MD Filed by Filing Services Canada Inc. (403) 717-3898

EX-99.1 2 technical.htm TECHNICAL REPORT MD Filed by Filing Services Canada Inc. (403) 717-3898

NI 43-101 Technical Report for a Preliminary Economic Assessment on the Elk Gold Project, Merritt, British Columbia, Canada

Report prepared for:

Almaden Minerals

Suite 1103-750, West Pender St, Vancouver, BC, Canada


Effective Date: Date of Issue:	10 December 2010 14 January 2011

Project Code: ALM003

Authors:

Roger Pooley, BSc (Eng), MAusIMM, Senior Mining Consultant

SRK Consulting (Australasia) Pty Ltd, Level 1, 10 Richardson St, West Perth, WA, 6005, Australia

Susan Lomas, P.Geo (BC), Principal

Lions Gate Geological Consulting Inc., 7629 Sechelt Inlet Rd, Sechelt, BC, V0N 3A4, Canada

Gary Hawthorn, BSc, P.Eng (BC), Principal

Westcoast Mineral Testing Inc., 2806 Thorncliffe Drive, North Vancouver, BC, V7R 2S7, Canada

Robert Brian Alexander, P.Geo (BC, Nunavut), Principal

Alexploration Inc., #503-2759 Carousel Crescent, Ottawa, ON, K1T 2N5, Canada

SRK Consulting		ALM003 Elk Gold Project		NI 43-101 Technical Report	January 2011


Table of Contents

1.	Summary				1
	1.1	Report Purpose			1
		1.1.1	Property Location and Ownership		1
		1.1.2	History		1
		1.1.3	Geology and Mineralization		1
		1.1.4	Exploration		2
		1.1.5	Mineral Resources Estimate		3
		1.1.6	Preliminary Economic Assessment		4
		1.1.7	Qualified Person s Conclusions and Recommendations		5

2.	Introduction				7
	2.1	Terms of Reference			7
	2.2	Report Authors			7
	2.3	Data Sources			8
	2.4	PEA Scope			8
	2.5	Effective Date			9
	2.6	Units of Measure			9

3.	Reliance on Other Experts				10

4.	Property Description and Location				11
	4.1	Property Location			11
	4.2	Tenement Details			11
	4.3	Royalties and Taxes			15
		4.3.1	Royalties		15
		4.3.2	Taxes		15
	4.4	Environmental Considerations			15
		4.4.1	Background		15
		4.4.2	Related Environmental Studies		16
		4.4.3	Environmental Impact and Management		19
		4.4.4	Permitting, Community Consultation & First Nations		19
		4.4.5	Initial Conceptual Reclamation Plans		20
		4.4.6	Summary		20
		4.4.7	Current Environmental and Social Liabilities		21

5.	Accessibility, Climate, Local Resources, Infrastructure and Physiography				22
	5.1	Access to Property			22
	5.2	Physiography and Climate			22
	5.3	Power and Water Resources			22
	5.4	Local Resources and Infrastructure			22
		5.4.1	Buildings and Ancillary Facilities		22
		5.4.2	Camp Site		23

6.	History of Property				24
	6.1	Exploration History			24
		6.1.1	Copper and Molybdenum Exploration		24
		6.1.2	Gold Exploration		24
	6.2	Historical Resource and Reserve Estimates			26
		6.2.1	Historical Mineral Reserve Estimates Completed in 1995		26
		6.2.2	Mineral Resource Estimates Completed by Giroux Consultants		26

7.	Geological Setting				28
	7.1	Geological setting			28
	7.2	Regional Geology			28
	7.3	Elk Project Geology			28

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8.	Deposit Types			30

9.	Mineralization			31
	9.1	B and WD Vein Complex (Siwash North Area)		31
	9.2	Other Exploration Areas at the Elk Project		31

10.	Exploration			33
	10.1	B and WD Vein Complex		34
	10.2	Other Showings on the Property		35
		10.2.1	Bullion Creek Area	35
		10.2.2	Siwash East Area	36
		10.2.3	Gold Creek East and West Areas	37
		10.2.4	Lake Zone Area	39
		10.2.5	Great Wall Zone Area	40
		10.2.6	End Zone Area	41
		10.2.7	Discovery Showing (North Showing Area)	42
		10.2.8	South Showing Area	42
		10.2.9	Elusive Creek	43
		10.2.10	Agur Option Area	44

11.	Drilling			45
	11.1	Surface Drill Holes		45
		11.1.1	2010 Drilling	46
	11.2	Underground Drill Holes		48

12.	Sampling Method and Approach			49
	12.1	Diamond Drill Hole Samples		49
	12.2	Trench and Rock Sampling		49
	12.3	Soil Samples		49

13.	Sample Preparation, Analyses and Security			50
	13.1	Drill Core		50
		13.1.1	1989-2006	50
		13.1.2	2007	51
		13.1.3	2010	51
	13.2	Trench and Rock Sampling		51
	13.3	Soil Samples		51

14.	Data Verification			53
	14.1	Database Audit by LGGC in 2009		53
	14.2	QAQC Data and Review		53
		14.2.1	Standard Reference Material (SRM)	55
		14.2.2	Core Duplicates	58
		14.2.3	Blank Samples	58
		14.2.4	Acme Pulp and Coarse Reject Reruns	59
		14.2.5	Acme Pulp Resubmission Duplicates	60
		14.2.6	Chemex Laboratory Check Samples	60

15.	Adjacent Properties			62

16.	Metallurgical Testing and Mineral Processing			63
	16.1	Introduction		63
	16.2	Metallurgical Testing – Historical		63
		16.2.1	Placer Dome – 1989 - 90	63
		16.2.2	Bacon Donaldson (BD) - 1992	63
		16.2.3	Brenda Process Technology - 1992	64
		16.2.4	Nickel Plate Mine (NPM) - 1992	65

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		16.2.5	Andre LaPlante – 1993	66
		16.2.6	G&T Metallurgical Services – Report KM 2121 – 2008	66
	16.3	Summary of Historical Metallurgical Testing		66
	16.4	Current Testing Program – G&T Metallurgical Services (2009 – 2010)		67
	16.5	Testing of Processing Plant Products for Acid Generating Potential		68

17.	Mineral Resources and Mineral Reserve Estimates			70
	17.1	Geology Model		70
	17.2	Evaluation of Extreme Grades		75
	17.3	Compositing		77
	17.4	Block Model Definition		78
	17.5	Bulk Density Values		79
	17.6	Variography		80
	17.7	Grade Model and Interpolation Plan		80
	17.8	Model Validation		81
		17.8.1	Visual Inspection	81
		17.8.2	Global Means	81
		17.8.3	Swath Plots	82
	17.9	Resource Classification and Summary		84
	17.10	Comparison of Updated Mineral Resource Estimation and Historical Resource
		Estimations		90
	17.11	Mineral Reserves		90

18.	Other Relevant Data and Information – Preliminary Economic Assessment			91
	18.1	Introduction		91
	18.2	Mining Method		91
		18.2.1	Pit Wall Slope Angle	92
		18.2	Ore Loss and Dilution	92
		18.2.3	Waste Rock Disposal	92
	18.3	Processing		93
		18.3.1	Process Flow Sheet	93
		18.3.2	Design Criteria	95
		18.3.3	Tailing Storage Facility	95
	18.4	Infrastructure		95
		18.4.1	Power Supply	95
		18.4.2	Water Supply	95
		18.4.3	Assay Laboratory	96
		18.4.4	Miscellaneous Infrastructure	96
	18.5	Product Marketing		96
	18.6	Operation Cost Model		96
		18.6.1	Mining	96
		18.6.2	Processing	97
		18.6.3	Administration and Overheads	98
	18.7	Open Pit Optimisation		98
		18.7.1	Introduction	98
		18.7.2	Block Model	99
		18.7.3	Input Parameters	99
		18.7.4	Optimisation Results	99
	18.8	Production Scheduling		100
		18.8.1	Alternative approaches considered	100
		18.8.2	Scheduling Results	101
	18.9	Operating costs		102
	18.10	Capital Costs		102
		18.10.1	Mining Capital Costs	102
		18.10.2	Processing Capital Costs	103
		18.10.3	Offsite Infrastructure	103
		18.10.4	Other	103

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	18.11	Financial Modelling		104
		18.11.1	Modelling Results	104
		18.11.2	Sensitivities	104
		18.11.3	Project Risks	105
	18.12	Other Cases Considered		106

19.	Interpretation and Conclusions			109
	19.1	Geology and Resource Estimation		109
	19.2	Mining		109
	19.3	Mineral Processing and Metallurgy		109
	19.4	Environmental Considerations		110
	19.5	PEA Conclusion		110

20.	Recommendations			111
	20.1	Geology and Mineral Resources Estimates		111
	20.2	Mining		112
	20.3	Mineral Processing and Metallurgy		113
	20.4	Infrastructure		113
		20.4.1	Overall Project Layout	113
		20.4.2	Water Supply	113
		20.4.3	Electrical Power	113
		20.4.4	Environmental Matters	114
	20.5	Cost of Recommended Work Program		114

21.	References			116

22.	Date and Signature Page			119

23.	Additional Requirements for Technical Reports on Development and Production Properties			125


List of Tables

Table 1-1:	Mineral Resources for the B and WD Veins at the Elk Project	3
Table 1-2:	Key assumptions and results for Base Case and $1200 Case studies	5
Table 1-3:	Summary of recommended tasks to complete PFS	6
Table 2-1:	Summary of Qualified Persons	8
Table 4-1:	List of titles with areas	14
Table 6-1:	Historical open pit and underground mine production	25
Table 6-2:	Resource Estimate, Giroux Consultants, 2004	26
Table 6-3:	Resource Estimate, Giroux Consultants, 2007	27
Table 10-1:	Summary of exploration work completed from 1986 to 2007	33
Table 10-2:	Significant Intersections in Bullion Creek Area	35
Table 10-3:	Significant drill hole intersections in Siwash East area	36
Table 10-4:	Significant drill hole Intersections in Gold Creek West area	37
Table 10-5:	Significant drill hole intersections in Gold Creek East area	38
Table 10-6:	Significant drill hole intersections in Lake Zone area	40
Table 10-7:	Significant drill hole intersections in the Great Wall area	40
Table 10-8:	Significant drill hole intersections in the End Zone area	41
Table 10-9:	Significant drill hole intersections in the Discovery Showing area	42
Table 10-10:	Significant drill hole intersections in the South Showing area	43
Table 11-1:	Drilling completed from 1986 to 2007	45
Table 11-2:	Difference between Collar Coordinates in the Elk Project Database and the 2010 Surveyed Coordinates	47

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Table 14-1:	List of SRM used at the Elk Project		55
Table 16-1:	Processing comparison		68
Table 16-2:	Test results from ABA testing of Flotation Concentrates and Tailings		69
Table 17-1:	2007 drill holes in the Elk Deposit Resource Estimate		74
Table 17-2:	Summary of capping strategy and number of assays capped by zone		76
Table 17-3:	Preliminary variogram models for the 1300 Vein of the B Vein Complex		80
Table 17-4:	Block Model interpolation parameters for B and WD Vein Au grades		81
Table 17-5:	Comparison of ID2, OK and NN Global Mean values		82
Table 17-6:	Inputs used to build the pit shell for low grade and high grade cut-offs Mineral Resources estimation tabulation of open pit		85
Table 17-7:	Mineral Resources for the B and WD Veins at the Elk Project, reported using 0.5 g/t Au for blocks within the Resource reporting pit shell and 5.00 g/t Au for blocks below the pit shell, 7 September 2010		88
Table 17-8:	Resource Estimation reported at various Au g/t cut-offs for comparative purposes		89
Table 17-9:	2010 Updated Resource Estimate Results for the B and WD Veins reported at a Global 1 g/t cut-off for comparative purposes only	90
Table 17-10:	2009 Updated Resource Estimate Results for the B and WD Veins reported at a Global 1 g/t cut-off for comparative purposes only	90
Table 17-11:	B and WD Vein Resource Estimate Results reported at a Global 1 g/t cut-off (includes both near surface and deeper grade blocks) (Giroux, November 2007)		90
Table 18-1:	Simplified design criteria		95
Table 18-2:	Mining Supervision and Overhead costs		97
Table 18-3:	Commination, gravity and flotation operating costs		97
Table 18-4:	Cyanidation and electrowinning operating costs		98
Table 18-5:	Administration and overheads costs		98
Table 18-6:	Whittle optimisation parameters		99
Table 18-7:	Open pit mining schedule for Chosen Case (RF 0.70)		101
Table 18-8:	Unit Operating costs		102
Table 18-9:	Mining capital cost estimate		102
Table 18-10:	Processing Capital Cost Estimate		103
Table 18-11:	Project Financial Indicators		104
Table 18-12:	Two-parameter sensitivity analysis		104
Table 18-13:	Base case outcomes with $US1200/tr.oz gold price		106
Table 18-14:	Pit Shell at $US1200 Au/tr.oz ($US1200 case) Financial Indicators		106
Table 18-15:	Details of $US1200 case		107
Table 20-1:	Cost of PFS Work Program		114


List of Figures

Figure 4-1:	Property location and geology		11
Figure 4-2:	Location of Agur Option Area and other prospective zones		12
Figure 4-3:	Site plan of existing facilities		13
Figure 4-4:	Titles plan		14
Figure 9-1:	Location map of the Exploration Targets identified on the Elk Project to date		32
Figure 10-1:	Elk Property vein locations		34
Figure 10-2:	Anomalous gold mineralization location map for Bullion Creek area		35
Figure 10-3:	Anomalous gold mineralization location map for Siwash East area		36
Figure 10-4:	Anomalous gold mineralization location map for Gold Creek West area		37
Figure 10-5:	Drill hole location map for Gold Creek East area		38
Figure 10-6:	Anomalous Gold mineralization location map for Lake Zone East area		39
Figure 10-7:	Detailed results of exploration completed at the Great Wall Zone, End Zone and the Discovery Showing (North Showing) areas		41

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Figure 10-8:	Anomalous Gold mineralization location map for South Showing area	42
Figure 10-9:	Anomalous gold mineralization location map for Elusive Creek area	44
Figure 11-1:	Plan of 2010 Infill drilling completed in November 2010	46
Figure 14-1:	SRM CDN-GS-10A	55
Figure 14-2:	SRM CDN-GS-30A	56
Figure 14-3:	SRM CDNGS-6	57
Figure 14-4:	SRM CDN-GS-8	57
Figure 14-5:	Scatter plot of core duplicates for 2000 to 2006 drill holes fire assay results (Au g/t)	58
Figure 14-6:	Blank samples for Elk Project	59
Figure 14-7:	Acme Laboratory pulp re-assay and reject re-assay results for 2000 to 2006 samples	59
Figure 14-8:	Acme resubmission duplicate results (Au g/t)	60
Figure 14-9:	Scatter plot of ACME Original FA against the Chemex Laboratory Check FA (Au g/t)	61
Figure 14-10:	Scatter plot of ACME Original FA plotted against the ACME blind submissions (Au g/t)	61
Figure 17-1:	Plan view of B vein domains	71
Figure 17-2:	Long section view of B vein domains - looking south	71
Figure 17-3:	Section view of B vein domains (looking east)	72
Figure 17-4:	Section view of B vein domains (looking west)	72
Figure 17-5:	Plan view of WD vein domains	73
Figure 17-6:	Long section view of WD Vein Domains (looking south)	73
Figure 17-7:	Box plot and summary statistics for the B vein domains	74
Figure 17-8:	Box plot and summary statistics for the WD Vein domains (2000s), Waste domains (900 and 0)	75
Figure 17-9:	Cumulative probability plot for gold assay data 1300 Domain (B Vein)	76
Figure 17-10:	Cumulative probability plot for gold assay data 1400 Domain (B Vein)	77
Figure 17-11:	Cumulative probability plot for gold assay data 2500 Domain (WD Vein)	77
Figure 17-12:	Cumulative probability plot of Whole Vein composites (m)	78
Figure 17-13:	Cumulative probability plot of Whole Vein composites (Au g/t)	78
Figure 17-14:	Block Model definition for the B and WD Vein Block Model	79
Figure 17-15:	Swath plot for 1300 Domain of the B Vein - Eastings - Au (g/t)	83
Figure 17-16:	Swath plot for 1300 Domain of the B Vein - Northings - Au (g/t)	83
Figure 17-17:	Swath plot for 1300 Domain of the B Vein - Elevations - Au (g/t)	84
Figure 17-18:	B Vein with classification blocks showing location of Measured, Indicated and Inferred Mineral Resources and location of Pit Shell used to declare Open Pit and Underground extraction	85
	methods - Longsection looking north
Figure 17-19:	B Vein with classification blocks showing location of Measured, Indicated and Inferred Mineral Resources and location of Pit Shell used to declare Open Pit and Underground extraction methods and location of the underground decline - Longsection view from below the deposit, looking south	86
Figure 17-20:	WD Vein with classification blocks showing location of Indicated and Inferred Mineral Resources and location of Pit Shell used to declare Open Pit and Underground extraction methods - Longsection looking north	86
Figure 17-21:	Plan View of B and WD Veins and Resource Estimation Pit Shell to segregate Open Pit potential and Underground potential for extraction method	87
Figure 18-1:	Initial Conceptual Project Layout $US1200 case	93
Figure 18-2:	Preliminary Process Flow	94
Figure 18-3:	Whittle results – graph of tonnes by revenue factor	100
Figure18-4:	Base case Pit Shell (green) with included mineralization (red)	101
Figure 18-5:	$US1200 Pit Shell and Base Case pit shell with included mineralization (red)	108

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List of Appendices

Appendix 1:	Base Case Financial Projection	(Gold at $US1000/tr.oz)
Appendix 2:	Base Case Financial Projection	(Gold at $US1200/tr.oz)
Appendix 3:	$US1200 Case Financial Projection (Gold at $US1200/tr.oz)
Appendix 4:	Summary of ABA testing on Mine Waste Rock

SRK Consulting

ALM003 Elk Gold Project

NI 43-101 Technical Report

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1.	Summary
1.1	Report Purpose

Almaden Minerals (Almaden) wishes to provide information to shareholders concerning the economic viability of the Elk Gold Project, and this report is a means to achieve this. This report entitled “NI 43-101 Technical Report for a Preliminary Economic Assessment on the Elk Gold Project, Merritt, British Columbia, Canada” presents the results of a preliminary evaluation that incorporates the updated resource estimation (2010), discussion relating to mineral processing, and the results for the Preliminary Economic Assessment (PEA).

1.1.1 Property Location and Ownership

The property is within the Similkameen Mining District and consists of 27 contiguous mineral claims and one mining lease covering 16,566 hectares. Except for the Augur Option block, Almaden has a 100% interest in all claims. A 1% NSR production royalty is payable on production from the Agur Option block, located approximately 4 km south of the area of estimated resources, and is not relevant to this report.

1.1.2 History

Prospecting activities date back to the early 1900s but detailed work in the area began in 1960s and 1970s by several companies who focused on copper and molybdenum. Fairfield Minerals investigated the area for gold in 1986. Approximately 51,500 ounces of gold were produced between 1992 and 1995 from a test pit and underground mining exploration ( a decline was excavated for underground drilling activity, but no stoping or underground mining occurred). Almaden Resources Ltd. amalgamated with Fairfield Minerals in 2002 to form Almaden Minerals (Almaden), which thereby became the sole owner of the Elk gold property and has continued exploration activity on the project since this time.

1.1.3 Geology and Mineralization

The Elk property lies within the intermontane tectonic belt of south-central British Columbia. Nicola group andesites and massive basalts of Upper Triassic age cover the western part of the project area and the NNE-trending Middle Jurassic granites and granodiorites of the Osprey Lake Batholith cover the eastern part.

The B and WD Vein Complex (historically called the Siwash Vein) is emplaced within a fault / fracture zone that strikes east-northeast and dips moderately to steeply southward. Most of the previous mine production occurs within the granodiorite border phase of the batholiths (Lewis, 2000).

Gold mineralization occurs within quartz-sulphide veins and stringers most often within altered granite and occasionally within the adjacent volcanics. Pyrite is the most common sulphide (Conroy, 1994), ranging from 5 to 80% with higher grades associated with chalcopyrite and tetrahedrite. Mineralization occurs as fine grained native gold (typically less than 50 microns) in quartz, in quartz-pyrite boxwork, and in fractures within veins (King, 2001). Gangue minerals include quartz and altered wall rock clasts, with minor amounts of ankerite, calcite, barite and fluorite.

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1.1.4

Exploration

Exploration has been focused on finding additional gold-bearing quartz veins on the Elk property for more than 20 years. Three companies are primarily responsible for this work and the project has benefited from overlap of personnel from the late 1980s to 2007. The work completed throughout the property area includes 17,400 soil surveys, 5,728 lithology geochemical samples, geological mapping of 4,110 hectares, more 8,000m of trenching, about 16 km of road building,1,824 km of legal surveys, 121 km of magnetic surveys, 4.5 km of IP surveys and 1.8 km of UTEM surveys.

There are ten identified exploration targets beyond the main North B and WD Complex including Bullion Creek, Siwash East, Lake Zone, Great Wall Zone, End Zone, Gold Creek West, Discovery Showing (North Showing), South Showing, Elusive Creek and Agur Option. These exploration targets have had varying levels of work completed, including geophysical surveys, trenching, chip sampling and 68 diamond drill holes have been completed for a total of 5,111 m. Positive results from the existing work suggest potential to find additional resources beyond the B and WD Vein Complex in the project area.

1.1.4.1 2010 Drilling Program

Almaden completed a diamond drill program in November 2010 that consisted of 87 drill holes in the resource area, for a total of 12,749 m. The purpose of the program was primarily to provide infill drill holes to increase the drill hole density in the areas of Inferred Mineral Resources. The drill hole locations were selected in consultation with LGGC. A new camp was permitted and constructed for the geological and support staff. Also in 2010 aerial photography and Lidar planimetric mapping was conducted by Eagle Mapping over the property in order to update the base map for the general project area. The finished product includes one meter contours and a 24 cm pixel orthophoto (1:20,000 scale color photography). Control for this survey and 2010 drill hole locations were surveyed by BC registered land surveyor.

Drilling for 2010 was supervised by Mr R. Brian Alexander, P.Geo and Qualified Person as defined by NI 43-101. Mr Alexander acted as QP for the technical information released by the company relating to the 2010 drill program in Almaden news releases dated 9 September and 21 October.

At the conclusion of the drill program, a BC registered land surveyor completed the survey of the 2010 drill collar location coordinates. The surveyor was asked to check the coordinates of 25 older holes in the area of the resource estimate. Upon review of the data differences in 16% of the X and Y coordinates were found of the historical drill holes and a difference between the mine grid elevation (reference grid used for locating all drill collars prior to 2010) and the UTM elevation was discovered. The difference may be a result of two different datum points being used.

The mine grid was established using a surface datum or survey control point and was assigned an estimated elevation from 1:50,000 topography map. The same datum was used to close all surveys, prior to 2010, of both the surface and underground drill hole collar coordinates. That datum was labeled as CP-1.

In 2010, the legal land surveyors chose an alternate datum or control point. This was labeled originally as LSM2N1E and later renamed BCLS 356. The 2010 aerial photography, Lidar planimetric mapping survey and the check of the 25 historical drill holes were all surveyed using BCLS 356.

