March 2010 Report No. 09-1117-6013

Golder Associates Ltd.
2390 Argentia Road
Mississauga, Ontario L5N 5Z7
Telephone: (905) 567-4444
Fax: (905) 567-6561

OLIVIER TAVCHANDJIAN

CONSENT OF AUTHOR

TO: British Columbia Securities Commission

Alberta Securities Commission

Saskatchewan Financial Services Commission (Securities Division)

Manitoba Securities Commission

Ontario Securities Commission

Autorite des marches financiers

New Brunswick Securities Commission

Nova Scotia Securities Commission

Superintendant of Securities, Prince Edward Island

Securities Commission of Newfoundland

AND TO:      HudBay Minerals Inc.

I, Olivier Tavchandjian, Ph.D., P.Geo., (the “Author”) do hereby consent to the public filing of the technical report titled, “Update to the Mineral Resource estimates of the Fenix Project Izabal Guatemala” (the “March 31, 2010 Technical Report”), with an effective date of March 31, 2010 and to any extracts from or a summary of the March 31, 2010 Technical Report in the press release of HudBay Minerals Inc. dated March 31, 2010 (the “Press Release”).

I also certify that I have read the Press Release, including the extracts from or the summary of the Technical Report contained in the Press Release and it fairly and accurately represents the information in the Technical Report that supports the disclosure.

Dated this 31^st Day of March, 2010

Original Signed and Stamped by:

Olivier Tavchandjian, Ph.D., P.Geo.

OFFICES ACROSS NORTH AMERICA, SOUTH AMERICA, EUROPE, AFRICA, ASIA AND AUSTRALIA

OLIVIER TAVCHANDJIAN

CERTIFICATE OF QUALIFICATION

I, Olivier Tavchandjian, Ph.D., P.Geo., do hereby certify that:

1. I am employed as a Senior Project Manager at:

Golder Associates Ltd.

2390 Argentia Road Suite 200, Ontario L5N 5Z7

Telephone: 905-567-4444(ext. 1344); Fax: 905-567-6561; Email: Olivier_Tavchandjian@golder.com

2. I graduated with a Bachelor’s degree in Geology from the University of Aix-Marseille (France). In addition I have obtained a Master in Geology from the University of Aix-Marseille (France) and a Ph.D. in Mineral Resources from the Universite du Quebec a Chicoutimi (Canada).

3. I am a member in good standing of the Association of Professional Geologists of Ontario (APGO).

4. I have practised my profession continuously since graduation. My relevant experience with respect to this project is over twenty years of experience in exploration and mineral resource evaluation of mineral projects nationally and internationally including over ten years Ni laterites operations in Indonesia, New Caledonia, Australia and Brazil.

5. I am responsible for Sections 1 to 3 and 16 to 19 (that are excerpts from the March 31, 2010 Technical Report) of the technical report titled “Update to the Mineral Resource estimates of the Fenix Project, Izabal, Guatemala” (the “March 31, 2010 Technical Report”) with an effective date of March 31, 2010. I visited the Fenix Project site for one week in August 2009.

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

7. I have not had prior involvement with the property that is subject to the March 31, 2010 Technical Report.

8. I am not aware of any material fact or change with respect to the subject matter of the March 31, 2010 Technical Report as of the date of this certificate, to the best of my knowledge, information and belief that sections I am responsible for as stated in Bullet 5 contains all scientific and technical information that is required to be disclosed to make the March 31, 2010 Technical Report not misleading.

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

10. I have read NI 43-101 and form 43-101F1, and the March 4, 2010 Technical Report has been prepared in compliance with that instrument and form.

Dated this 31^st Day of March, 2010.

Original Signed and Stamped by:

Olivier Tavchandjian, Ph.D., P.Geo.

Golder Associates Ltd.