The X and Y coordinates for 4 of the 25 historical holes were different from that included in the project database by 2.2 to 8 m. The UTM coordinates in the historic data were derived from a mathematical conversion of mine grid coordinates. This mathematical conversion may not be correct. It is recommended that a qualified surveyor be commissioned to survey all the surface drill hole locations into the UTM coordinate system prior to updating the current resource estimation or undertaking a pre-feasibility study on the Elk Project. A complete resurvey of the underground drill holes may be too difficult to undertake given that the workings are flooded but the currently accessible survey stations in the decline should be included in the resurvey program and a reliable conversion factor be determined to convert the underground drill holes to the UTM coordinate system. The survey must also reconcile the elevation difference in the two separate datums used, as this will allow the integration of the aerial photography and Lidar planimetric mapping. The difference in survey elevations between the mine grid and the UTM coordinate system does not impact either the current mineral resource estimation or the PEA study as all data included in these studies were located using the mine grid only.

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ALM003 Elk Gold Project

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ALM003 Elk Gold Project

NI 43-101 Technical Report

January 2011

The combined Measured and Indicated mineral resources for the B and WD veins, both in the pit shell (reported at 0.50 Au g/t) and below the pit shell (reported at 5.00 Au g/t) are estimated to be 2.2 Mt with a gold grade of 4.26 g/t for 300,000 contained ounces and Inferred mineral resources reported with the same criteria of 1.2 Mt with a gold grade of 7.13 Au g/t and an estimated 263,000 contained ounces.

1.1.6 Preliminary Economic Assessment

In May 2010, Almaden commissioned SRK (Australasia) Pty Ltd to complete a PEA to define an economic mining scenario for the Elk Project using an open pit methodology whilst also identifying areas of risk.

This included:

Developing an appropriate cost model for a suitable operation Open pit optimisation to define an appropriate in-pit inventory Conceptual scheduling and financial modelling.

This mining study and PEA are at a conceptual level where different options can be considered and a broad understanding of the potential project performance can be gained.

SRK and Almaden consider this project to be preliminary or “green field” in nature as previous mining activity on the property was largely for exploration purposes and the property has not been the subject of a detailed pre-feasibility study as is defined in NI 43-101. The trial open pit operation that occurred in the 1990s mined a small portion of vein material. The scenario developed as a recommendation for further study is summarised in the following table. Known as the Base Case, it is based on using augmented process equipment that Almaden already owns. This limits the throughput to 500 tpd.

The Base Case is a conservative and low risk scenario in the light of the current gold price, and in practice the project could be expanded to mine a much larger part of the known resources if current gold prices are sustained. To show the effect of this, an alternative case known as the $US1200 case was also studied.

The $US1200 case assumes that a gold price of $US1200/tr.oz will be maintained for eight years. The mine processing plant production is doubled, to 1000 tpd.

The underground resources declared in Table 1-1 are not considered for production in this report. It is believed that if the project proposed goes ahead, then these resources will have a much improved chance of being mined, because access can be gained from within the open pit, and because the treatment plant will have been built, and will be ready to accept underground production without further capital expense. This matter can therefore safely be left for consideration at a later time.

The key details of both cases are given in Table 1-2.

The reader is advised that this report contains an economic assessment that is preliminary in nature and includes inferred mineral resources. These are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorised as mineral reserves, and there is no certainty that the preliminary assessment will ever be realised, in whole or in part.

Table 1-2 shows the quantities of feed processed during the scenarios under study. These are all drawn from resources declared in Table 1-1 with allowance for mining dilution. It is emphasized that these do not constitute a reserve, being based in part on Inferred Resources and estimated from preliminary data.

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Table 1-2: Key assumptions and results for Base Case and $1200 Case studies

Project summary	Base Case	$1200 Case		Unit
Assumed gold price	1000	1200	$US/tr.oz
Tonnes per day treated	500	1000	tpd
Life	7	9	years
Total tonnes treated	1.1	2.6	MT
Grade	4.14	3.89	g/t
Waste: Ore ratio	16.4	30.1
Plant recovery	92	92		%
Ounces Au produced	139,198	297,23	Tr.oz
Initial capital expense	9.91	17.50	$CADM
Working and preproduction capital	2.27	9.60	$CADM
Waste mining	2.42	1.90	$ CAD/tonne waste
Ore mining	8.38	5.87	$CAD	/tonne ore
Processing	20.68	14.74	$CAD	/tonne ore
Administration and overheads	2.07	1.27	$CAD	/tonne ore
Total operating cost	70.30	78.91	$CAD / tonne / ore
Pre-tax NPV @ 8%	28.7	67.9	$CADM
Pre-tax IRR	51%	39%
Max Exposure	13.66	33.53	$CADM
Payback, years from start production	1.85	3.30	years
ratio, gross earnings: max exposure	5.02	6.00
ratio, NPV: max exposure	2.10	2.03

1.1.7 Qualified Person s Conclusions and Recommendations

Overall, SRK believes that this preliminary study demonstrates that a viable project could be launched on the Elk property, and therefore, that further work is justified.

The favoured strategy is to start production using the currently- owned processing plant augmented as required, along the lines of this document s Base Case. This should minimise capital costs and time for procurement. Once production is established, the ongoing outlook for the gold price will determine when an expansion goes ahead. The above execution plan is to be preceded by a Prefeasibility Study (PFS); which is the next step. Specific recommendations for ongoing work are included in this report. Table 1-3 is a simplified task list for the proposed study. The estimated cost of the items in the list is $CAD5.4M, which does not include Almaden s overhead costs in the period covered. Seasonal conditions and permitting uncertainties make it difficult to estimate the time required for this program.

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Table 1-3: Summary of recommended tasks to complete PFS


Work Item		Cost K$CAD
Design a preliminary project layout		15
Complete Environmental baseline studies		150
Consult with first Nation Groups		70
Apply for permission to reroute and upgrade access road		25
Obtain Permission for powerline to site		50
Initiate a drilling Program to:-
Locate ,quantify and permit project water supply		100
Determine slope angles in the proposed pit, define vein strength and friability and separation from walls		250
Convert Inferred to Indicated Resources by Infill Drilling		1600
Obtain further waste bulk densities and conduct further ABA testing		20
Obtain samples of early production for metallurgical testing		150
Condemnation drill Program for waste dump, TSF, plant site		1500
Update Mineral Resources estimate		150
Make early-years confirmatory metallurgical tests. Update plant parameters, design flow sheet		30
Survey existing processing equipment, make plant layout and cost estimate for plant		300
Complete surface rainfall/runoff study		25
Design tailings disposal facility, estimate cost		50
Obtain initial mining costs		30
Reoptimise pit and schedule pit and dumps, study the integration of underground mining into the overall
mining plan, adjust dumps and surface layout		370
Obtain final costs iterate planning		150
Design and estimate site infrastructure		100
Investigate marketing of doré vs. flotation concentrate and estimate costs		50
Document PFS		250
	Total Costs K$CAD	5435

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Introduction

Almaden Minerals (Almaden) wishes to provide information to shareholders concerning the updated mineral resource estimation and preliminary economic viability of the Elk Gold Project. This report entitled “NI 43-101 Technical Report for a Preliminary Economic Assessment on the Elk Gold Project, Merritt, British Columbia, Canada” presents the results of the preliminary evaluation that incorporates the updated resource estimation (2010), discussion relating to mineral processing, and the results for the Preliminary Economic Assessment (PEA).

The Elk Project (formerly known as the Siwash Project) is located 45 km southeast of Merritt, southern British Columbia, Canada. To date, eight gold bearing veins, in four groups, have been identified in the B and WD Vein Complex that is the subject of the mineral resources and the PEA documented in this report.

Between 1992 and 1994, a small open pit and underground exploration operation on the B vein produced just over 50,000 oz of gold to a depth of 40 m from a bulk sample of 16,500 tonnes (averaged grade of material was 97 g/t Au).

Since the cessation of the explorative mining, additional exploration work, including diamond drilling, trenching, geophysical studies and surface sampling has led to multiple mineral resource estimations for the B and WD veins; most recently in 2010.

2.1 Terms of Reference

This Technical Report includes an updated Mineral Resource estimation and presents the findings of a Preliminary Economic Assessment (PEA) of the Elk Project. It has been prepared to support Almaden Minerals Ltd s disclosure relating to the PEA as defined by item 4.2 (j) (i) of National Instrument (NI) 43-101. As such, the intended audience for the Report are all the Stakeholders of the Elk Project.

2.2 Report Authors

In the fall of 2008, LGGC was commissioned by Almaden to complete an updated geology model of the B and WD Vein Complex and estimate a mineral resource that included drilling data gathered in 2007. The mineral resource estimate was completed in September, 2009. In September 2010, LGGC updated the 2009 mineral resource estimate using larger block sizes to support a larger open pit extraction mining method, as the previous estimate was built to support a potential small open pit and expanded underground operation. The 2010 mineral resource estimate was completed in September, 2010 and was used to support the PEA study completed by SRK in Perth, Australia. GEMS®, a commercially available exploration and mining software package, was used to estimate the mineral resources.

Susan Lomas (P.Geo.) of LGGC reviewed the pertinent geological data in sufficient detail to support the data incorporated into the updated mineral resource estimation. Susan Lomas carried out a site visit to the Elk Project on June 23, 2009. While at site, a general review of the project was provided by Mr Morgan Poliquin, President and CEO of Almaden and drill core collection, sampling procedures and existing drill core relative to the computer database were reviewed. Samples for independent assaying were not collected during the site visit. Susan Lomas is the Independent Qualified Person (QP) for the mineral resource estimate as defined by NI 43-101.

Several source documents covering geology, previous exploration work and drilling were used to prepare this report including yearly drilling and assessment reports by Mr Wojtek Jakubowski, P.Geo., and Qualified Person under the rules of NI 43-101, a geological report prepared by Mr Leo King, P.Geo. (2001), and a Resource update report by Mr Gary Giroux, P.Eng., and Qualified Person (2007).

In November 2010, Almaden completed 12,749 m of infill diamond drilling in 87 drill holes. The drilling was completed in the area of the mineral resources and results for most of the drill holes are still pending at the time of issue of this report. Drilling for 2010 was supervised by Mr R. Brian Alexander, P.Geo and Qualified Person as defined by NI 43-101. Mr Alexander of Alexploration Inc. acted as QP for the technical information released by the company relating to the 2010 drill program in Almaden news releases dated 9 September and 21 October.

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Metallurgical testing and the provision of preliminary capital and operating costs for a treatment plant was undertaken by Mr Gary Hawthorn, P.Eng a senior metallurgical consultant with Westcoast Mineral Testing Inc.

In May 2010, Almaden commissioned Mr Roger Pooley, of SRK Consulting (Australasia) Pty Ltd (SRK) to complete a PEA to define potential economic mining scenarios for the Elk Project using open pit methodology only, whilst also identifying areas of risk and opportunity. SRK relied on the other authors named in Table 2-1 for data in the areas of Geology, Resources, and Processing.

The authors have written sections of this Technical Report and are acting as the Qualified Person (QP) for those sections. Roger Pooley is responsible for the mining section and the financial analysis and has also compiled the Technical Report into a single document, suitable for SEDAR filing. A summary of the authors is shown in Table 2-1 which includes their level of personal inspection of the Project.

Table 2-1: Summary of Qualified Persons

Name		Company	Qualifications and Affiliation	Sections authored	Personal inspection
				1.1, 1.1.6, 1.1.7, 2, 3, 4, 5,
				18.1, 18.2, 18.4, 18.5, 18.6,
			BSc (Eng),	18.6.1, 18.6.3, 18.7, 18.8,
Roger Pooley		SRK		18.9, 18.10, 18.11, 18.12,	Nil
			MAusIMM	19.2, 19.4, 19.5, 20.2, 20.4,
				20.5, 21, 22 and 23, part
				Table 20-1, Appendices 1-4
				1.1.1, 1.1.2, 1.1.3, 1.1.4,	Site visit 23 June 2009 to
				1.1.5, 6, 7, 8, 9, 10, 11, 12,	inspect core, review
Susan Lomas		LGGC¹	P.Geo. (BC)	13.1, 13.1.1, 13.1.2, 13.2	sampling procedures and
				and 13.3, 14, 15, 17, 19.1	gain an overview of the
				and 20.1, part of Table 20-1	project
Gary Hawthorn		WMT²	P. Eng (BC)	16, 18.3, 18.6.2, 19.3, 20.3	Nil
		Alexploration			On site July - November
R. Brian Alexander			P Geo (BC)	1.1.4.1, 11.1.1 and 13.1.3
		Inc.			2010


2.3	Data Sources

The data used in the preparation of the work contained in this Technical Report are appropriately described and referenced in the related sections.

2.4 PEA Scope

In summary, the objective of the PEA was to define a potential financial scenario for the Elk Project using an open pit methodology whilst also identifying areas of risk and upside.

This included:

Developing an appropriate cost model for the operation

Open pit optimisation with some sensitivity analysis to examine risk Conceptual scheduling and financial modelling.

The final element was to incorporate the PEA into a suitable Technical Report.

¹Lions Gate Geological Consulting Inc.
²Westcoast Mineral Testing Inc.

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2.5	Effective Date
The effective date of this Technical Report is 10 December 2010.


2.6	Units of Measure

SI (metric) units of measurement are used throughout this Technical Report. All monetary values refer to Canadian dollars (CAD) unless specifically stated otherwise. All maps in this report are on projection UTM Zone 10-NAD83.

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Reliance on Other Experts

In compiling this report, the QP s have relied on material or reports from experts who are not QPs as follows: Section 4.4.2 Related Environment Studies which concerns legal/regulatory, environmental, and socio-political issues as per NI 43-101 F1 Item 5; and which were provided by Marke Wong, E.P. (Canada) Environmental Consultant, MRI Management Consultants. This section is summarised in Section 19.4 and recommendations based on it are in Section 20.4.4.

Sections 4 and 5 geographical data was provided by Mr Marke Wong.

Status on mineral claims and leases was provided by Mr Marc Blythe, P.Eng (BC) VP Mining of Almaden and Ms Dione Bitzer, Controller of Almaden. SRK has not checked on the validity of these claims and leases.

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4.	Property Description and Location
4.1	Property Location

The Elk Property is located in southern British Columbia, Canada approximately 325 km northeast of Vancouver and 55 km west of Okanagan Lake, approximately midway between the towns of Merritt and Peachland, as indicated in Figure 4-1. The property is at latitude 49º50 N and longitude 120º19 W and can be found on National Topographic Sheets (NTS) 082E071, 092H078 to 80 and 092H089.

Figure 4-1: Property location and geology

4.2 Tenement Details

The property is within the Similkameen Mining District and consists of 27 contiguous mineral claims and one mining lease covering a total of 16,566 hectares. Almaden owns 100% of the claims. The Agur Option block is subject to a 1% NSR payment on production from it. This block is over 3 km south of the Mineral Resources estimated in this Report; see Figure 4-2: showing the existing trial open pit location, while the old waste dumps are shown in Figure 4-3.

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Figure 4-2: Location of Agur Option Area and other prospective zones

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Figure 4-3: Site plan of existing facilities

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Figure 4-4: Titles plan

The claim block may be maintained by continuing to conduct work on the property, or by payment in lieu. The expiry date of each claim is listed in Table 4-1. The small mines permit may be maintained by providing an annual reclamation report on the property that is acceptable to the Ministry of Energy, Mines and Petroleum Resources.


Table 4-1: List of titles with areas

Claim	Type	No units	Record no	Expiry date	BCGS	Hectares
ELK06A	Cell	24	524944	1/12/2011	092H089/88	500.07
ELK06B	Cell	24	524945	1/12/2011	092H089/88	499.90
ELK06C	Cell	24	524946	1/12/2011	092H089/88	499.73
ELK06D	Cell	24	524947	1/12/2011	092H089/88	499.56
ELK06E	Cell	24	524948	1/12/2011	092H089	499.56
ELK06F	Cell	24	524949	1/12/2011	092H089	499.73
ELK06G	Cell	13	524950	1/12/2011	092H089	270.75
ELK06H	Cell	13	524952	1/12/2011	092H089/99	520.33
ELK06I	Cell	24	524954	1/12/2011	092H099/98	499.43
ELK05A	Cell	1	516781	1/12/2016	092H089	20.85
ELK05B	Cell	2	517116	1/12/2016	092H089	41.65
No Name	Cell	1	517045	1/12/2016	092H079	20.86
No Name	Cell	25	516717	1/12/2016	092H089	520.57

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Claim	Type	No units	Record no	Expiry date	BCGS	Hectares
No Name	Cell	30	516725	1/12/2016	092H089	624.98
No Name	Cell	25	516727	1/12/2016	092H089	521.05
No Name	Cell	25	516731	1/12/2016	092H089/79	521.26
No Name	Cell	71	516732	1/12/2016	092H079	1,481.07
No Name	Cell	45	516733	1/12/2016	092H089	938.03
No Name	Cell	30	516739	1/12/2016	092H089	624.69
No Name	Cell	70	516740	1/12/2016	092H089	1,458.28
No Name	Cell	8	516743	1/12/2016	092H089	166.61
No Name	Cell	61	516750	1/12/2016	092H089	1,271.49
No Name	Cell	57	516755	1/12/2016	092H079/89	1,188.84
No Name	Cell	49	516757	1/12/2016	092H079/89	1,021.84
No Name	Cell	54	516759	1/12/2016	092H089/88	1,125.59
No Name	Cell	30	516761	1/12/2016	092H089	625.03
No Name	Cell	5	519105	1/12/2016	092H079	104.30
Siwash North	lease	1	308695	9/14/2027	092H089
			Total Area			16,566.05

4.3	Royalties and Taxes
4.3.1	Royalties

The private royalty on the Agur Option area is disclosed in Section 4.2. production scenarios proposed in this document.

4.3.2 Taxes

It has no significance to the

Federal and provincial income taxes would be expected to apply to Almaden. The rates are of the order of 19% of net taxable profits by federal government, and 11% of net taxable profits by the provincial government. There are many offsetting credits. These matters were not considered in this preliminary assessment.

4.4	Environmental Considerations
4.4.1	Background

Characterization of existing environmental conditions is an important component of the risk management and permitting process for the Elk Gold Project. Environmental monitoring studies were begun at Elk Gold in 1992. This data includes surface water quality that provides a long term record of the relative chemical stability of the project area. Baseline studies to date included long-lead time baseline studies including climate, hydrology, water quality, fish & aquatic studies; as well as community consultation activities. Environmental and socio-economic data was summarized by Golder (2007)3 and more recently Almaden Minerals (2010)4. Environmental baseline studies and community consultation activities are ongoing at the Elk Gold Mine in preparation for site development and related permitting; and are summarized in the following sections.

³	Golder, 2007 “Report on Review and Data Gap Analysis of Environmental Baseline Studies, Elk (Siwash) Gold Mine” Prepared for Almaden Minerals.
4	Almaden Minerals, 2010. Siwash Gold Deposit Elk Gold Project Reclamation Report 2009, Prepared for the Ministry of Mines, Similkameen Division.

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4.4.2	Related Environmental Studies
4.4.2.1	Physical Resources

Climate, Air Quality and Noise

Characterization of the microclimate at the project site is required to allow for simulation of long-term records. These data are required for engineering designs for water balance, surface water management, waste management and water quality predictions.

Climate data for 1971-2000 is available from the nearby Environment Canada climate station No. 1126077-Peachland Brenda Mines located 25 km east of the site. An on-site automated weather station was installed at Elk Gold Mine on September 21, 2006 to collect site specific data for temperature, rainfall, wind speed and direction. Selected climate parameters for design of a water storage dam by Klohn Crippen Berger (2010)⁵reported an average annual precipitation of 653 mm, average annual lake evaporation 560 mm.

The project area is located in a relatively remote area and air and noise quality are assumed to be good with low levels of particulates and contaminants; and low levels of noise. Almaden plans to collect site specific data for air and noise quality during baseline studies in 2011.

Geology & Geochemistry

Environmental considerations from geology and geochemistry are related primarily to the assessment of acid rock drainage (ARD).

Past ARD testing has been limited; however, operational evidence suggests that the large majority of the unaltered waste rock in the existing dumps has good neutralization potential while only a small proportion of the waste rock immediately adjacent to the vein, has shown moderate levels of sulphur and Potentially Acid Generating (PAG) material. Additional testing for Acid Base Accounting (ABA) of representative samples of waste rock from the open pit area was completed in the fall of 2010, and did not show any acid generating potential, refer to Appendix 4.

Results of early metallurgical testing indicate that flotation tailings would not be expected to be acid generating. However, cyanide leaching of flotation concentrate will produce tailings that are potentially acid generating⁶. Trace metal analysis of simulated tailings MN2-049 showed elevated concentrations of arsenic, copper, silver and zinc.

Hydrology

Siwash Creek lies in the headwaters of Hayes Creek in the Similkameen-Okanagan-Columbia River System. Instream flow data is required to support various engineering studies including stage discharge curves, site run-off management plans, withdrawal and discharge limits, water balances; and to establish minimum flows for aquatic life.

The past hydrology monitoring program was initiated in 1992 and established 14 stations including weirs, staff gauges and flow monitoring stations. In 2004 a Global Water data logger was installed on Siwash Creek downstream of the Bullion Creek confluence and in 2006 another was installed downstream of the Don Creek confluence . Baseline hydrology studies through 2011 are being conducted by Knight Piesold Ltd (2004).

Hydrogeology

Hydrogeological information is limited to historical observations of the pit and mine workings; and a few wells. The open pit and underground workings hold approximately 43,000 m³of water and took an estimated 19 months to fill (Jakubowski, 2004). Groundwater seepage is estimated at 15 L/min.

⁵	Klohn Crippen Berger, 2010
⁶	Bacon Donaldson, 1992b, as outlined in Fairfield Minerals (1993) Siwash North Gold deposit – Compilation of Metallurgical Results, May 15, 1993.

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There are three known wells in the area: the 11 m-deep “Portal Well” drilled onsite in 1993-94 and two wells (152 and 76 m depth) at the Lodge located 2 km to the east. The yields of the two wells are typical of deep bedrock wells in the area, and indicate very low permeability.

Additional hydrogeological studies are currently planned to support tailings storage facility and other waste management designs; as well as to support environmental impact assessment studies.

Surface Water & Sediment Quality

Water quality monitoring was initiated in 1992, and as a result the project has established a long term monitoring record for the site at several stations. A review of all background water quality data in comparison to BC Approved and Working Water Quality Guidelines (BCAWQG) for the Protection of Aquatic Life was conducted by Golder (2007)^7.Results showed mean exceedances for turbidity, sulphate, dissolved aluminium, arsenic, cadmium, copper, lead, manganese and zinc at various sites. Sediments samples show pre-existing elevated levels of copper, arsenic and zinc in Bullion and Siwash Creeks. Cyanide is known to occur naturally at detectable analytical levels on the project site.

In mine water from the pit and underground workings, total cadmium was found to exceed the BCAWQG by a factor of 79, the largest factor. It is estimated that there would be more than enough natural drainage in the accepting creeks to dilute this mine drainage to acceptable levels, but Almaden has already commissioned the design of an earth fill dam to safely accept twice the estimated initial dewatering volume (Klohn Crippen Berger, 2010)⁵. This is to allow for the initial phase of operations, when stored water in the pit and underground workings will have to be removed.

In the SRK s opinion, there is scope to integrate all water storage with the TSF, and it appears that a significant part of the processing plant s annual water demand may be available from the mine drainage.

Water and sediment quality studies are ongoing at Elk Gold. In preparation for permitting monthly water quality sampling was started in the spring of 2010, and will be ongoing throughout 2011.

4.4.2.2 Biological Resources

There are no listed records in the Conservation Data Center (CDC)⁸of protected flora and fauna species in the project area. The only CDC element occurrence records in the region are at lower elevations more than 25 km from the project area and include one Yellow-listed plant, Regal s rush (Juncus regelii) a riparian vascular monocot; and a Red-listed bird, western screech owl (Megascops kennicottii macfarlancei). Occurrence of these species in the region does not suggest that the species occurs on the project site. A summary of terrestrial and aquatic biological resources is provided below.

Terrestrial – Vegetation and Wildlife

The Elk Gold Mine is located in the Okanagan Highlands ecosystem and falls within the very dry and cold Montane Spruce biogeoclimatic zone and within the transition between Interior Douglas Fir (IDF) and the Engelman Spruce (ESSF) at 1450 and 1650 ASL. The mine is located in an active forestry area and has been actively logged since 1993. An estimated 80% of the surrounding area has been clear cut. The remaining forest is primarily young secondary growth of lodgepole pine (Pinus contorta) and hybrid white spruce (Picea glauca x englemanni) and hemlock (Tsuga sp.) in protected and wetter areas. The understory is dominated by pine grass (Calamagrotis rubescens) and grouse berry (Vaccinium scoparium), arctic lupin (lupines lyallii) and turn flower (Phleum pretense).

According to plans from the local forestry licence holder⁹, the remaining mature secondary growth is scheduled to be cut in 2011 The pre-existing forestry impacts have impacts on water quality (e.g. erosion, sedimentation and nutrient enrichment), soils (e.g. soil loss and nutrient leaching); as well as, impacts on local wildlife noise disturbance and habitat loss.

⁷	Golder, 2007 “Report on Review and Data Gap Analysis of Environmental Baseline Studies, Elk (Siwash) Gold Mine” Prepared for Almaden Minerals.
⁸	www.srmwww.gov.bc.ca/cdc/gis/eo_data_fields.htm
9	Aspen Planers Ltd. – ortho map showing cut plan for winter 2011

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Wildlife resources have been described for the area by Wildstone Resources (1993)¹⁰and by Stantec Consulting (1999)^11.Moose and deer are reported to use the area as a summer range. Riparian area provide important habitat for foraging, security cover and calving for moose and deer. Black bears, deer and osprey have been observed in the project area by Almaden staff.

Grizzly bear may occur in the area however no sightings have been reported. Siwash Lake located 2 km SE from the project site is important habitat for local wildlife, given the dry climate.

Wardrop Engineering is planning for baseline vegetation and wildlife surveys during baseline monitoring during spring 2011.

Aquatic Studies

Aquatic studies are ongoing and include fish and fish habitat studies (Golder 2006)^12,benthic macroinvertebrate and periphyton surveys by Knight Piesold (2004)¹³, Golder (2008)¹⁴. Results showed healthy diverse benthic communities suitable for use in environmental monitoring. Additional information was sourced from the Fish Wizard database for BC and information from provincial and regional district offices of the Ministry of Environment.

No salmon species or bulltrout are known to live in the project area and productive fish habitat within the actual project area is limited. Siwash Lake located approximately 2 km SE of the project site is known to have a self-sustaining population of rainbow trout and relatively high recreational potential, some recreational cabins are located on the SE shore of the lake. Stream, lakes and marshes are habitat for amphibian species such as the spotted frog, western toad and long-toed salamander^15..

Almaden reports that environmental baseline studies of aquatic resources were ongoing through 2010 and will continue through 2011.

4.4.2.3 Socio-economic Resources

The region s economy is resource-based with forestry, agriculture, mining and tourism as the primary industries^16..

The project is located in a rural area within the traditional territories of three First Nation communities; the Upper Similkameen Indian Band (USIB), Upper Nicola Indian Band (UNIB) and the Westbank First Nation (WFN). A privately owned lodge is located approximately 2 km northeast of the project site. A few seasonal cabins occur on the southeast side of Siwash Lake. Other fishing and hunting lodges in the area includes Paradise Lake Resort located 9.5 km North of Elkhart road- 97C junction.

A Remote Access to Archaeological Data (RADD) application of the Provincial Heritage Register was conducted for the project area¹⁷. The nearest archaeological site is located at the outlet of Siwash Lake, approximately 1.4 km southeast of the mine site. The site record D1Rb-5 indicates surface scatter of stone flakes and a rock feature/dam (Rousseau, 1992)¹⁸. An updated Archaeological Overview Assessment (AOA) for the project this will be conducted with participation from relevant First Nations.

10 Wildstone Resources, 1993. Description of Wildlife Resources on and adjacent to the Siwash North Gold Property. Prepared for Fairfield Minerals Ltd. and Cordillearan Engineering Ltd., October 28, 1993.

11 Taylor, M.E. and PcKee, P. 1999. Wild Ruminant Study at Brenda Mine. Prepared by Stantec Consulting Ltd.

¹²Golder, 2006. Elk Gold Mine Baseline Fall Fisheries site Report. Technical Memorandum dated October 25, 2006. ¹³Knight Piesold Ltd. 2004. Environmental Baseline Studies. (Ref. No. VA102-147/02-1). December 2004

¹⁴Golder Associates Ltd. 2008. Water quality and Benthic Invertebrate Baseline Study, Elk Gold Mine, Siwash Lake, BC. Prepared for Almaden Minerals Ltd. dated July 4, 2008

¹⁴	Klohn Crippen Berger, 2010. Elk Gold Project – Water Storage Dam Design, Prepared for Almaden Minerals Ltd., dated March 10, 2010.
¹⁵	Golder, 2007. Report on Review and Data Gap Analysis of Environmental Baseline Studies, Elk (Siwash) Gold Mine” Prepared for Almaden Minerals.
¹⁶	www.westbankchamber.com and www.merritt-chamber.com
¹⁷	Golder, 2007 “Report on Review and Data Gap Analysis of Environmental Baseline Studies, Elk (Siwash) Gold Mine” Prepared for Almaden Minerals.
¹⁸	Rousseau, 1992. An archaeological Resource Overview Assessment for the Proposed Elk-Siwash Gold Deposit Project Between aspen Grove and Kelowna South Central BC Prepared by Antiquus Archaeological Consultants Ltd. 1992.

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Almaden reports having initiated early stage consultation activities with relevant First Nations and has begun a communication log and the collection of initial socio-economic baseline information in anticipation of permitting. Studies planned for 2011 include socio-economic data collection, public meetings, site visits and ongoing consultation.

4.4.3 Environmental Impact and Management

The Elk Gold Mine project is currently envisaged as an operation with a 10 to 15 year life operation. The first 8 to 10 years will see open pit mining with a strip ratio of about 25:1 and an ore production of between 180,000 and 360,000 tpa, 500-1000 tonnes per day. Underground mining at lower rates is likely to continue after cessation of open pit operations.

Project components are expected to include: An open pit, underground workings, waste dumps, tailings storage facility (TSF), water storage dam and impoundment, road upgrades, mine site facilities and a new power line.

Engineering designs are not even started, and environmental impact assessment studies and related management plans are not yet prepared. However, Almaden is well aware of the requirements of this process and has begun preparation of a Project Description which will detail all aspects of each mine structure and activity, potential environmental impact, and mitigation and management strategies.

4.4.4 Permitting, Community Consultation & First Nations

Permitting

Almaden currently holds a Provincial small mines permit (Permit M199 and MX-4-387) for the Siwash North Lease in 1994 remains current allowing for production of 10,000 tonnes per annum (27.4 tonnes per day). The Reviewable Projects Regulations¹⁹specify that new mineral mines²⁰are automatically reviewable if they have a planned production capacity of greater than 75,000 tonnes per year (205 tonnes per day) of mineral ore. The current development plan for production of 180,000 tonnes per year and an open pit the project will require an environmental assessment under the British Columbia Environmental Assessment Act (BCEAA). This is process is initiated once a Project Description (PD) for the mine is submitted and accepted by the BC Environmental Assessment Office (EAO).

Based on a preliminary review of the current mine concept the project is not expected to require a review under the Canadian Environmental Assessment Act (CEAA). The federal environmental assessment process is applied whenever a federal “trigger” is applied. In the case of a development project this means that

CEAA is triggered when the federal government provides a licence, permit or an approval that is listed in the Law List Regulations, such as: Fisheries Act subsection 35(2)²¹– i.e. causing alteration, disruption or destruction of fish habitat Navigable Waters Protection Act, paragraph 5(1)(a) ²²– i.e. no work shall be built or placed in, on, over, under, through or across any navigable water Explosives Act R.S.C. – E15²³- i.e. authorization for the manufacture of explosives on site Canadian Environmental Protection Act 1999, subsection 127(1)²⁴– i.e. overall approval of project concept if any federal triggers apply Species At Risk Act –Section 73(1) – i.e. alteration to critical habitat or impact to species listed under Schedule 1 of SARA on federal land Migratory Birds Convention Act – i.e. to protect or compensate for wetland habitat supporting migratory birds.

19	http://www.bclaws.ca/EPLibraries/bclaws_new/document/ID/freeside/13_370_2002
20	“Mineral mine” as defined by the Mineral Tenure Act http://www.bclaws.ca/EPLibraries/bclaws_new/document/ID/freeside/00_96292_01
21	http://laws.justice.gc.ca/eng/F-14/page-5.html#anchorbo-ga:s_34
22	http://laws.justice.gc.ca/en/N-22/
23	http://laws.justice.gc.ca/eng/E-17/page-4.html#anchorbo-ga:s_7
24	http://laws.justice.gc.ca/en/c-15.2/text.html

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Approvals of the Project under the BCEAA and CEA Agency are not the sole authorizations required for permitting. Additional permits, licences approvals, consents, authorizations and potential amendments are before receiving an Environmental Assessment Certificate (EAC). Almaden is aware of the requirements for these types of permit and plans to conduct this permitting concurrently with the EIA process.

Community Consultation & First Nations

Environmental assessment is an open process, which includes participation by all interested community stakeholders and First Nations. The British Columbia Environmental Assessment Act (BCEAA) includes provisions for public notification, access to information, consultation and issue resolution.

As per Section 5.3 of the Guide to the BC Environmental Assessment Process (EAO, March 2003), the Public Consultation Policy Regulation requires that the proponent; 1) conduct a public consultation program that is acceptable to the Environmental Assessment Office (EAO), and 2) report back to the EAO on the consultation activities and results. In addition the Elk Gold Project lies within the overlapping traditional territories of the Upper Nicola Indian Band (UNIB), the Upper Similkameen Indian Band (USIB) and the Westbank First Nation (WFN).

Almaden understands that effective consultation will be a key driver to project permitting success; and once the Project Description is accepted by the EAO, the first step in formal consultation will be to prepare the Terms of Reference (ToR) in participation with relevant community stakeholder and First Nations. The ToR defines the means and methods for the project s baseline and impact assessment studies. The second step will be to address stakeholder and First Nations interests and concerns on the Environmental Certificate application. In anticipation of the permitting and consultation process, Almaden has initiated a communications log and begun preparation of a Project Description for submission to the EAO. Pre-consultation activities has included referrals, letters, interviews and meetings regarding advanced exploration drilling activities and future plans for the project with relevant First Nations.

4.4.5 Initial Conceptual Reclamation Plans

Almaden has been engaged in progressive reclamation activities since acquiring the project. A review of past reclamation activities and exploration, development history is reported in Annual Reclamation Reports²⁵.

Almaden s plans for Elk Gold Mine include staged closure over the mine life including appropriate contouring and re-vegetation of waste dumps; flooding or capping of tailing facilities, and closure of roads and facilities; as well as post- closure monitoring plans for an appropriate period. It is intended that initial design of the TSF, water storage, mine drainage containment, mine waste dumps .and other facilities be integrated with closure plans for each facility and for the mine as a whole.

4.4.6 Summary

Almaden has been continuing to collect a wide variety of environmental data from the site since 1992 and is well prepared to integrate the design of all mining, processing, water supply, drainage, and infrastructure works with each other and with the environment.

Mine waste rock samples have been tested and found to be not acid generating. The dumps will be covered with topsoil and reseeded, as has already been trialled on the trial pit dumps. Environmental baseline of potential dump sites and a waste management plan will be drawn up, to include all the normal activities such as condemnation drilling, stripping topsoil, and reclamation procedures. Permitting of the final plan will be required.

A tailings dam site must be geologically examined to eliminate the chance of covering resources of value. Geotechnical foundation testing will also be required. A project water balance must be made, involving the development of an acceptable climatic profile. As a result, the dam can then be sized so as to avoid discharge to the environment. It is assumed that water will be recycled to the plant. The mine drainage will be discharged to the TSF, thus providing some process plant water, and avoiding direct discharge.

²⁵Almaden Minerals Ltd. 2010. Siwash Gold deposit Elk Gold Project Reclamation Report 2009. Permit M-199 and Amended Mines Act Permit MX-4-387 Almaden Minerals Ltd.

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Initially, the mine dewatering must precede mining, and the TSF may not be ready to accept the water at that time. Therefore, the water will need to be held within a purpose-designed drainage containment which has already been designed (Klohn Crippen, 2010). Release from this may never be necessary. If it is, it will be controlled so as not to exceed any quality limits in the accepting stream.

The plant tailing has been tested and found to be not acid- generating. The source rock for the dam wall must be tested for suitability.

Work on the tailings and surface and mine water management will be a part of the task of locating a make-up water source and obtaining permission to abstract the quantity required for mining and processing.

Permission to build a power line to the site would be desirable before proceeding with the project. Permission to upgrade and re- route the access road from the adjacent highway will be required. The site infrastructure must be sited and permitted.

The First Nation Groups will be consulted on an ongoing basis.

The submission of a Project Description, partly based on this document, to the BC Environmental Office will be the initial step to initiate the formal review process. The British Columbia Environmental Assessment Office will then issue a procedural order which will identify specific requirements.

4.4.7 Current Environmental and Social Liabilities

Final remediation of the current disturbance on the property was estimated to cost $150,000 in 2007 (Golder 2008b). As subsequent exploration and mining activities have taken place on the property, disturbed sites have been remediated.

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5.	Accessibility, Climate, Local Resources, Infrastructure and Physiography
5.1	Access to Property

Access to the claims is excellent, by following the Okanagan Connector (HWY 97C) east from Merritt for 50 km to the Elkhart Road interchange. If approaching from the east, the same highway would be followed 50 km west from Peachland. This highway passes through the northernmost claims in the property. From the Elkhart Road interchange, gravel roads and trails provide access to most parts of the property.

5.2 Physiography and Climate

The Elk property is located within the Trepanege Plateau highlands on heavily forested hilly terrane. Elevations range from 1,300 to 1,750 m above sea level. The area is blanketed by a layer of glacial till of varying thicknesses, and as a result, outcrop is scarce.

Forest cover consists mainly of lodge pole pine with some balsam, sub-alpine fir and spruce. Alders are found along streams and in marshes.

The property is within the Okanagan Highlands and has relatively short warm summers and cold winters. Average daily temperatures for May through September are 12º to 18ºC, with extremes ranging from 5º to 39ºC. Average daily temperatures during the winter months of November to February are 0º to -5ºC, with extremes ranging from 14º to -9ºC.

The average yearly rainfall is 235 mm, with most of it falling from May to October. Monthly rainfall averages are between 20 and 25 mm with June being the wettest month with an average rainfall of 34 mm. The yearly snowfall average is 83 cm with most of it falling between December and February. The source of the climatic data noted here is Environment Canada from data collected over the last 30 years.

The main industries within the area are cattle ranching and logging. Recreational fishing is available on small lakes within the area. The dense forest cover also supports hunting of deer, moose and game birds.

5.3	Power and Water Resources
A	single-phase power line ends at Elkhart Lodge 2 km from the mine site. Three-phase power is located

approximately 20 km east and west of the project. For the Base Case Scenario developed in Chapter 20, it is proposed that power will be generated by Almaden s existing diesel generating sets, augmented as appropriate. The $1200 case assumes that three phase power will be brought to the site by a new power line.

The availability of water for ore processing and mining use is not certain. Further study is required to confirm that adequate supplies exist.

5.4 Local Resources and Infrastructure

Kelowna is located 62 km east of the property and has a large labour force, airport, supplies and equipment. Merritt and Peachland are both about 40 km from the property and have adequate facilities for a commuting workforce.

5.4.1 Buildings and Ancillary Facilities

Almaden recently commissioned an accurate topographic survey of the whole area.

There are several small buildings at site, comprising drill core storage, drill logging building, and storage of rock and pulp samples. A weather monitoring station has been established to collect baseline climate information. The existing trial pit and dumps can also be seen.

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5.4.2	Camp Site

There is a 20-man camp established at site, consisting of a mess and kitchen, generator building and water treatment plant, along with tent frame dwellings for sleeping quarters.

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6.	History of Property

Very detailed descriptions of historical exploration on the Elk property are included in reports by King (2001), Jakubowski (various years) and Giroux (2007). This section is a summary from these reports.

6.1 Exploration History

The first reported exploration is from the early 1900s when a few short adits were driven on narrow sulphide veins. One of the adits was driven on a quartz vein hosting lead, zinc, silver and gold mineralization in volcanic rocks. A minor amount of placer gold was found near the old adit in the Siwash Creek area.

6.1.1	Copper and Molybdenum Exploration
A	number of companies staked and explored the area for copper and molybdenum mineralization in the

1960s and 1970s during development of the copper- molybdenum Brenda Deposit located 22 km east of the Elk property.

No significant discoveries were made from the historical exploration listed below:

1955-1995, Don Agur, of Summerland, BC, prospected the north and west parts of the present Elk property and to the south along Siwash Creek.

1972, Phelps Dodge Corporation of Canada, Ltd conducted mapping and soil geochemistry in search of copper mineralization in the present claim tenure #516759 and #516757.

Utah Mines Ltd conducted mapping, geochemistry, IP geophysics and trenching in search of copper mineralization on part of what is now tenure #516759.

1979 to 1981, Brenda Mines Ltd conducted copper exploration including mapping, soil geochemistry, geophysics, trenching and diamond drilling. Work was done in the southern part of the Elk property, currently covered by tenure # 516755, 516757 and 516759.

1980, Cominco Ltd conducted geological mapping and soil geochemistry in search of molybdenum mineralization. The work was completed in current tenure #515727, 516731, 516733 and 516740.

6.1.2 Gold Exploration

From 1986 to 1989, Cordilleran Engineering Ltd staked claims on what is now the Elk Property on behalf of Fairfield Minerals Ltd. In 1988-1990, Placer Dome Inc. entered an option agreement with Fairfield whereby Placer paid exploration costs. In 1993, Placer Dome Inc. transferred their interest to Placer Dome Canada Ltd and in 1994, Placer relinquished their interest in the property (King, 2001). Fairfield Minerals amalgamated with Almaden Resources Corporation in February 2002 and the claims were transferred to the amalgamated company Almaden Minerals Ltd (Almaden).

The following is a summary of the exploration work completed during the late 1980s and into the early 2000s:

1987: Soil geochemistry, geophysical surveys and trenching identified gold-bearing quartz vein structures extending for hundreds of meters.
1988: Trenching in B and WD Vein Complex (historically called the Siwash North Area) and Elusive Creek area.
1989: Trenching and stripping of overburden exposed quartz veins which extended the gold-bearing vein system 750 m along strike and 12 HQ-sized diamond drill holes (778 m) were completed to confirm down-dip continuity in the B and WD Vein Complex. Geophysical surveys and soil sampling were done in other areas of the property.
1990: Additional stripping of overburden was completed at the B and WD Vein Complex, detailed (5-m spaced) sampling of the B vein was completed, and 58 HQ-sized diamond drill holes (5,158 m) on 50-m centres along the vein were drilled. Several gold shoots and locally stacked, parallel veins were identified. In the Siwash Lake area trenching, overburden stripping and drilling of four HQ-sized diamond drill holes (259 m) were completed. Soil sampling was completed in other areas of the project and geophysical surveys were done on the Agur Option.

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1991: 37 HQ-sized diamond drill holes (6,587 m) were drilled in the B and WD Vein Complex, trenching was completed in the End Zone (200 m southwest of Siwash Lake), and aerial photos were taken.
1992: A 2,040 t bulk sample grading 137 g/t was excavated from the „Mother Shoot (Main Vein in B Vein Complex), and 79 reverse circulation drill holes (2,683 m) were drilled to test the area of the open pit bulk sample.
1993: Open pit bulk sampling continued, 11 reverse circulation drill holes (942 m) were drilled in the pit area and a portal was collared and a decline (480 m) driven for underground exploration.
1994: Underground exploration continued, 84 NQ-sized diamond drill holes (5,011 m) were drilled underground, the main decline was extended 330 m with a branch of it extended an additional 185 m, testing of underground mining methods (long-hole stoping and shrinkage stoping) were pursued, and following receipt of a mining permit, the open pit was expanded.
1995: 217 NQ-sized diamond drill holes (7,612 m) were drilled underground from the decline and a total of 98 HQ-sized diamond drill holes (6,289 m) were drilled from surface in the project area, 70 (4,645 m) of the 98 were drilled in the Siwash North area.
1996: 88 NQ-sized diamond drill holes were drilled (6,946 m), five of these holes (1,120 m) were drilled in Siwash North Deep B and 38 of the holes (1,997 m) were drilled to the east of the open pit area. 1997: Mining reclamation and site clean-up were the main activities at site.
1997-1999: Limited prospecting, sampling and environmental monitoring.
2000: 12 NQ-sized diamond drill holes (1,414 m) were drilled to test the WD, B and Gold Creek veins. 2001: Six trenches (202 m) were completed in the Siwash East area.
Exploration and drilling between 2001and 2007 are presented in Section 10 of this report.
Open Pit Bulk Sample and Underground Exploration
Explorative mining completed on the property included both open pit bulk sampling and underground exploration between 1992 and 1994. Open pit and underground exploration was only completed on the B Vein.
The mineralized material extracted from the pit and underground exploration in 1992 was shipped to the Horne smelter in Rouyn Noranda, Quebec; and mineralized material extracted in 1993 and 1994 was shipped to the ASARCO smelter in Helena, Montana for processing as high-grade silica flux.

Table 6-1: Historical open pit and underground mine production

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6.2	Historical Resource and Reserve Estimates

A	number of historical resource and reserve estimates have been made. The amount of data has increased

significantly over the years and the methodologies and assumptions used to generate these estimates have varied. The estimates reported in 2004 and 2008 were calculated by Mr Gary Giroux, P.Eng, and NI 43-101 reports were issued for these. The earlier estimates are not considered NI 43-101 compliant.

6.2.1 Historical Mineral Reserve Estimates Completed in 1995

In February 1995, Rowe (1995) reported an “indicated reserve” of 156,000 t at 36.5 g/t gold. Minimum grade and thickness criteria were applied to the drill hole data for the polygonal estimate. The estimate was based on 100 diamond drill holes on a spacing of 50 x 50 m. This estimate is not considered to be NI 43-101 compliant.

In November 1995, Roscoe Postle Associates (Roscoe, 1995) estimated “Probable Reserves” of 29,400 t grading 26.0 g/t gold and “Possible Reserves” of 66,400 t grading 31.5 g/t gold. In the Roscoe report, for “Probable Reserves”, drill spacing is 10 x 10 m and has 15% dilution applied, “Possible Reserves” (a term no longer used) is based on 25 to 50 m spaced data and has no dilution applied. This estimate was focused on the „Mother Shoot of the B Vein and was prepared using grade times true thickness projected to a longitudinal section. Stoping blocks were outlined using a gold grade cut-off of 14 g/t. The Roscoe report states that the tonnages in their estimate are considerably smaller than the Rowe estimate because “detailed

1995 drilling demonstrated that mineralized shoots within the vein system are smaller than previously assumed.”