2390 Argentia Road, Mississauga, Ontario L5N 5Z7

Tel: +1 (905) 567 4444 Fax: +1 (905) 567 6561 www.golder.com

Golder Associates: Operations in Africa, Asia, Australasia, Europe, North America and South America

	Golightly Geoscience Ltd.
	488 Lakepoint, Sudbury, Ontario P3E 6J3
	Tel. 1-705-688-1503	E-mail: golgeo@sympatico.ca

J. PAUL GOLIGHTLY

CONSENT OF AUTHOR

TO: British Columbia Securities Commission

Alberta Securities Commission

Saskatchewan Financial Services Commission (Securities Division)

Manitoba Securities Commission

Ontario Securities Commission

Autorite des marches financiers

New Brunswick Securities Commission

Nova Scotia Securities Commission

Superintendant of Securities, Prince Edward Island

Securities Commission of Newfoundland

AND TO:        HudBay Minerals Inc.

I, J. Paul Golightly, P.Geo., (the “Author”) do hereby consent to the public filing of the technical report titled, “Update to the Mineral Resource estimates of the Fenix Project, Izabal, Guatemala” (the “March 31, 2010 Technical Report”), with an effective date of March 31, 2010 and to any extracts from or a summary of the March 31, 2010 Technical Report in the press release of HudBay Minerals Inc. dated March 31, 2010 (the “Press Release”).

I also certify that I have read the Press Release, including the extracts from or the summary of the Technical Report contained in the Press Release and it fairly and accurately represents the information in the Technical Report that supports the disclosure.

Dated this 31^st Day of March, 2010.

J. Paul Golightly, Ph.D., PGeo

	Golightly Geoscience Ltd.
	488 Lakepoint, Sudbury, Ontario P3E 6J3
	Tel. 1-705-688-1503	E-mail: golgeo@sympatico.ca

J. PAUL GOLIGHTLY

CERTIFICATE OF QUALIFICATION

I, J. Paul Golightly, President of Golightly Geoscience Ltd., 488 Lakepoint, Sudbury, Ontario, P3E 6J3, do hereby certify that:

1. I graduated with an Honours Degree in Bachelor of Science Geology from McGill University in 1961.1 graduated with a degree of Doctor of Philosophy in Geology & Crystallography in 1968.

2. I am licensed as a Professional Geoscientist with the Association of Professional Geoscientists of the Province of Ontario (APGO# 0879)

3. I have worked as a geologist in mineral exploration and resource evaluation for INCO and in Consulting practice continuously for 40 years, including geology and resource evaluation of nickel laterite deposits for at least 20 years.

4. 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.

5. I am responsible for Sections 4 to 15 of the Technical Report titled “Update to the Mineral Resource estimates of the Fenix Project, Izabal, Guatemala” (the “March 31, 2010 Technical Report”) with an effective date of March 31, 2010.

6. I was consulted by INCO on a number of technical issues of the Fenix Project area deposits from 1970-1981, and prepared several mineralogical & geochemical reports on samples from the area in 1973-1977. I am the author of the technical report: Fenix Nickel Project: Nickel Laterite deposits of the Lake Izabal Region, Guatemala: Initial Assessment of activity required to re-evaluate the resources, 5 August, 2005 and am a co-author of the Technical Report: Mineral Resource Estimate, Fenix Project, Izabal, Guatemala, 4 July 2006 filed by Skye Resources.

7. I have visited the Fenix Project property eleven times and my last visit occurred from November 6 to 13, 2007.

8. As of the date of this certificate, 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.

9. I am independent of the issuer as defined in section 1.4 of the Instrument.

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

Dated at Sudbury, Ontario, 31 March, 2010

J. Paul Golightly, Ph.D., PGeo

At Golder Associates we strive to be the most respected global group of companies specializing in ground engineering and environmental services. Employee owned since our formation in 1960, we have created a unique culture with pride in ownership, resulting in long-term organizational stability. Golder professionals take the time to build an understanding of client needs and of the specific environments in which they operate. We continue to expand our technical capabilities and have experienced steady growth with employees now operating from offices located throughout Africa, Asia, Australasia, Europe, North America and South America.	Africa	+ 27 11 254 4800
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