This estimate is considered a historical estimate and is not considered to be NI 43-101 compliant.

6.2.2 Mineral Resource Estimates Completed by Giroux Consultants

In May 2004, Giroux Consultants estimated resources (Table 6-2) using a two-dimensional (2-D) approach for the B and WD Veins and a 3-D block model for the B Vein in the open pit area. In reference to the 3-D model, Giroux states, “the tonnage and volume contained in this model would include significant parts of the B Flat, B Steep and B East vein resource estimated in the 2-D approach...” It is the opinion of LGGC that this indicates possible serious issues with the 2004 estimate as these areas overlap in the two estimation methods used.

Giroux Consultants (2007) completed an updated resource estimate (Table 6-3) for the project using drilling completed through 2006. A total of 853 drill holes were used in the estimate. Almaden prepared the solids models representing 25 mineralized structures which were used for the basis of the estimate. Composites were created from capped (top cut) assays and gold grades were estimated using ordinary kriging. Resource categories were assigned based on variogram ranges. A review of this model by Almaden and LGGC showed vein intersections missing from the composite inventory due to the vein intersections not being tagged or tied to the drill holes in 3-dimensions during the grade compositing process. This situation lead to many intersections of vein material not being included in the composites file therefore many block values were informed using waste intervals and not vein intervals.

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Table 6-3: Resource Estimate, Giroux Consultants, 2007

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7.	Geological Setting
7.1	Geological setting

The information for this section is sourced from reports by King (2001), Jakubowski (2000-2006) and Giroux (2007).

7.2 Regional Geology

The Elk Property lies within the Intermontane tectonic belt of south-central British Columbia. Upper Triassic magmatic arc sequence volcanic rocks and sediments of the Nicola Group cover the western third of the project area. The eastern two-thirds is covered by Middle Jurassic age intrusive rocks of the Osprey Lake Batholith, see Figure 4-1. The contact between the Nicola andesite and basalt and the Osprey Lake granite and granodiorite trends NNE across the western part of the claim block. Early Tertiary quartz-feldspar-porphyry dykes and stocks from the Otter Intrusive occur within the property. The youngest units mapped are andesite dykes which cross cut all other lithologies.

Breccias of rounded volcanic, diorite and granite fragments in a granite matrix cross cut the Nicola, Osprey Lake and Otter formations. The breccias strike north-east and may identify major fault structures.

7.3 Elk Project Geology

The geological events leading to the Elk Project deposit occurred in the following order:

1.	Deposition of Nicola volcanic
2.	Emplacement of Osprey Lake Batholith
3.	Emplacement of Otter intrusions
4.	Fracturing possibly from Osprey Lake and/or Otter intrusion of andesite dykes
5.	Precipitation of quartz veins with pyrite, base metal sulphides
6.	Late stage gold mineralization with associated hydrothermal alteration
7.	Erosion to present topography (Jakubowski, 2007).

The western claims area is underlain by steeply west-dipping andesitic to basaltic flows, agglomerates, tuffs and minor siltstone and limestone units of the Upper Triassic Nicola Group. The eastern half of the property is underlain by Jurassic granitic rocks of the Osprey Lake Batholith. The contact between these two assemblages trends approximately north-northeast.

Upper Cretaceous to Tertiary feldspar porphyry and quartz-feldspar porphyry stocks and dykes of the Otter Intrusions cut both of the above. Breccias containing rounded volcanic, dioritic and granitic fragments in a granitic matrix crosscut Nicola Group rocks, Osprey Lake and Otter Intrusions. Andesite dykes are the youngest units mapped, post-dating all of the above. Mineralization appears to be spatially associated with these andesite dykes, which are locally cut by quartz veins.

Nicola Group lithologies mapped on the Elk property dip 60º to the west, forming the east limb of a north-south trending syncline. The rocks are dark greyish green massive, basalts and andesites (locally porphyritic texture with pyroxene and/or amphibole phenocrysts and 0.5 mm lamina of sand=sized black grains), pale grey-green siliceous laminated tuff; and brownish green to pale green agglomerates containing 5-50 cm sized fragments.

The Osprey Lake Batholith is composed of pinkish grey, medium- to coarse-grained equigranular quartz monzonite to granodiorite. Pink, sugary textured aplite dykes often cut the quartz monzonite and are thought to be a late phase of the intrusive event. Quartz diorite stocks are less prevalent and may be related to the batholith.

Quartz monzonite and hornblende-biotite-quartz monzonite are also present locally and have been mapped as dykes.

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The granodiorite border phase of the Osprey Lake Batholith contains most of the previous mine production (Lewis, 2000).

The Otter Intrusions are quartz feldspar porphyry, feldspar porphyry and quartz-biotite-feldspar dykes and stocks. Quartz feldspar porphyry is strongly clay altered and contains feldspar phenocrysts up to five cm but averaging about five mm. The altered groundmass is beige in colour and can be strongly friable. Feldspar porphyry rocks range from medium grey to red and contain feldspar phenocrysts 2 to 5 mm in size that vary in quantity from 3 to 40%. Petrographic examination of the red, medium packed feldspar porphyry indicated a syenitic composition. Quartz-biotite-feldspar porphyry is greyish beige and is typified by small biotite grains with equal quantities of fine quartz and feldspar phenocrysts.

The younger andesite dykes are fine-grained, dark greyish green and 30 cm to 5 m thick. Mineralization appears to be spatially associated with these (possibly Tertiary) andesite dykes which are locally cut by quartz veins.

Nicola volcanics are locally silicified, carbonatised or altered to epidote the Osprey Lake batholith has weak to strong propylitic, argillic, phyllic and silicic alteration, the Otter quartz-feldspar porphyry is extensively altered to clay and the andesite dykes are altered to muscovite.

According to logging reports, alteration zoning extends symmetrically grading outward from the quartz veins through advanced-argillic, phyllic to potassium-stable-phyllic, argillic and propylitic assemblages (Fairfield, 2000). Strong to very strong phyllic alteration encloses the vein with some areas close to the vein less intensely altered according to underground mapping (Conroy, 1994).

The B and WD Vein Complex was emplaced within a fault/fracture zone that strikes east-northeast and dips moderately to steeply southward (Lewis, 2000).

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8.	Deposit Types

The Elk project is a gold-bearing mesothermal quartz + sulphide vein deposit. This depositional environment is characterised by deep circulation and evolution of meteoric water in structures associated with major, strike-slip fault zones (Nesbitt, et al., 1986). Fluid inclusion studies (Geiger, 2000) indicate a lithostatic pressure of 2.5 kilo bars and a minimum formation temperature of 250ºC, corresponding to an approximate 7 km depth of formation.

Gold mineralization occurs within structurally controlled pyritic quartz veins. Vein widths range from a few centimeters to several meters wide.

Other deposits of this type include the Mother Lode District, California and the Bralorne-Pioneer District, BC.

In these deposits, high gold grades tend to be continuous over large vertical ranges.

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9.	Mineralization

Gold mineralization has been identified in multiple locations on the property, hosted primarily by quartz veins and stringers in altered granitic and, less frequently, volcanic rocks. Cross-cutting relationships indicate that the veins are Tertiary in age possibly related to the Tertiary Otter Intrusions.

9.1 B and WD Vein Complex (Siwash North Area)

Mineralization occurs as fine grained native gold (typically less than 50 microns) in quartz, in quartz-pyrite breccias and in fractures within veins (King, 2001). Gangue minerals include quartz and altered wall rock clasts, with minor amounts of ankerite, calcite, barite and fluorite. From surface to a depth of several meters, oxidised groundwater has leached most of the sulphides with some pyrite and chalcopyrite remaining (Jakubowski, 2007).

Veins occur along east-west to east-northeast-striking, south dipping fault/fracture zones (Lewis, 2000). Cross-cutting relationships suggest the veins are Tertiary in age and possibly related to the Tertiary Otter Intrusions (Giroux, 2007). Mineralization occurs in single high-grade veins and also as a tabular zone containing several closely-spaced, sub-parallel mineralized veins (Lewis, 2000).

Three principal styles of veining related to the primary lithologies were identified by Lewis (2000) from observations in the open pit. These include the western en-echelon segment which strikes 065º-070º, has an overall vein envelope width up to 10 m and is hosted within the granodiorite border phase of the intrusive. The central segment is a single, discrete structure with minor secondary veining and is hosted mainly in the granodiorite but extends into the quartz monzonite. Vein orientation is variable with step-over zones (abrupt vein shifts of 1.5 m) containing high-density secondary veins. These step-over zones were highly favourable during mining. In both the western and central segments, the veins dip shallowly near surface then abruptly steepen (50º to 60º) and change character at about 1610 m elevation. The eastern segment consists of several sub parallel east-northeast-striking, moderately south dipping splays. The transition from a single, thick vein to a zone of sub parallel veins occurs just east of the granodiorite/quartz-monzonite contact (Lewis, 2000).

Small scale open pit and underground mining and extensive exploration and drilling have been completed in the B and WD Vein Complex. Mineral resources were estimated by LGGC for this area and are presented in Section 17 of this report.

9.2 Other Exploration Areas at the Elk Project

Ten other prospective locations, Figure 9-1, and Figure 4-2: have been identified (King, 2001 and annual assessment reports) on the property including Bullion Creek, Siwash East, Lake Zone, End Zone, Gold Creek West and East, Great Wall Zone, Discovery (North) Showing, South Showing, Elusive Zone and the Agur Option area. The exploration work completed in the areas beyond the B and WD Vein Complex suggests the mineralization system trends north to northeast. Soil sampling, trenching and drilling have traced the rock types and alteration favourable for mineralization for 2 km along the north-south strike direction.

These areas are discussed in detail in Section10.2 of this report.

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Figure 9-1: Location map of the Exploration Targets identified on the Elk Project to date

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10.	Exploration

Exploration activity on the project from 1986 to 1994 was by Cordilleran Engineering Ltd. From 1995 to 2002 the work was done by Fairfield Minerals and from 2002 to present the work was done by Almaden. Mr Wojtek Jakubowski (P.Geo and Qualified Person as defined by NI 43-101) was responsible for exploration and drilling from 1987 to 2007. A full description of exploration work is available in yearly drilling reports authored by Mr Jakubowski as well as in King (2001) and Giroux (2007). These reports are the source information for data presented in this section.

A summary of all non-drilling exploration since 1986 is shown in Table 10-1.

Table 10-1: Summary of exploration work completed from 1986 to 2007

Type of work	Year	Extent of work
Geochemical Soil Surveys	1986-1996	17,400 samples collected
Litho Geochemical Samples	1987-1992	5,728 samples collected
Metallurgical Samples	1990 and 2007	Samples collected and sent for metallurgical test work
Geological Mapping	1997-98, 1990, 1992	4,110 hectares mapped
		7,905 m
Trenches	1986-1996, 2001, 2004	202 m
		20 m
		14.9 km of road built
Road Building	1987-88, 1990-91, 2004, 2005	500 m
		350 m
Legal Surveys	1990-92	1,824 km
	1987, 1989, 1990,	Mag. 105 km, IP 4.5 km
Geophysical Surveys	2004,2009	Mag. 15.8 km
	1992	UTEM 1.8 km
Underground Development	1993-1994	Portal and decline driven 995 m
Open Pit Bulk Sampling & Limited Mining	1992-1994	480,000 m3 waste 16,230 tons mined 47,600 oz Au produced
Reclamation	1989	15 hectares
From King (2001), Giroux (2007) and Jakubowski (2002-2007)

Exploration on the property has been done in a logical sequence from geophysical surveys and soil and trench sampling followed up by widely spaced diamond drilling to drilling on a regularly spaced grid for delimiting resources. Trial open pit and test-scale underground mining in 1992-1993 by Fairfield Mining provided very detailed information about the character of the veins.

The primary exploration activity conducted by Almaden since 2002 has been diamond drilling from surface, with most of the drilling in the B and WD veins.

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10.1	B and WD Vein Complex

Siwash North and WD veins have been the primary focus in the project area in terms of the amount of drilling. LGGC has prepared an updated Mineral Resource for these veins. A list of exploration activities completed by Almaden in the B and WD Vein Complex including drilling is included below: 2002 – 26 NQ-sized diamond drill holes (4,996 m) were completed in Waste Dump (WD) (7 holes), Deep B (11), Gold Creek West (GCW) (4) and Bullion Creek (4) zones (Jakubowski, 2003). Refer to Figure 4-2: for vein locations.

2003 – 30 NQ-sized diamond drill holes (6,570 m) were completed which tested strike and down-dip extensions of the WD zone; and claim posts in the south destroyed by logging were located with GPS (Jakubowski, 2004).

2004 – 500 m of roads were built, 40 m of road-cut was mapped and sampled as a trench, 15.8 km of ground magnetometer geophysical survey was carried out to the east of the mine site and existing drill grid, additional claim posts were surveyed by differential GPS and 44 NQ-sized diamond drill holes (10,265m) were drilled in the WD, B and BC areas (Jakubowski, 2005).

2005 – 350m of road building and 36 NQ-sized diamond drill holes (8,395 m) in the WD, B and Siwash Lake zones. The drill pattern in WD Vein was brought to 25 m by 50 m and the vein system was tested down-dip and to the west (Jakubowski, 2006).

2006 – 58 NQ-sized diamond drill holes (8,873 m) were drilled to test the WD (27 holes), B (17 holes) and Siwash East (4 holes) zones. Trenching and soil sampling was done on new claims to the northwest side of the property (Jakubowski, March 2007).

2007 – 9 HQ-sized diamond drill holes (2,470 m) were drilled for samples for metallurgical testing (Jakubowski, October 2007). A, B, C and the deeper WD zones were targeted with this drilling. Metallurgical testwork was completed by G&T Metallurgical Services (G&T 2008).

2010 – 87 diamond drill holes (12,749 m) were drilled as part of the infill drilling program in the area of the mineral resources, and for exploration in the South Showing area (4 holes).

Figure 10-1: Elk Property vein locations

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10.2	Other Showings on the Property
10.2.1	Bullion Creek Area

The Bullion Creek structure is 700 m north of the existing open pit. Six holes have been drilled in the area testing hydrothermal alteration anomalies identified from soil sample results and an east-west trending topographic low. The anomalies are illustrated in Figure 10-2. The holes intersected fine-grained granodiorite with strong phyllic alteration, which is typical of the intrusive rocks in contact with the Nicola volcanic in other parts of the property.


Figure 10-2:	Anomalous gold mineralization location map for Bullion Creek area
Note: DDH locations on Soil Anomaly Base Map

Two holes (165 m) NQ sized core holes were drilled in 2002 and four holes (395 m) were drilled in 2004 to test a geochemical anomaly identified from soil sampling with two holes on three sections spaced 50 m apart. Strong phyllic altered, fine grained granodiorite, typical of the intrusive rock near the contact with the Nicola volcanics and narrow quartz veins were intersected in the drill holes. The best intercepts from the drilling are included in Table 10-2.

Table 10-2: Significant Intersections in Bullion Creek Area

	Au	Downhole width
Drill hole ID			Includes
	(g/t)	(m)
SND-02-314	Nothing Significant
SND-02-315	8.43	0.65
SND-04-376	1.56	1.96
SND-04-377	2.73	2.47	13.66 g/t over 0.32 m and 5.50 g/t over 0.43 m
SND-04-378	Nothing Significant
SND-04-379	1.24	1.68

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10.2.2 Siwash East Area

Siwash East is located 2.5 km northeast of the B and WD Vein area. Quartz veins occur adjacent to an andesite dyke that trends 65º and dips 65º to the south. The veins are discontinuous and are found on the north side of the dyke, Figure 10-3.


Figure 10-3:	Anomalous gold mineralization location map for Siwash East area
Note: DDH locations on Soil Anomaly Base Map

In October 2001, six trenches (totalling 202 m total length with four trenches on 100 m step outs) intersected the quartz veins adjacent to the andesite dyke. Anomalous and high grade gold were found in narrow zones in the trench samples.

In 2004, a ground magnetometer survey (15.8 line km) was completed and a road cut that exposed 40 m of bedrock was mapped and sampled.

In 2006, four NQ-sized core drill holes (505 m) were drilled on two fences spaced 50 m apart to test the continuity of the mineralized quartz veins exposed by trenching. Discontinuous, low grade mineralized quartz veins were intersected in all four holes.


Table 10-3:	Significant drill hole intersections in Siwash East area
	Au	Downhole width
Drill hole ID		Includes
	(g/t)	(m)
SED-06-446	1.43	0.40
SED-06-447	1.05	1.50
SED-06-448	1.27	0.30
SED-06-449	1.03	0.40

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10.2.3 Gold Creek East and West Areas

Gold Creek West is 450 m southwest of the B and WD Vein area. Mineralization has been identified in a secondary fracture set trending 110º to 160ºand dipping 45º to 60ºsouth west, Figure 10-4.


Figure 10-4:	Anomalous gold mineralization location map for Gold Creek West area
Note: DDH locations on Soil Anomaly Base Map

In 1996, anomalous soil geochemistry in the Gold Creek West area was followed up with 7 NQ-sized core drill holes (557 m).

In 2000, five NQ-sized core drill holes (343 m) were drilled to test the vein continuity at 50 m spacing. The vein was intersected at the projected location.

In 2002, four NQ-sized holes were drilled (333 m) in two 50-m step-outs to the west. The vein was intersected at the projected locations but recovery of the core was reported as poor in two of the drill holes.


Table 10-4:	Significant drill hole Intersections in Gold Creek West area
	Au	Downhole width
Drill hole ID			Includes
	(g/t)	(m)
SND-96-291	1.77	5.57	4.49 g/t 1.79 m
SND-96-292	Nothing significant
SND-96-293	7.03	0.94	19.68 g/t 0.30 m
SND-96-294	Nothing significant
SND-96-295	2.40	0.63
SND-96-296	Nothing significant
SND-96-297	Nothing significant
SND-00-304	27.2	0.36
SND-00-305	5.25	0.37
SND-00-306	Nothing significant
SND-00-307	3.22	0.32
SND-00-308	0.690	0.40
SND-02-321	Nothing significant
SND-02-322	1.33	0.90
SND-02-323	1.31	0.30
SND-02-324	1.78	0.62

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Gold Creek East zone is located 600 m southeast of the B and WD Vein area, Figure 10-5.

Figure 10-5: Drill hole location map for Gold Creek East area

Note: DDH locations on Soil Anomaly Base Map

In 1996, four NQ-sized drill holes (400 m) were drilled in the Gold Creek East area to follow up on anomalous soil sample results and a VLF conductor. No significant gold results were intersected.


Table 10-5:	Significant drill hole intersections in Gold Creek East area
	Au	Downhole width
Drill hole ID		Includes
	(g/t)	(m)
SND-96-284	Nothing significant
SND-96-285	Nothing significant
SND-96-286	Nothing significant

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10.2.4 Lake Zone Area

In the Lake Zone Area, located 800 m south of the B and WD Vein Area, mineralization appears to occur mainly in quartz stringers and veins up to 35 cm thick, hosted by strongly argillic- to phyllic-altered granitic rocks in contact with an andesite dyke. The zone trends east-west and dips about 50º to the south. At surface and in drill core, the gold mineralization is associated with pyrite, chalcopyrite, and locally high concentrations of galena and sphalerite. Tetrahedrite and possibly maldonite are also locally present, Figure 10-6.

Figure 10-6: Anomalous Gold mineralization location map for Lake Zone East area

Note: DDH locations on Soil Anomaly Base Map

In 1989, trenches were excavated in the Lake Zone area where quartz veins and stringers were exposed.

In 1990, diamond drilling in the Lake Zone Area consisted of 259 m of HQ core in four drill holes (SLD90-56 to 59). Six trenches and one stripped area totalling 607 linear meters of bedrock exposure were excavated. Soil sampling on the northern Elk claims was concentrated in the Lake Zone Area where 250 infill samples were collected.

In 1995, drilling at the Lake Zone was done on three north-south drill fences spaced 50 and 100 m apart. Seven holes totalling 477 m were drilled to the north to intersect quartz vein mineralization encountered in 1989 trenching and 1990 drilling.

In 2005, five diamond drill holes (509 m) were completed in the Lake Zone Area on 50-m centres to test for grade and continuity to the west and at depth along the south dipping structure. Mineralization was intersected in all holes adjacent to a south dipping andesite dyke. The dyke dips about 45 degrees on the east part of the drill grid and steepens to 70 degrees to the west. High gold values in narrow intersections were returned from sampling and the structure remains an interesting exploration target.

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Table 10-6: Significant drill hole intersections in Lake Zone area

	Au	Downhole width
Drill hole ID			Includes
	(g/t)	(m)
SLD-90-056	2.95	2.60
SLD-90-057	1.13	5.75
SLD-90-058	Nothing significant
SLD-90-059	4.46	0.50
SLD-95-148	2.26	1.18
	4.06	0.94
SLD-95-149	1.69	4.56
	2.46	4.35
SLD-95-150	9.99	2.57	55.44 g/t/0.55 m and 18.45 g/t/0.36 m
SLD-95-151	Nothing significant
SLD-95-152	Nothing significant
SLD-95-199	Nothing significant
SLD-95-200	7.70	1.44	15.91 g/t/0.67 m
SLD-05-436	1.11	4.48
	2.87	2.60
SLD-05-437	3.05	1.60	13.73 g/t/0.30 m
SLD-05-438	6.87	1.2	18.43 g/t/0.30 m and 8.97 g/t/0.30 m
SLD-05-439	9.01	2.1	14.26 g/t/0.50 m and 21.26 g/t/0.50 m
SLD-05-440	3.12	0.30

10.2.5 Great Wall Zone Area

The Great Wall Zone is 500 m south of the Lake Zone area where a quartz vein boulder float was found during road construction, Figure 10-7.

In 1995, a 5 m by 25 m trench was dug in the Great Wall area to locate the source of pyritic quartz vein float. The trench exposed moderately sericitic and propylitic altered quartz monzonite cut by a 10 to 20 cm quartz vein trending 55° and dipping 65° to the south. Two 0.5 m square panel samples were taken across the vein returning values of 0.51 g/t and 0.99 g/t Au.

Two drill holes, totalling 102 m, were drilled in 1995 on a north-south drill fence below the trench and in an area of coincident geochemical and geophysical anomalies. The holes intersected the zone at 22 m and 24 m down dip. Narrow stringers with gold mineralization were intersected.

Table 10-7: Significant drill hole intersections in the Great Wall area

	Au	Downhold width
Drill Hole ID	(g/t)	(m)	Includes
GWD-95-155	0.93	0.33
GWD-95-156	3.81	0.29

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10.2.6 End Zone Area

Mineralization near the southwest end of Siwash Lake, known as the End Zone, Figure 10-7, is similar to that in the Lake Zone, but trends approximately northeast dipping about 55º to the south. The quartz veins are 1 to 20 cm in thickness and are hosted in strongly to moderately altered quartz monzonite. The dominant sulphide minerals noted in the quartz veins include pyrite, galena, sphalerite, chalcopyrite, tetrahedrite and arsenopyrite. Silver to gold ratios were also elevated and are similar to the Lake Zone.


Figure 10-7:	Detailed results of exploration completed at the Great Wall Zone, End Zone and the Discovery Showing (North Showing) areas

In 1991, one trench was dug in the End Zone to further expose a quartz vein discovered by trenching in 1990. The vein is continuous across the entire length of the 45 m trench.

In 1995, four holes totalling 187 m were drilled on three north-south drill fences spaced 60 and 90 m apart. These holes tested mineralization encountered in the 1990 and 1991 trenching programs, and geochemical and geophysical anomalies (Table 10-8).

Table 10-8: Significant drill hole intersections in the End Zone area

	Au	Downhold width
Drill Hole ID	(g/t)	(m)	Includes
EZD-95-153	10.05	0.56
EZD-95-154	10.56	0.30
EZD-95-157	Nothing significant
EZD-95-158	24.41	0.50

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10.2.7 Discovery Showing (North Showing Area)

The Discovery Showing area, historically referred to as the North Showing, is the original gold discovery on the property that led to the staking of the mineral claims. It is located in a clear-cut area 2.0 km south of B and WD Vein area and hosts an irregular northeast-trending quartz vein ranging from 10 to 80 cm thick, with locally up to 20% pyrite and local minor chalcopyrite and galena. The vein is spatially associated with an andesite dyke that cuts granitic host rock and an irregular body of clay-altered feldspar porphyry. The vein structure has been traced by trenching for a length of 200 m, see Figure 10-7 and Table 10-9.

During 1987, nine trenches, totalling 1528 m were excavated to test soil geochemical targets, exposed quartz veins and altered breccias hosted in granite. IP, magnetometer and VLF-EM geophysical surveys were carried out over the trenched areas.

In 1995, six holes totalling 397 m, on three north-south drill fences spaced 50 m apart were drilled to test mineralization uncovered during trenching.

Table 10-9: Significant drill hole intersections in the Discovery Showing area

Drill hole ID	Au	Down hole width	Includes
	(g/t)	(m)
DSD-95-159	13.60	0.68	22.01 g/t/0.37 m
	4.80	0.33
DSD-95-160	Nothing significant
DSD-95-161	4.46	0.30
DSD-95-162	Nothing significant
DSD-95-163	Nothing significant
DSD-95-201	1.03	0.40

10.2.8 South Showing Area

In the South Showing area, located 2.7 km south of the B and WD Vein area, mineralization occurs mainly in quartz stringers in altered quartz monzonite in association with breccia or with intensely argillic andesite dykes. Gold is rarely visible and is associated with pyrite and base-metal sulphides. The highest grade sample interval is from a zone of quartz stringers paralleling the breccia, within weak sericitic alteration. A strong, consistent shear structure hosting the local veining and breccia has been traced by trenching over a length of 800 m (Figure 10-8).


Figure 10-8:	Anomalous Gold mineralization location map for South Showing area
Note: DDH locations on Soil Anomaly Base Map

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During 1987, widespread and detailed grid soil sampling programs were undertaken to define areas anomalous in gold. Nine trenches, totalling 1528 m were excavated in two areas (Discovery and South Showings) to test soil geochemical targets, exposed quartz veins and altered breccias hosted in granite. IP, magnetometer and VLF-EM geophysical surveys were carried out over the trenched areas.

In 1989, trenches were excavated in the South Showing.

In 1995, two test pits were dug in the South Showing area to determine the source of anomalous soil geochemistry and to locate the source of a quartz vein float. The pits were dug at grid locations 1800E/125N and 2250E/100N to depths of seven and two meters, respectively. Pit wall soil samples were taken at 1 m intervals. Featureless quartz monzonite bedrock was exposed at the bottom of the pit at 2250E and a diamond drill hole was drilled under the test pit. No significant features were intersected and both pits were backfilled.

Eight holes were drilled in 1995 on three fences oriented at 330°, and spaced 75 and 100 m apart. These holes tested mineralization encountered in 1987 and 1989 trenching programs. A ninth hole was drilled at 330° at the southern end of the showing area, and tested a geochemical anomaly associated with an east-west trending gully. A total of 481 m were drilled (see Table 10-10).

Four holes were drilled in 2010 for a total of 300 m. The purpose of these holes was to verify the 1995 anomalous assays on two separate profiles. SSD-10-003 was targeted to intersect the mineralization between SSD-95-166 and 167. SSD-10-004 was targeted to undercut mineralization intersected in SSD-95-169. The exploration drilling results were included in a press release dated September 9^th, 2010. The significant results are included in Table 10-10.

Table 10-10: Significant drill hole intersections in the South Showing area

Drill hole ID	Au	Downhole width	Includes
	(g/t)	(m)
SSD-95-164	2.48	0.30
	1.35	2.78
SSD-95-165	Nothing significant
SSD-95-166	2.87	1.97	15.47 g/t/0.22 m
SSD-95-167	4.32	0.20
	4.25	0.58
	1.13	9.43
SSD-95-168	2.21	1.40
SSD-95-169	2.18	11.69	32.13 g/t/0.40 m
SSD-95-170	1.58	0.49
SSD-95-171	Nothing significant
SSD-95-202	2.08	0.60
SSD-10-003	0.65	9.29
SSD-10-003	1.14	3.20
SSD-10-004	1.04	7.81

10.2.9 Elusive Creek

In the Elusive Creek area, 4.5 km southwest of the B and WD Vein area, wide-spaced trenches exposed northeast-trending, altered granitic dykes cutting volcanic rocks. Siliceous alteration and quartz vein masses returned elevated gold values, Figure 10-9.

The soil geochemical anomaly for gold in the Elusive Creak area was tested with 5 trenches (1227 m). Quartz veins were chip sampled and returned values up to 5480 ppb Au over 0.5 m. Correlation of the narrow mineralized veins between trenches was not clear.

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In 2006, Almaden completed a program of trenching on the Elusive Creek to follow-up on anomalous soil sample results. No drilling has been completed at Elusive Creek.

Figure 10-9: Anomalous gold mineralization location map for Elusive Creek area

10.2.10 Agur Option Area

The Agur Option area, 3.7 km south of Siwash North, is underlain by a large gold soil geochemical anomaly which has yet to be evaluated by trenching or follow-up work.

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11.	Drilling

Drilling from 1989 through 2007 was supervised by Mr Wojtek Jakubowski, P.Geo and Qualified Person as defined by NI 43-101. LGGC did not observe any aspects of drilling, as there was no active drilling at the time of the site visit.

A total of 833 diamond drill holes (79,321 m) and 90 reverse circulation (RC) (3,626 m) drill holes were drilled on the Elk property. The RC holes were not used in this resource estimate because they were mostly in the pit area which has been mined out and the quality of the logging and assaying was unknown.

The drilling completed to-date (through 2007) on the Elk Property is summarised in Table 11-1.

Table 11-1: Drilling completed from 1986 to 2007

Type of Work	Year	Number of DDHs and total meters drilled
Surface Diamond Drilling	1989-91, 95-96	298 NQ (206 DDH) and HQ (92 DDH) holes 26,019 m
	2000	12 NQ holes 1,414 m
	2002	26 NQ holes 5,007 m
	2003	30 NQ holes 6,570 m
	2004	44 NQ holes 10,298 m
	2005	36 NQ holes 8,395 m
	2006	58 NQ holes, 8,873 m
	2007	9 HQ holes, 2,470 m for resource & metallurgy
Surface Diamond Drilling	2010	87 HQ holes, 12,749 m infill drilling – not included in 2010
		resource estimation
Reverse Circulation Drilling	1992, 1993	90 holes 3,626 m – not used in2010 resource estimation
Underground Diamond Drilling	1994-1995	301 NQ holes -10,018 m

Drill hole identifiers have a three letter prefix (e.g. SND=Siwash North Surface Diamond Drilling and SUD=Siwash Underground Drilling), then two numbers to identify the year followed by a sequential number.

11.1 Surface Drill Holes

The drill hole sites were located on existing roads, drill pads, or in existing cleared areas. Drill hole sites were prepared using an excavator and the drill sumps were dug to contain the drill cuttings. The drill was moved between sites using a D5 crawler tractor. Water was pumped to the drill from Gold Creek or from water collection ponds in the mine area.

Drill collar locations and orientations were initially set using the ground grid coordinates and a Brunton compass. The down hole survey data was collected using a Sperry-Sun single shot camera. Once a drill hole was completed, the collar location was marked with a large PVC pipe, the collar names were engraved onto a metal plate and the locations were surveyed.

Upon receipt from the drill company, the core was washed, footage blocks were converted to meters, and the recovery, RQD (rock quality determination) and hardness were measured. The entire length of the drill core was photographed in groups of four core boxes to each frame, and selected intervals were photographed at five frames per core box for more detailed review. The geology, geotechnical information, and sample intervals were logged onto hand-held HP200LX palm-top computers or (less frequently) graphical logging forms and were later down-loaded onto a desktop computer.

Rock type and alteration logging has been done in great detail, frequently with multiple codes for the same rock type found in the database. To make the geology database more useful, the logging codes should be simplified and made consistent.

Drill hole orientations were measured at surface with a transit or Brunton compass, and down-hole with a Sperry-Sun single shot camera (1989-2003) and with an Icefield down hole survey tool (2004-2005).

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The drill hole collar locations were surveyed relative to pre-established survey control points using a Wild-Leitz T1000 transit.

During the 1995 and the 2003 drill programs, bulk density measurements were made on selected mineralized zones by weighing samples in air and water. There are no further details available regarding the procedures used to measure or calculate the densities. In 1995, 47 measurements were taken and in 2003, 36 results were collected. The density of the waste rock was not measured. During future drilling campaigns, LGGC strongly recommends that Almaden begin a program of regularly capturing the bulk density of vein material and waste rock material during the course of the drill program.

11.1.1 2010 Drilling

In 2010, the drill hole sites were predominantly located on existing roads, drill pads, or in existing cleared areas. Drill hole sites were prepared and the drill was moved between sites using a D5 crawler tractor. The pre-existing drill sumps were used to contain the drill cuttings or shallow trenches were used for containment. Water was pumped to the drill from water collected in the old open pit.

Figure 11-1: Plan of 2010 Infill drilling completed in November 2010

Drill collar locations and orientations were initially set by measuring from old drill collars and a Brunton compass. The down hole survey data was collected using a Reflex digital instrument. Once a drill hole was completed, the collar location was marked with a large PVC pipe, and the collar names were engraved onto a metal plate.

In late November 2010, A BC registered land surveyor was retained by Almaden to survey the collar locations for the 2010 drill holes. The surveyor was asked to survey the collar coordinates for 25 of the historical drill holes to confirm their coordinates in the database (Table 11-2).

The current collar locations in the project database for the historical drill holes were surveyed by the project geologist, W. Jakubowski, using the mine grid only. . The mine grid was established by Jakubowski in 1988 using control point coordinates -2350E and 3000N. The elevation of this point was established from a 1:50,000 topography map (92H16W) and assigned an elevation of 1630.68 MASL. This control point or datum was identified as CP-1.

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In 2004, Jakubowski used a differential GPS unit to determine UTM coordinates and a new elevation for the control points CP1 and BCLS356 (LSM2N1E). In his report entitled “2004 Claim Post Differential Survey” the elevation of the control point BCLS356 (LSM2N1E) is given as 1618.632 MASL and control point CP1 is given as 1620.721 (a difference from 1988 CP1 mine grid elevation of 9.959 m).

In 2010 a registered land surveyor located the control point (survey pin BCLS356) and determined that the elevation of this point is 1618.297 MASL. The control point CP-1 was also located and the elevation was determined to be 1620.422 MASL (a difference with the mine grid elevation of 10.258 m). This is significantly similar to the elevation differences recorded in the 2010 check survey. The average difference in the 25 elevations re-measured was 10.184 m.

Both the discrepancies in the collar elevations and the 2 to 8 m discrepancies in Northing and Eastings of 4 of the 25 (16% error rate) historical drill holes warrants a survey program to include all the surface drill hole locations into the UTM coordinate system prior to updating the current resource estimation or undertaking a pre-feasibility study on the Elk Project. A complete resurvey of the underground drill holes may be too difficult to undertake given that the workings are flooded but the currently accessible survey stations in the decline should be included in the resurvey program and a reliable conversion factor be determined to convert the underground drill holes to the UTM coordinate system. The survey must also reconcile the elevation difference in the two separate datums used, as this will allow the integration of the aerial photography and Lidar planimetric mapping. It is the opinion of LGGC that there is a minimal impact on the current resource estimation that is the subject of this report since all drill holes included in the study were located using mine grid only.

Table 11-2: Difference between Collar Coordinates in the Elk Project Database and the 2010 Surveyed

Coordinates

	Year	Name	Northing	Easting	Elevation
	1990	snd90-73	0.06	-0.41	-10.18
	1995	snd95-138	0.40	-0.03	-10.80
	1995	snd95-146	0.02	0.11	-10.10
	1996	snd96-230	0.05	0.12	-10.19
	1996	snd96-256	0.00	-0.66	-10.11
	2000	snd00-303	-0.05	0.20	-10.09
	2000	snd00-309	0.35	-0.35	-10.28
	2002	snd02-313	-0.32	0.00	-9.41
	2002	snd02-316	0.29	0.12	-10.24
	2002	snd02-324	3.17	0.91	-8.67
	2003	snd03-346	0.05	-0.06	-10.26
	2003	snd03-352	-0.14	0.14	-10.31
	2003	snd03-357	0.08	0.09	-10.36
	2004	snd04-370	0.00	0.09	-10.26
	2004	snd04-381	-8.09	0.84	-10.42
	2004	snd04-394	-0.11	0.05	-10.17
	2005	snd05-410	-0.31	0.38	-10.53
	2005	snd05-418	0.08	0.10	-10.38
	2005	snd05-444	0.10	-0.02	-10.12
	2006	snd06-459	0.09	0.29	-10.14
	2006	snd06-470	-0.01	-0.14	-10.17
	2006	snd06-487	-0.05	0.26	-10.17
	2007	snd07-508	2.61	2.39	-10.27
	2007	snd07-510	2.28	2.77	-10.56
	2007	snd07-511	-0.13	0.13	-10.41

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Upon receipt from the drill company, the core was washed, footage blocks were converted to meters, and the recovery and the RQD (rock quality determination) were measured. The entire length of the drill core was photographed in groups of four core boxes to each frame. The geology and sample intervals were logged onto laptop computers and the geotechnical information was entered on graphical logging forms and was later down-loaded onto a laptop computer.

The geological units and the type and degree of alteration described in the 2010 drill program were based on a core library established from 2006 drill core samples. Typical examples of rock types and alteration were collected and used for comparison to establish continuity between 2010 and historical descriptions.

LGGC had recommended a drilling program to increase the drill hole spacing in the area of detailed exploration and current mineral resources at the Elk Deposit. The list of recommended drill hole locations and orientations suggested drill hole sites in three categories of priority; 1, 2 and 3.

A category 1 drill hole was typically suggested as it would increase the drill hole spacing in an area of inferred mineral resources contained within the SRK base case and alternate case ($1200 USD Au/oz) pit shell outlines. LGGC had recommended infill drilling 121 drill holes for approximately 15,500 m of drilling in category 1 priority. Almaden completed 12,749 m of drilling in 2010 in 87 drill holes and an updated mineral resource may still benefit from drilling more drill holes in the northern and peripheral areas of the veins where drill density is not ideal and greater than 25 to 35 m between drill holes. LGGC had recommended the drilling of 70 further drill holes (13,000 m) to test the mineralization found below the current pit shells produced from SRK s PEA study. These drill holes are a lower priority than the Category 1 drill holes but should be considered for possible expansion of the mineral resources.

There is a list of Category 3 drill holes suggested by LGGC and these targets cover all the possible extensions and infilling potential for deeper vein intersections that may be pursued at a later date and have lower priority than other drill hole targets but should be considered if mineralization extends into these areas of the vein or possible vein material.

11.2 Underground Drill Holes

In 1994 and 1995, 301 underground drill holes were completed and were drilled in fans on north south fences that were spaced at an average 10 m spacing along the main (“A”) and “E” declines. A survey instrument was used to set up the drill hole locations and an inclinometer was used to check the drill hole dip. The down-hole orientation testing was done using acid tests and the final collar locations were surveyed.

Core handling procedures, core logging, sample collection and assaying of the underground drill core are described by Jakubowski (1995) “as per surface holes”.

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12.	Sampling Method and Approach

Drilling and sampling on the Elk property were completed under the supervision of Mr Wojtek Jakubowski, P.Geo. and Qualified Person as defined in NI 43-101 from 1987 through 2006. The procedures described here are taken from Jakubowski (2000) and Giroux (2007) and have been used continuously through all drilling on the property. LGGC did not directly observe core sampling as there was no active drilling during the site visit.

Drilling and sampling on the Elk property were completed under the supervision of Mr Robert Brian Alexander, P.Geo. and Qualified Person as defined in NI 43-101 in 2010. The same procedures as described from Jakubowski (2000) and Giroux (2007) were used in 2010.

12.1 Diamond Drill Hole Samples

How samples were selected for assay appears to be consistent throughout the 1989 to 2006 explorations according to the annual drilling reports (Jakubowski, multiple years). Whether drill core was assayed and which assay method was used was based on the amount of quartz veining and sulphide content in the sample. In places where the quartz vein is narrow, sample intervals straddle the vein and wallrock is included in the sample regardless of other logged geological attributes such as rock type and alteration.

Before the year 2000, the entire drill core sample was sent for preparation and assay so there is no drill core of quartz vein intersections remaining from the early drilling. This practice was common in the industry prior to the introduction of National Instrument 43-101. After 2000 when NI 43-101 had been introduced, all samples sent for assay were split in half using a manual splitting press and every 20^thsample was quarter-split for duplicate analysis.

12.2 Trench and Rock Sampling

As reported in 1991, panel and continuous rock chip samples were typically 0.5 m wide and both the panel and continuous chip samples were taken over intervals averaging 1 m in length using sledge and cold chisel. Panel samples were obtained by chipping five parallel lines of chips along the length of the outlined area to a depth of 2 to 4 cm, depending on the hardness of the rock. Panel samples of strongly altered or soft rock were chipped to a depth of 2 cm over the entire area of the sample. Continuous chip samples consisted of a single line along the sample interval. Aluminium tags with the sample numbers inscribed were nailed to the rock at all four corners of the panel samples and at both ends of the continuous chip samples. Panel samples (15 to 20 kg) and continuous rock chip samples (2 to 5 kg) were placed in numbered plastic sample bags and shipped to Acme Analytical Laboratories, Vancouver, for gold analysis.

Rock samples typically weighing 2 to 5 kg were collected and placed into plastic sample bags. Tyvex® tags with the sample number were attached to stable vegetation close to the sample site. Sites were located by GPS and tied into the local soil sample grid where possible.

12.3 Soil Samples

Multiple campaigns of soil sampling were completed over the Elk Project on recognisance scale and closer detailed spacing to follow up on anomalous results. Reconnaissance soil samples were collected along cut grid lines from the "B" soil horizon and placed in Kraft paper bags numbered with the location identification number. The location was marked with coloured ribbon and a Tyvex® tag labelled with the identification number.

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13.	Sample Preparation, Analyses and Security

The sources of information for this section are Jakubowski (1995-2006), Giroux (2007), Blythe (2007) and Hylands (2007). Information pertaining to the 2010 drill program was obtained from Robert Brian Alexander, P.Geo.


13.1	Drill Core

All drill core samples collected prior to 2010, at the Elk property were shipped to Acme Analytical Laboratories Ltd. in Vancouver, B.C. for sample preparation and gold assaying. ICP (Inductively Coupled Plasma) analyses for 30-35 elements were also performed on samples containing quartz veins.

In 2010, drill core was sent for analysis at ALS Laboratory Group, Minerals Division, in Vancouver, B.C. Analysis was done using 50g fire assay with AA finish for Au (0.005-10ppm) and 50g fire assay with gravimetric finish for Au (0.05-1000ppm). ICP (Inductively Coupled Plasma) analyses for 33 elements were also performed on all samples.

13.1.1 1989-2006

Between 1995 and 2006, sample preparation and assaying methods varied based on the expected grade of the sample, as follows: If the sample was expected to be high grade (typically quartz vein with or without wall rock, at least 10 cm wide with Ã 10% sulphides, +/-visible gold) then the whole core was sent to the lab for screened-fire-assay. The sample was crushed to 3/16” and coarse pulverised to 1/16”. Two kilograms of the 1/16” material was split out and pulverised to 99% passing -150 mesh by sieving on a 150 mesh screen.

One-Assay-Ton (1-AT) of the -150 mesh fraction was assayed for gold and silver, and was combined with the weighted result of gold and silver fire assays (FA) of the entire coarse fraction, to give total gold and silver values. ICP analysis for 35 elements was also carried out on a 0.50 g sample of -100 mesh material.
Samples expected to be of lower grade (typically quartz veins less than 10 cm thick with less than 10% sulphide, and/or wall rock) were sent to the lab for one-assay-ton (1 AT) fire-assay (FA). At the lab, the entire sample was crushed to 3/16”, then 2 kg was split out and coarse pulverised to 1/16”. A 250 g split was then taken and pulverised to -100 mesh. A 1 AT sample was fire assayed for gold and silver and 35 element ICP analysis was usually carried out. Higher grade intercepts were re-assayed using the screened-fire assay method described above.
Samples expected to carry low gold values, typically stringers with varying sulphide, strongly altered wall rock or wall rock flanking well mineralized intercepts, were sent for wet geochemical analysis. This procedure was later changed in 2006 when all the low grade samples were analysed by FA.
The sample was crushed to 3/16” and then a 250 g sample was split and pulverised to -100 mesh. A 20 g sample of the -100 mesh material was analysed for gold ICP-MS using acid extraction. Samples that returned higher than expected values were then assayed by FA, although LGGC found that this was not done consistently. The re-assays generally returned values lower than the originals and upon review, LGGC found that this is probably a function of a sample selection bias.

In 2006, the screened fire assay method was used on 40 samples selected by visual estimation of high grade or visible gold or because the original FA results were higher than expected.

LGGC assigned precedence to the screen assay results over the FA results in the database and the FA assay results took precedent over the ICP results. The original database contained averaged results; LGGC removed all averaged results from the database prior to the grade estimate runs.

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13.1.2 2007

The primary purpose of the 2007 drilling was to obtain fresh (unoxidised) samples from near-surface B vein and deep WD vein material for metallurgical testing (Hylands, 2007). For intersections containing the vein material, a minimum sample width of 2.5 m was used which included hanging wall and footwall material. Pieces of wood, labelled with the sample interval data, were placed in the core box in place of the missing core and the entire sample was shipped to G&T Metallurgical Services Ltd. in Kamloops, BC.

Other samples of altered rock with mineralized quartz veins that were possibly auriferous were also taken. The core was split in half (as in previous years) and the minimum sample width was 30 cm.

Un-mineralized core (blank) and gold standard pulps (from CDN Resource Laboratories Ltd.) were inserted into the sample stream after every thirtieth sample as a check of lab procedures. Split samples were shipped to Acme for preparation and assaying.

13.1.3 2010

One un-mineralized core (blank) and one gold standard pulps (from CDN Resource Laboratories Ltd.) were inserted into the sample stream in every thirty samples as a check of lab procedures. A duplicate sample was also taken in every thirty samples. Due to the high degree of nugget effect in the duplicate samples it was decided to send the entire second half of the drill core, instead of using a quarter split as done previously in the early portion of the 2010 program. Pieces of wood, labelled with the sample interval data, were placed in the core box in place of the missing core. Metal tags, labelled with the sample number, were stapled into the core boxes to mark the start of each sample interval. All split samples were shipped to ALS for preparation and assaying.

13.2 Trench and Rock Sampling

Sample preparation and analysis methods varied, as follows:

Samples that were expected to have significant gold content were crushed to 3/16" and then coarse pulverised to 1/16". In the case of large trench samples, 5 kg of the sample was split out and coarse pulverised. One kg of 1/16" material was pulverised to 99% finer than -100 mesh and sieved on a -100 mesh screen. The result of one FA on -100 mesh material and the weighted result of the FA on the entire coarse fraction were combined for an average gold value. ICP analysis for 30 elements was also carried out on a 0.50 g sample of -100 mesh material. Selected higher grade intercepts were re-sampled from the sample reject and assayed for gold by the same method.

Samples adjacent to those analysed by the screened fire assay method were crushed and coarse pulverised as above, then 250 g of the sample was split out, pulverised to -100 mesh and assayed for gold by FA techniques.

Samples of apparent lower grade were crushed to 3/16", 250 g of sample split out and pulverised to 100 mesh. A 20-g sample of the 100 mesh material was analysed for gold by MIBK (methyl isobutyl ketone) extraction and atomic absorption.

Trench samples were not used by LGGC in the resource estimate.

13.3 Soil Samples

The soil samples were partially dried in camp and shipped to Acme Analytical Laboratories in Vancouver for gold analysis. At the lab, the samples were dried and sieved to obtain 30 grams of - 80 mesh size fraction. This portion was then ignited to 600 degrees Celsius and digested with hot aqua regia. Gold was extracted by MIBK and the solution analysed for gold by graphite furnace atomic absorption.

In 1999, soil samples that were collected in 1987 and 1988 on large grid scale (50 m intervals on 200 m-spaced lines) were re-analysed using 30 element ICP package.

Starting in 1999, soil samples were partially dried at site and shipped to Acme for gold analysis. At the lab, soils were dried and sieved to obtain 20 g of -80 mesh size fraction.

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A 0.5 g sub-sample was digested with 3 ml 2-2-2 HCL-HNO₃-H₂0 at 95ºC for one hour and diluted with 10 ml water and then analysed for 5 or 30 elements by ICP. A 20 g sub-sample of -80 mesh material was digested by aqua regia / MIBK and the solution analysed for gold by graphite furnace atomic absorption. Grid soils were analysed for five elements and the reconnaissance soils were analysed for 30 elements.

Soil samples were not used by LGGC in the resource estimate.

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14.	Data Verification

In order to be satisfied that the project data is appropriate for inclusion in the resource estimation, LGGC reviewed the QAQC data available for the project assay data and completed a database audit. LGGC found the data and QAQC support to be reasonable for inclusion into a mineral resource estimation and has included recommendations to optimise the QAQC protocols for future sampling programs.

14.1 Database Audit by LGGC in 2009

LGGC completed an audit of the project database used to estimate the mineral resources. 10% of the drill holes were selected randomly and the data in the drill hole databases received from Almaden was compared to original data sources such as drill logs and assay certificates.

The error rates for collar survey, down-hole survey and assay databases are within acceptable limits for reporting mineral resources. Minor errors that were identified during the audit were corrected.

The database contained averaged results for the gold grades using assay duplicate data; LGGC removed all averaged results from the database prior to the grade estimate runs.

The geology database containing the detailed geology and alteration codes had a very high error rate. Even so, the database is acceptable for vein interpretations because only veining vs. non-veining and gold grades were used. LGGC recommends that the geology and alteration codes be simplified and made consistent.

14.2 QAQC Data and Review

Prior to NI 43-101 introduction in 2000, the samples did not benefit from a lot of QAQC support during the sample preparation or assaying process. Since 2000 Almaden has improved its QAQC protocols. LGGC has reviewed the available data and finds the database to be sufficiently supported for resource estimation purposes.

Acme made a duplicate pulp for every 20^thsample for internal review purposes and Almaden has been capturing this data in the project database since 2000. The pulp duplicate results appear to have reproduced well, but a minor bias towards the duplicate values returning higher results was observed. The re-assay of the reject material does not appear to show the same bias and the data plots evenly around the parity line. LGGC recommends that this same data be captured in the project database for the assay results from core drilled between 1989 and 1999.

Beginning in 2000, blanks were submitted on a one in twenty basis and were sourced from vein-free, unaltered granodiorite and quartz monzonite core pieces.

There are a few samples that exceed the accepted fail limit of 0.05 g/t Au but there do not appear to be any long periods of contamination problems during the sample preparation process.

Starting in 2000, Almaden took every 30^thsample, changed the sample number and sent the sample back to ACME for duplicate analysis. The results for these resubmission duplicates show good reproducibility of the gold grades but a slight bias is apparent at the higher grade ranges where Acme reported a lesser value on the duplicated samples for analysis results higher than 40 g/t Au. No Standard Reference Material (SRM) samples were submitted with the blind samples to determine if there was a bias at the lab during the duplicated analysis.

From 2000 to 2005, half core and quarter core duplicates were taken every 20^thsample while in 2006, Almaden began sending the other half of the cut core and the frequency of insertion changed from every 20^thto every 30^thsample. The ¼ core to ½ core duplicates show an apparent high bias in the results of the ¼ core samples over the ½ core samples. The results for the ½ core duplicate data compare well and do not appear to show any bias between the duplicate pair analysis results.

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Also starting in 2000, Almaden began sending laboratory check samples to ALS-Chemex Laboratory (ALS) in North Vancouver. These samples show reasonable reproducibility of the gold analysis results between the two laboratories.

The data show a slight bias in the results with the Acme assay values being slightly higher than the ALS-Chemex Laboratory (ALS) results at gold grade ranges above 40 g/t Au. No SRM samples were submitted to Chemex.

In 2003, Almaden began inserting standard reference material (SRM) into the sample stream but they were not assigned a consecutive sample number so there is no assurance that they were analysed within the sample stream. In 2004 sample numbers were assigned to each SRM sample. In 2006, the SRM procedures were modified again – the insertion rates changed from 1 in 20 to 1 in 30, and the weight of the packets of SRM sent to the laboratory changed from 10 and 15 g packets in the early parts of the program, to 30 g packets.

Based on the review of the QAQC data available for the Elk Project, LGGC makes the following recommendations to standardise the QAQC program:

SRM packages should weigh a minimum of 100 g so the laboratory has sufficient material to complete at least two analyses on the same pulp if re-assaying is required.
A wider range of grade values of SRMs should be purchased. LGGC recommends that SRMs with grade ranges around 1.0 g/t Au, 3.0 g/t Au, 5.0 g/t Au, 10 g/t Au and 30 g/t Au be purchased. Two SRMs should be purchased for each of the grade ranges mentioned above with similar expected grades. As an example, if CDN-GS-2E with an expected range of 1.52 g/t is purchased then CDN-GS-2F with an expected range of 2.16 would pair well.
SRM results should be reviewed for compliance immediately upon receipt of the assay data from the laboratory. If two consecutive SRM results fail by two standard deviations on the same certificate then the corresponding sample batches should be re-assayed immediately by the lab. If a single SRM result fails by three standard deviations, then the associated assays in the sample batch should be re-assayed.
Core duplicates should continue to be taken and should be the two halves of the core with a note placed in the sample box as to the nature of the missing sample. Almaden photographs the entire drill core so a photographic record will remain.
Coarse reject duplicates should be taken on a 1 in 20 basis. This is best accomplished by providing the laboratory with two consecutive sample tags to increase the chance that the samples are analysed together.
Pulp duplicates should be taken on a 1 in 20 basis so that the sample numbers are consecutive and Almaden should not rely on the duplicate results from the laboratory.
Laboratory check samples sent to a secondary laboratory and batches of blind resubmissions back to the primary laboratory need to include SRM samples with the shipments on a 1 in 20 basis to confirm the reliability of the results.
Screen analysis method should be used for all samples containing quartz vein material, especially the B and WD Vein Complex samples. Samples outside of the resource area should be analysed using fire assay method with AA finish with a subsequent reassay using gravimetric finish for samples that exceed10 g/t Au For samples that return higher than 10 g/t Au, fire assay method with a gravimetric finish should be used.
All vein samples should be bracket sampled for over 3 m above and below the vein to properly sample the hanging wall and footwall material.
The laboratory duplicate data produced by Acme (Pulp and Coarse Reject Duplicates) should be captured in the project database for the assay results from core drilled between 1989 and 1996.
The Elk samples may benefit by increasing the crush and pulverised sample sizes to 500 g from 250 g.
If Acme has the equipment, then a 1 kg sample size may optimise the sample preparation protocols.

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14.2.1 Standard Reference Material (SRM)

In 2003, certified standards from CDN Resource Labortories (CDN) in Delta BC, were submitted at the rate of one standard for every 20 drill samples.

Sample numbers were not assigned to the SRMs in 2003 so there is no assurance that they were analysed within the sample stream. Twenty-four SRM samples were submitted that year and they were numbered SNSTD-1 to SNSTD-23 and included a SNSTD-13a.

Note: SRM results highlighted in red should be re-assayed and those highlighted in blue are OK and associated with waste intervals

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15.	Adjacent Properties

There are several mineral exploration leases in the vicinity of the Elk property but none are considered significant.

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16.	Metallurgical Testing and Mineral Processing
16.1	Introduction

Metallurgical testing was first performed on this deposit in the period 1989 – 1993 by:

Placer Dome Inc.
Bacon Donaldson
Brenda Process Technology
Nickel Plate mine at Hedley, BC
Andre LaPlante, at McGill.

In 2006 G&T Metallurgical Services, of Kamloops, BC reviewed the previous metallurgical testing studies and prepared a report prior to the 2008 program.

G&T Metallurgical Services has performed all subsequent testing in two programs, the first one ending in April 2008, and the second in April 2010.

16.2 Metallurgical Testing – Historical

At the time of the 1989 – 93 testing, Elk was known to contain high grade gold, so the metallurgical samples were all of a significantly higher grade than at the current resource grades. Typical samples that were tested at that time graded from an infrequent low of 50 g/t Au to > 200 g/t Au.

During the period 1992-94, Fairfield Minerals direct shipped approximately 14,600 tonnes of open pit ore, grading 91 – 137 g/t Au to the Noranda, Horne smelter. In addition, about 1,500 tonnes of underground ore, grading approximately 75 g/t Au was produced.

16.2.1 Placer Dome – 1989 - 90

Placer studied four samples all from trenching, but that data is not available.

16.2.2 Bacon Donaldson (BD) - 1992

BD studied a single 15 kg sample from what appears to drill core produced in 1991. A table in the 2006 G&T report also identifies a 15 kg composite that appears to describe that sample, with a reported grade of 74 g/t Au.

BD however reports grades of 130 g/t Au and 99 g/t Ag. BD performed a single gravity + flotation and a single gravity + cyanidation test. They also reported:

Work Index of 12.0 kwh/tonne
An ICP analysis of the feed
Acid base accounting (ABA) analysis of the feed.

Optical microscopy determined that the dominant sulphide mineral was pyrite, with < 1% as galena, arsenopyrite, chalcopyrite, sphalerite, and pyrrhotite.

The following was determined:

Gravity concentration recovered 62.7% of the gold into 0.05% of the feed weight.
Flotation testing recovered 36.8% of the gold from the same feed sample, to report a combined gravity + flotation (GF) recovery of 99.5%.

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The above flotation concentrate represented 11.9% of the feed weight producing rougher concentrate grading 404 g/t Au.
Gravity + cyanidation reported a total gold recovery of 98.3%, with 63.2% recovered in the gravity stage.
Cyanide consumption was satisfactory at 1.07 kg/t.
Virtually identical gravity stage gold recovery was reported in both tests.
The ABA analysis reported 2.97 % S, with an acid potential of 92.8 kg/t, significantly exceeding the neutralization potential of 18.8 kg/t. From a permitting perspective, this metallurgical sample is PAG. However, after recovering the pyritic flotation concentrate, the flotation tailing was sufficiently depleted of pyrite so that it was non-PAG. The implication is that the flotation concentrate will be strongly PAG, and will need to be stored separate from the flotation tailing if it is not shipped off-site.

16.2.3 Brenda Process Technology - 1992

Brenda received 5 samples including at least one panel sample from open pit mining that was producing high grade ore for direct shipping to the Noranda, Horne smelter. The report provided comments about both oxide and sulphide samples, but they were not specifically identified.

The samples were all high grade in the context of the current resource grade, ranging 58 – 210 g/t Au.

An important observation in the Brenda tests was that, “The ore contains very large volumes of clay material and appears to adversely affect process test work. In particular the filtering of flotation tailing and leach residues was very difficult”.

The report also stated that “tailing” samples (flotation tailing or leached tailing was not stated) were acid consumers.

Brenda reported somewhat lower gravity + flotation recoveries to a maximum of < 92%,

Cyanidation extraction was typically 98.5%, after overcoming liquid / solid separation characteristics associated with “standard” bottle roll cyanidation.

18.1.3.1 Gravity + Flotation Testing

Ten gravity + flotation tests reported the following:

Gravity stage recoveries ranging 24 – 45%, and were essentially independent of the feed grade.
The gravity + rougher flotation recovery, with one exception ranged from 88.0% to 91.7%.
The exception occurred in a single test on the very high grade sample 48-115 (210 g/t Au) reporting only 73.7% gold recovery with rougher tailing grading 58 g/t Au.
The five tests that were performed on “test pit composite” investigated grind sensitivity, reporting reasonably consistent but somewhat random tailing grades with grinds ranging from 50% to 84% minus 200 mesh, and grading 11.9 – 14.3 g/t Au.
Brenda did not report the flotation “test conditions”.

16.2.3.2 Cyanidation Testing

Eight bottle roll cyanidation tests were performed, all on the high-grade “test pit composite”. The first three tests used the “standard” bottle roll procedure, but very poor liquid / solid separation characteristic because of “clay” resulted in a very high retention of pregnant solution in the leached residue.

Subsequent testing used what was described as the CIL procedure. They were not CIL tests, since carbon was added subsequent to the 24-hour sampling interval. The “carbon” procedure was used to facilitate the removal of the leached gold into a conveniently assayable product.

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The five “CIL” tests were run with the following conditions:

Nominal pH of 10.5
Cyanide concentration of 0.7 g/l
72 hours
Grinds ranging from 50% to 80% passing 200 mesh
No gravity concentration stage.

The data reported the following:

Gold extraction in the very narrow range of 98.2 – 98.4%
Consistent cyanide consumption at 3 kg/t
The reported lime consumption was highly variable and unexplained, and may be as high as 5 kg/t
Back calculated feed grades ranging from 83.3 – 121.8 g/t Au
Leaching was essentially completed in 24 hours
Randomness in the leached residue grades ranging 0.67 – 1.93 g/t Au, with these values reflecting the finest and the coarsest grinds, respectively. The data does indicate that the tailing losses decrease at finer grinds, but even the coarsest grind reported an excellent gold extraction. The lack of a prior gravity concentration stage probably somewhat compromised the leached tailing grades in some of the tests.
The testing data implied that this composite was “oxidised”.

In most cases, the laboratory tests included gravity concentration. Given that the ore is reported to be nuggety, the inclusion of a gravity concentration stage in laboratory testing is appropriate to avoid having a “nugget” report to the tailing, and thereby understate what could be accomplished in an operating plant.

16.2.4 Nickel Plate Mine (NPM) - 1992

Samples were collected from the ongoing open pit mining (exploration) and prepared as seven composites and two samples, all of which were high grade ranging 90 – 130 g/t Au with similar silver grades.

Two samples described as “grab” and “ferricrete” grading nominally 130 g/t Au were also included.

The samples were identified as “oxide”, “oxide with sulphide”, and “sulphide”.

The text of the report was brief, but the following was relevant:

All tests were with an “added” 2.5 kg/ tonne ore of NaCN
Cyanide concentrations were not reported
11.0 pH stated, but not reported
Leaching time unstated
Grinds were variable, ranging from 33 to 72% minus 400 mesh.

Although NPM was an operating mine, the testing and reporting was not of a calibre that would be expected of a commercial testing laboratory. Comments about low residual cyanide concentrations were disturbing in the sulphide tests, since that may have been the cause of low gold extractions.

However, the following useful data was reported:

The oxide samples / composites appear to be quite sensitive to grind as noted in eight grind vs extraction tests on “grab”, in which the extraction was only 71.1% with a 10 minute grind (33% minus 400 mesh) increasing steadily to a duplicated > 98% extraction after 120 minutes of grinding to 53% minus 400 mesh. That data was consistent with later testing at G&T.
The gold extraction from the sulphide samples was low with 2.5 kg/t of NaCN, increasing substantially at 7.5 kg/t NaCN, but still only 60 – 70%.

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16.2.5 Andre LaPlante – 1993
This was a brief series of gravity plus staged rougher flotation tests on feed grading 120 g/t Au.

LaPlante reported 66.4% gold recovery in gravity concentration followed by staged rougher flotation increasing it to 95.5%.

16.2.6 G&T Metallurgical Services – Report KM 2121 – 2008

Almaden provided G&T with sufficient samples, all from drill core, to prepare eight composites grading 6 – 47 g/t Au, plus a master composite grading 11.5 g/t Au.

The program included 38 gravity + cyanidation or cyanidation-only tests, followed by two flotation tests.

The relevant gravity + cyanidation testing on the master composite determined the following:

The material is very grind sensitive with the lowest reported leached residue grade of 0.43 g/t Au at P80 = 40 microns, and with the expectation of reduced losses with even finer grinding.
Gravity + cyanidation tests on the 8 composites reported gold extractions ranging 91 – 99%, averaging 95%, of which the gravity stage accounted for an average of 44%., at an average grind of 87 microns.
Gold extraction was insensitive to cyanide concentrations in the range of 250 – 1,000 ppm.
Kinetic testing for 48 hours determined that leaching was complete or virtually complete within that time.
When gravity concentration was included prior to cyanidation, a somewhat higher gold recovery was reported compared to straight cyanidation. That was probably due to batch grinding that of necessity is used in laboratory testing, unlike the superior closed circuit grinding in an operating plant.

The two gravity + flotation tests, both on the master composite, at P80 = 120 microns reported overall gold rougher stage recoveries of 97.1 and 98.0%, decreasing to 94.8 and 95.3% after a single cleaning stage. The gravity stage recoveries were 41 and 47%.

The test determined that bulk sulphide flotation using MIBC frother at natural pH and potassium amyl xanthate (PAX) was effective in recovering gold at a much coarser grind than was required in whole ore cyanidation.

Differential flotation, in an attempt to selectively recover gold away from pyrite, recovered a large portion of the gold in the pyrite concentrate, thus endorsing bulk sulphide flotation to maximise gold recovery

Screen assaying of the rougher tailing in test 40 determined that the material is not grind-sensitive in flotation and that even coarser primary grinding should be evaluated.

16.3 Summary of Historical Metallurgical Testing

At this stage of the cumulative testing programs, it is quite apparent that the following characteristics are relevant to the processing of the ELK deposit:

Both whole ore cyanidation and flotation, with or without prior gravity concentration, represent excellent opportunities to recover gold.
The data determined that cyanidation benefits from very fine grinding that may be finer than P80 = 40 microns.
Based upon very limited data from the G&T KM2121 report at P80 = 120 microns, flotation is not particularly grind sensitive and will achieve >95% gold recoveries into <5% of the plant feed.
The grade of the flotation concentrate (after gravity concentration), at an anticipated 200 g/t Au, is potentially sufficiently high that it could be sold to a regional cyanidation plant. The current project assumes that the cyanidation plant will be on site, but selling concentrate needs to be evaluated at the next stage of the project, since cyanidation has permitting and cost implications.

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16.4	Current Testing Program – G&T Metallurgical Services (2009 – 2010)

The same sample set that was used in the 2008 program were used a second time to prepare two composites to reflect the forecast open pit grade of 5 g/t Au and 15 g/t Au grade from underground mining.

G&T performed several tests with various combinations of gravity and flotation concentration, and both whole ore and concentrate cyanidation, Because of the knowledge that derived from the previous testing program at G&T, this program focused on the influence of grind in both cyanidation and flotation concentration.

The initially cyanidation test did not demonstrate any significant potential benefit from fine grinding as had been demonstrated in the 2008 program. However, the three cyanidation tests on flotation concentrate all reported a very strong benefit from grinding to P80 = < 40 microns prior to cyanidation.

Since it was previously determined that flotation recovery of gold will be very high at potentially > 95% into a low weight percent, the initial whole ore cyanidation testing was performed mainly as a more economical method to determine the grind sensitivity of flotation concentrate to cyanidation than as an assessment of whole ore cyanidation.

This test program confirmed the processing variables that were reported in the 2008 program, as follows:

Gravity concentration will recover a variable 25 – 60% of the contained gold.
The optimum primary grind for flotation is quite coarse. A grind of P80 = 250 microns (or perhaps in operation somewhat coarser) is satisfactory. At that grind, the combined gravity + flotation recovery will be > 95%, with feed from either open pit or underground mining.
Cyanidation of the flotation concentrate can is significantly improved by grinding to P80 = 20 microns, obtaining gold extraction of> 95%.
From a technical perspective, cyanidation circuit gold recoveries will be effective in either the whole ore or concentrate cyanidation mode, but the overall grinding power will be significantly reduced by floating at a very coarse grind, then regrinding to the optimum conditions for cyanidation.
The above grind sensitivity confirms the preliminary selection of a flow sheet that includes: gravity concentration, flotation at P80 = 200 – 250 u, concentrate regrinding to P80 = 20 u, followed by concentrate cyanidation.

Although whole ore cyanidation after fine grinding has demonstrated gold extraction > 95%, the alternative use of flotation to achieve somewhat higher flotation recovery into a very low weight percentage of feed is the preferred option. This is because power consumption will thereby be lower. . The use of flotation also provides an opportunity to market the estimated 25 – 30 tpd of flotation concentrate.

From a technical perspective the very favourable process metallurgical results provide several flowsheet options all of which will include crushing, grinding plus gravity and flotation concentration. Beyond those stages the following all have technical merit, as follows:

1.	Sell both the gravity and flotation concentrates to a smelter.
2.	Sell the gravity concentrate (perhaps as doré) to a precious metal buyer, and sell the flotation concentrate to a cyanidation plant operator after regrinding to optimize the gold recovery from cyanidation.
3.	Sell the gravity concentrate as in 2 above, and sell the flotation concentrate to one of the autoclave plant operators in Nevada since they require additional sulphur as fuel in their whole ore reactors.
	Both have large capacity whole ore cyanidation circuits, and at least one of the two is currently purchasing gold bearing sulphide concentrate.
4.	Cyanide leach the concentrates (one or both of them) and sell only doré.

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From a chemical perspective, if the AP equals the NP, acid will either not be produced or will not escape the disposal site since it will be neutralized by the contained carbonate. In practice, to provide greater assurance that that acid generation will not occur, disposal rock is typically considered to be not potentially acid generating (non-PAG) until the ratio of NP to AP (expressed as neutralization potential ratio, NPR) exceeds 2.

Any material in which the NPR is less than 1, with be PAG.

The NPR in the range of 1 – 2 is neither non-PAG nor PAG, so the data needs to be reviewed on a case by case basis to determine its impact upon the overall site water rock disposal. In the case of the ELK project none of the submitted samples have NPR in the range of 1 – 2, and therefore are either PAG or non-PAG.

The ELK samples all have a paste pH close to the neutral value of 7, indicating that they have not entered an acid generating state.

Table 16-2 summarizes the data for the flotation products.

Table 16-2: Test results from ABA testing of Flotation Concentrates and Tailings



Sample Description	Paste pH	S % total	AP	NP	NPR
B&D MET sample - 1992	n/a	2.97	93	14	0.15
G&T - LG composite - 2010	6.8	2.26	71	11	0.16
G&T - HG composite - 2010	7.2	3.44	107	15	0.14
B & D flotation tails - 1992	5.8	0.10	3	19	6.03
G&T - LG flotation. tails - 2010	n/a	0.12	4	11	2.94
G&T - HG flotation tails - 2010	n/a	0.12	4	15	4.01

Three metallurgical composites all are clearly PAG based upon the less than 1 NPR value in Table 16-2. Note that virtually all of the sulphide sulphur is present as pyrite of which more than 95 % will be recovered into a flotation concentrate.

The mine production will be sent directly to the plant, from which two products will emerge, as follows:

Plant tailing, representing approximately 95 % of the plant feed, will be substantially depleted in pyrite that represents the source of sulphide sulphur, and as shown above has a non-PAG NPR ratio of 2.9 – 6.0.
The flotation concentrate will be highly elevated in sulphide sulphur with a pyrite content of more than 60 %. This product will have an AP greater than 1,000 kg/t, but with virtually no useful NP and therefore will be classified as highly PAG. That product will be either shipped off-site or be stored in a water flooded and membrane lined “pond” where, deprived of oxygen that is necessary for acid generation to occur, it will remain neutral.

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17.	Mineral Resources and Mineral Reserve Estimates

Lions Gate Geological Consulting Inc. (LGGC) was commissioned by Almaden Minerals (Almaden) to complete an updated mineral resource estimate in August 2010 using larger block sizes to support the open pit extraction mining method. This new resource estimate is an update to the one that LGGC completed in 2009 using an updated geology model of the veins and the 2007 diamond drill hole data. GEMS^®, a commercially available exploration and mining software package, was used to estimate the mineral resources.

Susan Lomas, P.Geo., of LGGC reviewed the pertinent geological data in sufficient detail to support the data incorporated into the updated resource estimation and visited the Elk Project on June 23, 2009. While at site, a general review of the project was provided by Morgan Poliquin P.Eng., President and COO of Almaden and drill core collection, sampling procedures and existing drill core relative to the computer database were reviewed. Samples for independent assaying were not collected during the site visit. Susan Lomas of LGGC is the Independent Qualified Person (QP) for the Mineral Resource Estimate as defined by NI 43-101.

17.1 Geology Model

The geology data for the B and WD Veins were reviewed on sections and plans. Each vein was segmented into different domains during the geological modelling process. The solids built to represent the vein sets in the B and WD Vein Complexes include some dilution as each drill hole intercept of the vein was given a minimum thickness of at least 1.5 m down-hole so that the vein solids would have a minimum 1.0 to 1.2 m true thickness. In the absence of any sampling of the footwall or hanging wall material, 0.001 g/t Au grades were used during compositing.

It is recommended that the project may benefit from a re-sampling program to test more hanging wall and footwall core samples to the veins so that any mineralized vein halos can be assessed more consistently. Historically, only the vein intervals were sampled so there are only a few samples in the database that represent the gold grades of the hanging wall and footwall material. When a vein was present in the drill hole but returned no gold values in the assaying, LGGC included these intervals in the vein solids to represent the full spectrum of vein gold grades intersected by the drill holes. When the vein was identified in the drill hole, LGGC did not force a solid through these areas.

The mineralized zones at the B Vein occur in eight sub-vertical to flat-lying horizons that overlie one another. The eight major horizons were assigned the following numeric codes; 1100, 1150, 1200, 1300, 1400, 1500, 1600 and 1700 while all veins in the B Vein have a main domain code of 1000. Domains 1600 and 1700 were modelled into multiple solids and each solid was assigned a sub-domain code; 1610, 1620, 1710 and 1720. Figure 17-1, Figure 17-2, Figure 17-3 and Figure 17-4, show the locations of B Vein domains from different viewpoints. Domains 1100 and 1150, previously referred to as the PC Veins, are deep lying in the western end of the vein set, they are found in the hanging-wall of 1300 and have a more shallow dip than the main vein in this area of the deposit. During data analysis, the assay data for 1100 and 1150 were combined together. Domain 1200 represents a hanging-wall vein found predominantly in the western end of the vein set. It appears that as the vein set flattens at the eastern end, the vein splays into at least four domains that are closely layered. Domain 1300 appears to weaken as it trends east and domain 1400 dominates in the shallow-dipping part of the vein set with three more veins stacked below; 1500, 1600, and 1700.

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Figure 17-1: Plan view of B vein domains

Figure 17-2: Long section view of B vein domains - looking south

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Figure 17-3: Section view of B vein domains (looking east)

Figure 17-4: Section view of B vein domains (looking west)

The mineralized zones at the WD Vein occur in four sub-vertical horizons that overlie one another. The four major horizons were assigned the following numeric codes; 2400, 2500, 2600 and 2700 and all WD Vein blocks have a main code of 2000. Domain 2600 was modelled into two solids and the smaller solid was assigned the same domain code, 2600. Figure 17-5 and Figure 17-6 show the locations of the WD Vein domains from different viewpoints. The main domain in this area is the 2500 vein, with the 2400 vein located in the hanging wall of domain 2500 at the west end of the deposit. Domains 2600 and 2700 are located in the footwall of the 2500 vein.

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Figure 17-5:

Plan view of WD vein domains

Figure 17-6: Long section view of WD Vein Domains (looking south)

The solids for each of the domains were considered hard boundaries during grade interpolation runs and grades from one domain were not allowed to influence the block grades in other domains.

The domain coding was assigned to the block model using the solids and the same integer coding (as shown in Figure 17-1 to Figure 17-6) was assigned to each block.

Drill Hole Data Analysis

LGGC added the data from nine drill holes that Almaden completed in 2007. The final database contains the drilling, assay and geology data for 447 surface diamond drill holes and 301 underground diamond drill holes and from these 419 surface and 290 underground diamond drill holes were included in the grade estimate. The database contained the results for 90 surface reverse circulation drill holes that were used to build the geology model but not for grade estimation purposes.

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A list of the new drill holes is given in Table 17-1.
	Table 17-1:	2007 drill holes in the Elk Deposit Resource Estimate
Hole ID	Easting	Northing	Elevation	Length	Azimuth	Dip
SND07504	2385.80	3133.20	1637.82	222.20	0	-90
SND07505	2261.21	3169.95	1649.78	172.82	0	-55
SND07506	2211.27	3161.33	1650.24	398.37	0	-65
SND07507	2211.45	3160.90	1650.25	288.04	0	-81
SND07508	2261.89	3169.24	1649.90	438.00	0	-60
SND07509	2547.86	3449.84	1648.75	234.09	0	-69
SND07510	2419.02	3426.59	1633.21	243.23	0	-66
SND07511	2203.55	3418.90	1640.01	188.37	0	-73
SND07512	2271.58	3380.51	1640.28	284.38	0	-82

Data analyses were completed using all the assay data for the B and WD Vein domains. Descriptive statistics and histograms were reviewed for the gold grades. The results were used to optimise the interpolation parameters for the mineral resource estimation.

Figure 17-7and Figure 17-8: contain the summary statistics for the uncut gold assay data (g/t) for both the B and WD Veins and waste domains. In B Vein, the 1300 and 1400 domains contain the most assay intervals with a total of 2323 and the remaining 730 assay intervals are distributed into the 1100, 1150, 1200, 1500, 1600 and 1700 domains. In WD Vein, the 2500 domain contains the bulk of the assay intervals with 402 and the remaining 125 were assigned to either the 2400, 2600 or 2700 domain.

Figure 17-7: Box plot and summary statistics for the B vein domains

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Figure 17-8: Box plot and summary statistics for the WD Vein domains (2000s), Waste domains (900 and 0)

17.2	Evaluation of Extreme Grades

Extreme gold grades were examined using histograms, probability plots and decile analysis. Based on their proximity to each other, all the B Vein (1000) domains and all WD Vein (2000) domains were combined for the decile analysis, which is an evaluation of the metal distribution as it relates to the assay frequency distribution and the resulting capping levels were then confirmed as reasonable based on review of the lognormal probability graphs. There are gold values in the 1100, 1300, 1400, 1600 and 2500 domains that exceed 100 g/t Au. LGGC implemented grade caps by zone as outlined in Table 17-2.

In addition to applying limited high grade capping to the assay database, LGGC used a restricted outlier strategy to allow the high grade composites to have a local influence but limit their grade beyond 25 m from a drill hole. The only zone that required this further restriction of high grade material was domain 1300 in the B Vein area. This restriction was put in place during grade interpolation and allowed composites higher than 100 g/t Au to influence blocks within 25 m of the drill hole. Beyond 25 m, the composite grade was reduced to 100 g/t Au.

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Table 17-2: Summary of capping strategy and number of assays capped by zone

Figure 17-9: Cumulative probability plot for gold assay data 1300 Domain (B Vein)

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Figure 17-10: Cumulative probability plot for gold assay data 1400 Domain (B Vein)

Figure 17-11: Cumulative probability plot for gold assay data 2500 Domain (WD Vein)



17.3	Compositing

Assays were composited across the entire width of the vein instead of into smaller lengths as the vein is narrow in places and with this method small composite remnants are not an issue.

Of the 1449 composites created from the assay data, the majority of the composites fall between 1.0 and 2.0 m long. Less than 3% of the composites are less than 1.0 m in length and about 25% of the composites are longer than 2.0 m and 10% of the composites exceed 3.0 m in length (Figure 17-12: and Figure 17-13: ).

With just under 90% of the composites having a length between 1 and 3 m, LGGC used the whole vein composite values to interpolate the grade into the block model.

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Figure 17-14: Block Model definition for the B and WD Vein Block Model



17.5	Bulk Density Values

Bulk Density (BD) measurements were made on core samples at site in both 1995 and 2003. The only description found for how the measurements were made states that core samples were weighed in air and in water and from that the bulk density was calculated. There is no description of how the samples were selected for measurement.

The 1995 data set contains 47 samples with veining which have an average bulk density of 2.84 and 4 samples without veining having a bulk density of 2.73.

The 2003 dataset contains 36 samples with an average bulk density of 2.84.

For tonnage calculations, LGGC used a bulk density of 2.8 for sulphide vein, which is reasonable based on the BD measurements made in earlier drill programs. For future resource estimation updates and refinements, LGGC recommends that Almaden continue to measure the bulk density of core samples. If enough samples are collected then the bulk density can be interpolated as part of the resource estimate and local variations can be realised and shown in the model. Bulk Density measurements using a wax coated method should be made regularly in any future drilling programs.

In previous resource estimates (Giroux, 2007 and others), bulk density values were assumed to be 2.75 for sulphide ore, 2.5 for oxide ore, or were calculated from the Fe, Pb, Cu, Zn contents of the samples when these element analyses were available.

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17.6	Variography

Correlograms were used for modelling the spatial continuity of gold grade composites on the Elk project. Correlograms (referred to as variograms in this report) are the measure of the correlation coefficient between two sets of data, comprising values at the heads and tails of vectors with similar direction and magnitude. The work was done in Sage2001, a commercially available software package.

Variograms were run using 0.75 m capped gold composites for the 1300 domain of the B vein only as the composites from this domain represent the bulk of the data for the deposit. Variograms were attempted on the WD vein composites but no reliable trends resulted from the study so inverse distance method squared (ID²) was used to estimate the grades into the blocks.

The 1300 Domain composites were tagged into two broad subdomains using the 2410 Easting on the property grid to segregate the steeply dipping western side of the deposit from the more shallowly dipping east side. More detailed domaining may be required for future resource estimates of this deposit to better segregate the veins into distinct spatially oriented domains.

To improve the quality of the resulting variograms, lag distances and capping levels were adjusted. The nugget effect was selected from the down-hole variogram. Automatic model fitting was initially used with the final models a result of manual refinement to better represent the geology of the deposit. The variogram models are summarised in Table 17-3.

The nugget effects and first structure account for 80 to 90% of the total variance. The B Vein has a particularly short first range.

Variogram models were used in preliminary confidence limit calculations and aided in determining appropriate search distances for gold grade estimation. Further work is warranted on the variography including subdividing the vein groups into structural domains.

Table 17-3: Preliminary variogram models for the 1300 Vein of the B Vein Complex

17.7 Grade Model and Interpolation Plan

The domain solids were used to assign rock codes (Zone-ID) to the block model. Blocks above the topographic surface were tagged as air (code 100) and blocks outside the B or WD Vein solids, but below topography, were tagged as background (code 900).

A percent model was used to record the percentage of the block inside the solid for each zone code. All blocks that contained greater than 5% vein material were included in the resource estimation.

Values from the vein composites for gold were interpolated into the block model using the inverse distance weighting method to the 2^nd(ID²), ordinary kriging (OK) and nearest neighbour (NN) methods. The NN method assigns the single closest composite grade to the block centre. The NN estimator is a good basis for checking the performance of different estimation methods. Silver values were not interpolated into the block model as there were insufficient data points to have any confidence in the interpolation.

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Interpolation was restricted by Zone-ID so there was no intermixing of grade values between the different zones except between sub-domains 1600, 1610 and 1620 and between 1700, 1710 and 1720.

Search ellipses for the B Vein Complex were divided into two domains, one for the steeply dipping west domain and one for the more gently dipping east domain for veins 1200, 1300, 1400 and 1700.

All the remaining veins were assigned to either the east or west search ellipses based on their being located completely to the east or west of 2410 Easting gridline (1100, 1150, 1500 and 1600.) One search ellipse orientation for the WD veins was sufficient as it is remarkably linear compared to the B Vein Complex.

For all domains a minimum of 4 and maximum of 10 composites were used to interpolate grade into a block, with an additional parameter for maximum composites per drill hole set to 3. This strategy forces the interpolation to require a minimum of two holes within the search ellipse before a grade value can be assigned to a block.

The search ellipse distances were determined using information from the variogram ranges with consideration given to the drill hole spacing. Table 17-4 lists the interpolation parameters used to assign gold grades to the block model for the B and WD Veins.

Table 17-4: Block Model interpolation parameters for B and WD Vein Au grades

17.8	Model Validation
17.8.1	Visual Inspection

LGGC completed a review of the Elk Project resource block model. The model was checked for proper coding of drill hole intervals, assays, composites and block model cells. The coding was found to be properly done.

Interpolated grade in the blocks was examined relative to drill hole composite values by inspecting the sections and plans. The checks showed good agreement between drill hole composite values and model cell values.

17.8.2 Global Means

LGGC also checked the block model estimates for global bias by comparing the average gold grades from the ID²model with means from OK and NN estimates. Results in Table 17-5 show no evidence of bias in the estimate.

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Table 17-5:	Comparison of ID², OK and NN Global Mean values

17.8.3 Swath Plots

LGGC also checked for local trends in the grade estimates by plotting the results from the ID², OK and NN estimate results on easting, northing and elevation swath plots. The ID and OK estimates should be smoother than the NN estimate.

LGGC has included the plots for the gold grades in the 1300 and 2500 domains using the Measured and Indicated Mineral Resource blocks only The 1300 Domain is shown in Figure 17-15: , Figure 17-16:, and Figure 17-17:.

For the 1300 domain of the B Vein Complex, LGGC also plotted the blocks classified into Measured and Indicated Mineral Resources separately.

Based on this review LGGC used the OK gold grades for the blocks classified as measured as these blocks represent the most closely spaced sampling in the deposit to date (area of underground drill holes).

The ID²gold grade in the blocks was used for the Indicated and Inferred blocks. These results are found in the block attribute called “Combined Au Final” of the May 2010 block model folder in the GEMS project.

The results for the gold grades show close tracking between the ID², OK and NN methods and local trends were investigated by reviewing the blocks and composite grades on sections and plans to determine the reason for any local bias. In areas where the NN model is grossly higher or lower than the ID models, sectional review of the block model grades and composites showed an over-influence of high (or low) grades in the blocks or an imbalance in the NN model values.

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Table 17-6: nputs used to build the pit shell for low grade and high grade cut-offs Mineral Resources estimation tabulation of open pit


Item	Assumed Value	Unit
Gold Price	1200	$US/tr.oz
Processing recovery	95	%
Processing Rate	1000	t/day
Mining Cost: waste	2.29	$CAD/t
Mining Cost: ore	8.25	$CAD/t
Processing cost	20.55	$CAD/t
Administration and overheads	2.07	$CAD/t

Table 17-7 reports the results of the resource estimation for the B and WD Vein gold values as of data available from December 2007. Almaden began a drilling program the Elk Deposit in August 2010 but none of this data has been finalised and was not available at the time of this resource estimation.

The 2010 resource estimate result for the Elk Project are being declared using 0.50 Au g/t cut-off for blocks that are within the resource estimation pit shell and a 5.00 Au g/t cut-off for blocks below the pit shell that may be amenable to underground mining methods. The combined Measured and Indicated mineral resources for the B and WD veins, both in the pit shell (reported at 0.50 Au g/t) and below the pit shell (reported at 5.00 Au g/t) are estimated to be 2.2 Mt with a gold grade of 4.26 g/t for 300,000 contained ounces and Inferred mineral resources reported with the same criteria of 1.2 Mt with a gold grade of 7.13 Au g/t and an estimated 263,000 contained ounces.

In late November 2010, at the completion of the 2010 drill program, a possible 10 m discrepancy was identified in the elevation coordinate for 25 out of 25 drill holes (includes drill holes drilled between 1990 to 2007) tested in the area of the resource estimation (Table 11-1). This is a serious concern to LGGC due to the inclusion of Measured Mineral Resources in the estimate inventory and we strongly recommend that no further updates be undertaken on the mineral resource estimate or detailed mine planning be done on the Elk Property until the survey issue is resolved.

For comparative purposes additional cut-offs for inside the pit shell (equal to 0.30, 0.75, 1.00, 1.25 and 1.50 Au g/t) and below the pit shell (at 3.00, 7.00 and 10.00 Au g/t cut-offs) are reported in Table 17-8.


Figure 17-18:	B Vein with classification blocks showing location of Measured, Indicated and Inferred Mineral Resources and location of Pit Shell used to declare Open Pit and Underground extraction methods - Longsection looking north

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Figure 17-19:	B Vein with classification blocks showing location of Measured, Indicated and Inferred Mineral Resources and location of Pit Shell used to declare Open Pit and Underground extraction methods and location of the underground decline - Longsection view from below the deposit, looking south


Figure 17-20:	WD Vein with classification blocks showing location of Indicated and Inferred Mineral Resources and location of Pit Shell used to declare Open Pit and Underground extraction methods - Longsection looking north

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Figure 17-21:	Plan View of B and WD Veins and Resource Estimation Pit Shell to segregate Open Pit potential and Underground potential for extraction method

The resources are reported to a depth of approximately 1300 m elevation, which is approximately 375 m below the surface.

Based on the study herein reported, delineated mineralization of the Elk Project is classified as a mineral resource according to the following definitions from NI 43-101.

“In this Instrument, the terms "mineral resource", "inferred mineral resource", "indicated mineral resource" and "measured mineral resource" have the meanings ascribed to those terms by the Canadian Institute of Mining, Metallurgy and Petroleum, as the CIM Standards on Mineral Resources and Reserves Definitions and Guidelines adopted by CIM Council on 11 December 2005, as those definitions may be amended from time to time by the Canadian Institute of Mining, Metallurgy, and Petroleum.”

“A Mineral Resource is a concentration or occurrence of diamonds, natural solid inorganic material, or natural solid fossilised organic material including base and precious metals, coal, and industrial minerals in or on the Earth s crust in such form and quantity and of such a grade or quality that it has reasonable prospects for economic extraction. The location, quantity, grade, geological characteristics and continuity of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge.” “An „Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drill holes.”

Due to uncertainty associated with Inferred Mineral Resources, additional exploration work on the property may or may not succeed in upgrading the portions of the deposit currently classified as Inferred Mineral Resource to an Indicated or Measured Mineral Resource. . Because confidence in these portions of the estimate is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure, the Inferred Mineral Resources must be excluded from estimates forming the basis of pre-feasibility or feasibility studies but are cautiously accepted for inclusion into PEA studies.

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17.10

Comparison of Updated Mineral Resource Estimation and Historical Resource Estimations

For the purposes of an equitable comparison of the 2010 and 2009 resource estimate by LGGC, to the last reported resource estimate in 2007, the gold grades in the blocks have been tabulated using a global cut-off of 1 g/t Au (Table 17-9 for 2010 LGGC Resource Estimate and Table 17-10 for 2009 LGGC Resource Estimate).

Giroux Consultants (Giroux, 2007) reported the results for the 2007 resource estimate using a 1.00 g/t Au global cut-off and a copy of his table has been included in Table 17-11.

The Measured and Indicated blocks in the 2010 resource estimate report a higher tonnage (2.24 M tonnes) but a lower grade (4.4 g/t Au) and approximately 320,000 contained ounces of gold. The Inferred Mineral Resources have been increased with 2.1 M tonnes reported in 2010 (with an average grade of 5.2 g/t Au) against 0.83 M tonnes reported in 2007 (with an average grade of 7.95 g/t Au). The difference between the two estimates in the Inferred Mineral Resources category is an increase of 130,000 ounces of Au.

Some of the increase in tonnage and loss of grade in the 2010 estimate can be attributed to an updated geological interpretation of the vein sets.

Table 17-9: 2010 Updated Resource Estimate Results for the B and WD Veins reported at a Global 1 g/t cut-off for comparative purposes only


Cut-off Au g/t		Vein	Method	Class	Tonnage	Au Oz
1.00		B&WD	OP&UG	Measured	200,000	55,000
1.00		B&WD	OP&UG	Indicated	2,040,000	263,000
1.00		B&WD	OP&UG	M&I	2,240,000	318,000
1.00		B&WD	OP&UG	Inferred	2,090,000	350,000


Table 17-10:	2009 Updated Resource Estimate Results for the B and WD Veins reported at a Global 1 g/t cut-off for comparative purposes only

Cut-off Au g/t		Vein	Method	Class	Tonnage	Au Oz
1.00		B&WD	OP&UG	Measured	170,000	50,000
1.00		B&WD	OP&UG	Indicated	1,400,000	240,000
1.00		B&WD	OP&UG	M&I	1,570,000	300,000
1.00		B&WD	OP&UG	Inferred	1,860,000	360,000


Table 17-11:	B and WD Vein Resource Estimate Results reported at a Global 1 g/t cut-off (includes both near surface and deeper grade blocks) (Giroux, November 2007)

17.11 Mineral Reserves

There are currently no mineral reserves outlined on the project. Historical mineral reserves have been discussed in Section 6.2 of this report.

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18.	Other Relevant Data and Information – Preliminary Economic Assessment
18.1	Introduction

In summary, the objective of the PEA was to define an economic mining scenario for the Elk Project using an open pit methodology, whilst also identifying areas of risk and upside.

This included:

Developing an appropriate cost model for a suitable operation Open pit optimisation to define an appropriate in-pit resource Conceptual scheduling and financial modelling

The final element was to incorporate the PEA into an NI 43-101 compliant Technical Report.

This mining study and PEA are at a conceptual level where different options can be considered and a broad understanding of the potential project performance can be gained.

18.2 Mining Method

SRK has considered only an open pit extraction method for mining the mineralized vein material at the Elk Project for this PEA study. However, the resource estimate contains resources reported at a 5.0 g/t Au cutoff that may be amenable to underground mining methods, and this must be kept in mind in the further development of the property. If the property is confirmed as a viable open pit, then access from the pit to underground workings should be integrated into the pit design. There will be an optimum level for the open pit below which an underground operation will be more economic, and this level should be studied further as part of a pre-feasibility study and allowed for in the pit design. Underground mining considered as an extension to the project could be attractive, because access from the pit could be relatively quick and inexpensive, and the processing plant and infrastructure would already be built and operating.

The understanding gained from drilling and trial mining the ore body indicates that a selective approach to mining is necessary to maximise recovery of the relatively small tonnages of valuable vein material.

It is assumed that the mine will use conventional drill and blast mining methods and mid-sized mining equipment such as 120 t excavators and 100 t trucks for mining most of the waste. The operation will work two eleven hour shifts seven days per week and will be carried out by a mining contractor.

When mining is close to the mineralized veins, flitches will be reduced in height to 2.5 m, and the excavators, assisted by bulldozers will clean off the hanging wall waste working from the top and bottom of the flitch. Cleaning the hanging wall waste immediately adjacent to the vein and then digging the vein will be carried out only on day shift .The contractor nominated a fleet of smaller equipment for vein mining (30-50 t excavators and matching trucks), although it may be found that a larger excavator will be better because of its superior horizontal reach. Such an excavator might work vein during dayshift and move to waste, with more trucks, at night.

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Based on previous experience at this property, it is believed that the vein material will not require blasting, but the costs allow for mechanical rock breaking and limited drilling and blasting if required. Blasting is simple to carry out, but may cause losses of vein, and possibly disproportionately large losses of gold in fines, so will be avoided if possible. The way the vein peels away from the hanging and foot walls will also influence the success of the selective mining .The mine policy will always be to include waste in the plant feed rather than to lose gold.

The vein material will be mined irregularly and only on day shift. It will be stockpiled as necessary, and reclaimed when pit output is inadequate for the needs of the processing plant.

Mining of the footwall will require a dedicated track-mounted drill to stand on the footwall and drill the part bench, or to drill horizontal holes from the flitch below. This drill can also be used for drilling and blasting vein if necessary.

Snowfall would impede selective operations and a shorter operating season of only eight months per year is envisaged. In summary, the vein mining is seen as a low –productivity, high- cost activity.

Waste will be hauled to external dumps; dumps will be re-contoured and revegetated.

18.2.1 Pit Wall Slope Angle

Golder Associates provided an inter-ramp slope angle of 50°. Using a road width of 25 m, suitable for the Caterpillar 777 trucks proposed by the contractor, gives an overall slope angle of 46.8° for a single ramp and 43.9° for a slope with two ramps. The inclusion of a ramp and berms is in practice what would occur and as such a slope angle of 45° is more realistic and was used. There is definitely scope for improvement in this, with potential to reduce the quantity of waste mined.

18.2.2 Ore Loss and Dilution

The selective approach described above is appropriate given the potentially very high value of the ore. It is preferable to dilute the ore with minor amounts of wall rock rather than incorrectly mine the ore as waste. SRK therefore assumed the mine would tolerate 10% dilution and avoid any ore loss.

18.2.3 Waste Rock Disposal

Waste will be hauled to external dumps. Open pit mining of the narrow veins will incur high stripping ratios at 16:1 for the Base Case scenario, increasing to 30:1 for the $US1200 scenario. These options will produce 21.6 and 77.9 million tonnes of waste rock respectively, of which only about 0.5 million tonnes will be consumed in tailing dam construction. The dump will occupy an area of 85-100 Ha. Almaden s recent accurate topographic survey indicates the there are several choices of location for dumps, but the overall surface design has only just started, and dump siting must be integrated with the location of the TSF and other facilities. An initial concept is presented as Figure 18-1. This concept has not been tested in any way.

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Figure 18-1: Initial Conceptual Project Layout $US1200 case

18.3 Processing

Processing is described in Chapter 16, Section 6, where several processing options are discussed. For the PEA it is assumed that: The processing will involve crushing and grinding followed by gravity and flotation circuits at a plant production rate of 500 tpd (180 ktpa).

The 25-30 tpd of gravity and flotation concentrates will be treated on site by regrinding and then cyanide leaching.

Overall, metallurgical recovery of gold is estimated to be 92%.

Tailings disposal will be in two facilities, a small, lined dam for the cyanide process tailings and a larger dam for the flotation tailings. Water will be recirculated from both dams back to the plant, and there will be no positive discharge of water from either dam during operation.

A full closure plan for these dams will be made as an integral part of their design during the Feasibility Study.

18.3.1 Process Flow Sheet

The flow sheet adopted is shown in Figure 18-2.

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Figure 18-2: Preliminary Process Flow

The plant will be constructed using a significant portion of the plant equipment inventory that was purchased several years ago at the disposal of the 100 tpd Willow Creek Mine in Alaska. That equipment is stored at Savona Equipment Ltd, in Savona BC, with the exception of the plant generators that were relocated to a storage facility in the Okanagan. Note that the purchase did not include the buildings, so other than a few sections of platform steel from the crushing plant, all other structural steel requirements will need to be designed, fabricated, and erected.

The purchase included good quality electrical motor control centres that will comply with the Canadian electrical code.

Also included in the purchase was the assay laboratory equipment that is discussed in Section 18.4.3.

A complete list and valuation of the Willow Creek plant and equipment including an identification of those items assumed to be re- used at Elk Gold is available (Willow Creek Gold Mill, June 9 2003, memorandum by Gary Hawthorn.)

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18.3.2 Design Criteria

The simplified Design Criteria is tabled in Table 18-1.

Table 18-1: Simplified design criteria

Item	Unit	Open Pit
Feed rate	dmt/day	500
Feed	Au – g/t	4-8
Work Index	kwh/tonne	14
Flotation feed	P80 = microns	>200
Flotation conc	Wt%	<.5
Gold Recovery – gravity + flotation	%	95
Concentrate regrinding	P80 = microns	< 40
Gold Recovery - cyanidation stage	%	96- 98
Combined recovery	%	92

18.3.3 Tailing Storage Facility

For purposes of this study, it is assumed that a 2.5 year starter dam will be constructed for which an allowance has been provided in the capital cost. All of the dam construction is anticipated to use non-PAG waste rock from the open pit as part of normal waste removal.

18.4	Infrastructure
18.4.1	Power Supply

At the proposed 500 tpd production rate, electrical power requirements that will be provided by Almaden are limited to the plant and to the site offices, since the mining contractor will be providing his electrical system. The power generation requirement is as follows: Plant - 700 kW @ 3/60/460 Site – 100 kW @ 3/50/460 partially transformed to 110 / 220 single phase.

Because of the high cost of diesel electric power generation it is important to both operate generators at their peak power generation efficiency and to match the generators to usage. Since the crushing plant will operate for only about 12 hours per day, it is proposed that the two Almaden owned 330 kW generators be dedicated to the crushing plant, with one on standby. For the balance of the plant two used 500 kW, low operating hour and “enclosed” generators will be purchased, with one on standby.

The site (including the assay office) will be serviced by a 2 - 100 kW 3 phase generators.

If the site is subsequently serviced by grid power, the installed generators will be retained for standby power.

18.4.2	Water Supply
A	reliable water supply has not yet been proven, but baseline measurements of rainfall and streamflow have

been made for several years. Water from the mine, or from nearby streams or lakes, or from underground sources may be used. This has to be investigated early in the next phase of development. The mine drainage, surface flows, and return water from TSF will probably all be used in an integrated water management system.

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18.4.3 Assay Laboratory
The purchase of the plant equipment at Willow Creek Alaska included the assay laboratory. Several major components are satisfactory, but others are now obsolete, and will have to be replaced.

A sample preparation and fire assay laboratory with gravimetric finishing will be provided on site to process the samples that will be assayed each day, including the samples from mine production and perhaps as many as 30 samples from the plant operation. At this stage of the project it is anticipated that atomic adsorption equipment will not be provided. That will change if routine assaying benefits for wet chemical assaying, or AA finishing, or there is a need to routinely assay very low grade gold samples, at < 0.5 g/t Au.

18.4.4 Miscellaneous Infrastructure

Allowance for some offices, workshops and stores including a fuel store has been made. It is allowed that the mining contactor will provide any infrastructure required for his fleet and personnel. Contractor and owner s personnel will live mainly in the Merritt or Peachland areas.

A Vancouver corporate office is allowed for, for product sales, logistics, and accounting.

18.5 Product Marketing

This study has assumed that doré gold will be produced on site. One reason for using this assumption is that doré gold can be readily marketed in small quantities to several competing refineries in Canada or USA, at a price that is significantly higher than for gold contained in sulphide concentrate. The freight and refining cost has been estimated as 2% of the gold spot price.

Flotation concentrate containing gold could be marketed in containers to a smelter in USA, Canada, or North Asia. Depending on the nature of the feed from their major customers, some smelters may be appreciative of the sulphur in the Almaden concentrate as a fuel source. Some will be already engaged in recovering gold from anode slimes and able to easily recover Almaden s gold. However, it is unlikely that a smelter would be prepared to pay highly for a spot shipment. They would expect a commitment of Elk production for a period of years.

The on-site cost of concentrate leaching and doré production is obviously higher than just producing concentrate. However if it was to be sold, the concentrate would then have to be dried, bagged containerised and freighted to a distant smelter. It may then be sold for about 80% of the gold spot price.

No sales contracts are currently in place.

18.6 Operation Cost Model

At this stage of project study only a simple cost model is possible, based on the assumptions stated above.

All costs quoted are in $CAD unless stated otherwise.

18.6.1 Mining

The mining operating costs are based on indicative pricing recently provided by a contract mining organisation with first-hand knowledge of the pit, the area, and the climate. The proposal is based on mining a total of 33 Mt of waste and 1 Mt of ore over a period of 5 years. A moderate-sized waste mining fleet and a smaller ore mining fleet will be used (see Section 18.2). SRK is of the opinion that this is a reasonable scenario and the costs provided are reasonable indications to use as a basis for the PEA study. However, the wide range of production rates and pits considered meant that the costs had to be adjusted for each different scenario examined.

In addition to the mining contractor, provision is made for $3.29/ per tonne of ore to cover the owner s supervision and overhead mining costs. These include site supervision, sampling and grade control, survey control, mine planning and scheduling; as shown in Table 18-2. Assaying is included in laboratory costs in the processing section.

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Table 18-2: Mining Supervision and Overhead costs





Item	No.	Unit $CAD per month	Total per month
Mine Supt	1	11000	11,000
Mining Engineer	1	7000	7,000
Secretary	1	2000	2,000
Surveyor	1	6000	6,000
Assistant	1	1800	1,800
Geologist	1	7000	7,000
Grade controller	2	2500	5,000
			39,800
Consumables	1	2000	2,000
Vehicle operation	3	2500	7,500
		Total $CAD /mo.	49,300
		Average per tonne	3.29

18.6.2 Processing

Based upon a plant feed rate of 500 tonnes per day, the operating cost is forecast in Table 18-3.

Table 18-3: Commination, gravity and flotation operating costs


Labour	$/month	$/tonne	Dist%
Mill superintendent	10,000
Crushing operators	13,680
Grinding Operators	16,416
Flotation Operators	20,064
Operations - day shift only	3,909
Electrician	4,777
Mechanics	9,554
Sub Total personnel	78400	5.23	31.0
Supplies
Liners	4,500
Balls	28,350
Xanthate	1,875
Frother	600
Miscellaneous - operating	15,000
Miscellaneous - maintenance	15,000
Sub Total supplies	65,325	4.36	25.0
Power	100,050	6.67	39.0
Assaying	9,630	0.64	4.0
Heavy Equipment fuel	1,500
Heavy Equipment repairs	1,500
Subtotal - services	112,680	7.51	44.0
Total – gravity + flotation	256,405	17.09	100.0

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Table 18-4: Cyanidation and electrowinning operating costs



Labour	$/month	$/tonne of plant feed
Cyanidation Operators	9,554	0.64
Supplies
NaCN	18,750
Lime	300
Refinery fluxes	200
Miscellaneous - operating	7,500
Miscellaneous - maintenance	3,750
Total supplies	30,500	2.03
Power	13,800
Total - cyanidation	53.854	3.59

The operating cost estimate includes a 25 % payroll burden.

It assumed that electrical power is provided by on-site diesel electric generators with an operating cost for the generators of approximately $ 0.23/kWh. If the capital cost to connect to grid power is justifiable, the unit power cost is expected to decrease to an estimated $ 0.05/kWh, thus reducing the operating cost by approximately $ 5.50/tonne of plant feed.

18.6.3 Administration and Overheads

In the cost model provision is made for $2.07 per tonne of ore to cover General and Administration (G & A) costs associated with the mine. These include head office salaries, and expenses for consumables and vehicles. Details are given in Table 18-5.

Table 18-5:

Administration and overheads costs


	Administration and overheads costs
Item	No.	Unit $CAD per month	Total per month
Gen Manager	1	14000	14000
CFO	1	10000	10000
Clerk	1	3000	3000
Consumables	1	2000	2000
Vehicles	2	1000	2000
		Total $CAD /mo.	31000
		Average per tonne	2.07

18.7	Open Pit Optimisation
18.7.1	Introduction

Open pit optimisation was used to identify the optimum economic pit shapes and pit limits, which would achieve the maximum project cash flow.

The pit optimisation process seeks a solution to a three-dimensional mathematical relationship involving the mineral resource model, geotechnical slope guidelines, product revenue, project constraints, and operating costs. The Whittle software package was used to carry out the open pit optimisation process for this study. Whittle software calculates the financial value of the blocks in the mineral resource model and then calculates a solution observing the geometrical constraints. Waste that must be mined to access the ore as a function of the overall pit slope angles is taken into account in the solution.

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18.7.2 Block Model

The resource block model (see Chapter 17) developed by LGGC was used in the optimisation. The model has a block size of 10 m x 10 m x 5 m (X Y Z) and had fields with gold grade (in ppm) reporting class, rock type and proportion of the block within a vein model. A bulk density of 2.7 was assumed for waste and 2.8 for mineralized vein material.

18.7.3 Input Parameters

The costs are discussed in Section 18.6. Input parameters are summarized in Table 18-6.

Table 18-6: Whittle optimisation parameters


Item	Value	Unit
Slope Angle	45	°
Mining Costs	2.29	$/tonne waste
	8.25	$/tonne ore
Mining Recovery	1.0
Mining Dilution	1.1
Processing Cost including G & A	22.75	$/tonne
Processing Recovery	92	%
Gold price	1000	USD
US Dollar / CAD Dollar exchange rate	0.95
Selling Cost	2%	Of nominal gold sale price

18.7.4 Optimisation Results

The optimisation output is a sequence of three-dimensional pit outlines called pit shells. These nested pit shells satisfy the overall pit slope factors and specific economic parameters applied to the pit optimisation.

The range of pit shells varies according to a Revenue Factor. At a Revenue Factor of 1.0, the optimum pit shell is found, where the marginal cost for an additional unit of product is equal to the net revenue received for that additional unit of product. This solution is specific to the input parameters used. Revenue factors greater than 1.0 generate larger pits that produce more gold, but decrease profits. Figure 18-3 shows the ore and waste tonnes for a series of different revenue factors for the optimisation results.

Significant changes to waste and ore tonnes at small changes in revenue factor indicate step changes to the mining with associated opportunities or risks. At about RF1.1 there is a large spike in the waste tonnes to be moved. This is caused by high grade ore at depth in previously unworked areas. When the price is high enough for this to become economic, a whole new area of the pit is opened from surface. The pit selected in this study is RF0.70, in a much more conservative and less risky part of the chart. This is discussed in the following sections.

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Figure 18-3: Whittle results – graph of tonnes by revenue factor

18.8 Production Scheduling

To provide an annual mining schedule for financial analysis, the material in the selected pit shell has to be scheduled.

The sequence of benches mined in each pushback is scheduled either to optimise the NPV or to meet mining and processing throughput limits or targets. The sequencing algorithm used by Whittle is called the Milawa algorithm. In practice the choice of pushback pits can influence how well the scheduling meets the target criteria and a number of iterations are normally performed.

18.8.1 Alternative approaches considered

Many alternative scenarios were examined, including:

Mining the RF 1 pit at various rates. The ore throughput to the plant is fixed given that the plant is already owned. However, the mining contractor proposed several schedules where the waste and ore were mined faster than required by the plant, which saves on indirect costs, but gives extra cost for reclaiming ore from stockpiles. This approach was adopted.
Mining a smaller pit (RF 0.70 was chosen) and scheduling this in various ways was attempted.
Building a 20 k power line to bring 1.5 Mw of power from the grid to site was considered. This would reduce power costs from $0.23/kwh when using diesel power, to $0.05/kwh for grid power.

While circumstances may change in future, it was found that under the gold price and costs assumed, the best project was one based on the RF 0.70 pit , mining and processing at matching rates, and using diesel power. Mining a smaller pit than the maximum does not preclude mining a pushback later, depending on the gold price outlook in a few years years time.

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A larger plant and a power line might be worthwhile once operating experience and more knowledge of the deposit has been gained. Meanwhile the proposed pit yields a seven year project if using the already-owned plant. It gives a robust return, and is a conservative model with relatively low risk.

18.8.2 Scheduling Results

The proposed annual schedule of ore and waste tonnes and grades is shown in Table 18-7.


	Table 18-7: Open pit mining schedule for Chosen Case (RF 0.70)

Year	Ore Tonnes	Waste Tonnes	Waste: Ore Ratio	Headgrade
				(g/t)
1	92214	3407786	36.96	3.26
2	179892	3320108	18.46	5.44
3	179941	3320059	18.45	3.04
4	179998	3320002	18.44	4.09
5	180000	2853957	15.86	3.46
6	180000	1297329	7.21	4.55
7	109105	503551	4.62	5.07
Total	1101150	18022792	16.37	4.14

Of the ore to be mined in the schedule, 9% is based on Measured Resources, 73% on Indicated and 18% Inferred Resources. The study and its conclusions are thus provisional. One outcome of a study like this is that those areas needing more drilling are defined.

The base case pit shell with its included mineralization is shown in Figure18-4:

Figure18-4: Base case Pit Shell (green) with included mineralization (red)

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18.9	Operating costs
The estimated operating costs are as per Table 18-8.
	Table 18-8:	Unit Operating costs
		Unit cost
	Cost centre		Unit
		$CAD
	Waste Mining	2.42	/tonne waste
	Ore Mining	8.38	/tonne ore
	Processing	20.68	/tonne ore
	Administration and Overheads	2.07	/tonne ore

18.10 Capital Costs

Capital costs are not used to generate the Whittle optimised shells, but are important when considering the overall financial performance of the project and selecting the shell to use. As with operating costs, this Preliminary Assessment does not require detailed estimates, but every effort was made to make them realistic.

18.10.1 Mining Capital Costs

By using contract mining, Almaden s mining related capital expenditure is minimised. The mining related capital costs that the owner will incur are in. Table 18-9. The needs of the processing plant staff in the office and change house have been allowed for.

Table 18-9:

Mining capital cost estimate


Mine	$CAD
Mine/Plant office	150,000
Change house(co only)	40,000
Store/Workshop	60,000
Light vehicle workshop/ service	30,000
Fuel Bowser & Tank	30,000
Explosives Magazines	200,000
Light vehicles	150,000
Computers & software	150,000
Admin/stores equipment	75,000
EPCM	48,896
Total Mine Admin plant and equipment	933,896
Consumables First Fill	11,250
Contractor Mobilization	1,697,152
Sub Total	2,642,298

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18.10.2 Processing Capital Costs

The capital cost shown in Table 18-10 is for a 500 tpd gravity and flotation concentration plant with cyanide leaching of the flotation concentrate, plus the assay laboratory, using a blend of plant equipment that was purchased from the disposal of the 100 tpd Willow Creek plant in Alaska and reconditioned equipment.

In addition, provision has been made for an assay laboratory and building of two tailings facilities.

A summary of the processing capital costs is shown in Table 18-10.


Table 18-10:	Processing Capital Cost Estimate

Processing plant		$
Site preparation		50,000
Building foundation		172,200
Machinery Foundations		86,100
Building Structure		430,500
Structural steel		344,400
Plant equipment		1,722,000
Installation labour		1,377,600
Electrical		344,400
Piping		430,500
Laboratory including building		75,000
Tailings disposal plant & equipment		200,000
Tailings disposal construction		290,684
Direct Capex		5,523,384
EPCM		828,508
Total		6,351,891
Contingency-8%		523,270
Total Process/Admin		6,875,161
First Fill Consumables		128,595
Sub Total Plant Power and Tailings		7,003,756

Sunk costs for equipment from the Willow Creek plant purchase, at $403,000, have been included above.

The above capital cost forecast is equivalent to $CAD14,000/tonne per day of plant throughput.

Included in the above is $75,000 to supplement the assaying equipment that was acquired with the Willow Creek equipment, and to construct the assay laboratory.

The cost of the main tailings dam wall is not included above, but it is allowed for in the mining costs, as it is part of the waste removal operation.

18.10.3 Offsite Infrastructure

No allowance for offsite infrastructure was made.

18.10.4 Other

No other capital costs have been included.

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18.11 Financial Modelling

A simple financial model of the project was constructed. The basis of the model was the production schedule, discussed in Section. This was combined with the operating and capital costs estimates.

A discount rate of 8% was used. The projection is included as Appendix 2.

18.11.1 Modelling Results

Some results of the modelling are given in Table 18-11 for gold at a market price of $US1000/tr.oz.


Table 18-11: Project Financial Indicators

Financial indicator		Value	Unit
Pre-tax NPV @	8%	28.7	$CAD M
Pre -tax IRR		51%
Payback, years from start production		1.85	Years
Times covered		5.02
Max exposure		13.66	$CAD M
Tr.oz Au produced		139,198	tr.oz
Op cost/tr.oz, $		528	$US/tr.oz

The reader is advised that this analysis is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves. There is no certainty that the preliminary assessment will ever be realised, in whole or in part.

18.11.2 Sensitivities

A “two-parameter” sensitivity analysis is shown in Table 18-12. Variation in gold price is about 1.4 times more significant than total working costs.


Table 18-12: Two-parameter sensitivity analysis
					Revenue
NPV ($CADM)					Sensitivity
	$US/tr.oz	700	800	900	1000	1100	1200	1300
		-30%	-20%	-10%	0%	10%	20%	30%
	-15%	10	20	29	39	48	58	67
	-10%	7	16	26	35	45	58	64
Total Working Costs	-5%	4	13	23	32	42	51	60
Sensitivity	0%	-0	10	19	29	38	48	57
	5%	-3	6	16	25	35	44	54
	10%	-6	3	13	22	33	41	50
	15%	-10	-0	9	19	28	38	47

It can also be seen that the project shows robustness to a drop in gold price. In fact, it could cover operating and financing costs at $US 695/tr.oz gold, and shows cash breakeven at $US 528/tr.oz.

With an initial capital expenditure of only $CAD 12.2M the project shows little sensitivity to capex.

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The project is relatively more sensitive to operating costs, in particular the cost of waste removal. Ore-related costs for mining or processing are of relatively less importance, and a 26% reduction in overall processing costs by using mains power is not a large benefit, although it may give more stable power supply and freedom from “oil uncertainty”, and the sovereign risk of carbon taxes.

The project s total earnings provide good cover for its capital expenditure; it has a reasonable payback period. There is scope for improvement in mine design and scheduling, once more of the ore has been converted to measured or indicated status.

The project is sensitive to slope angles, and these need careful study. The presently adopted 45° slopes are realistic in parts of the pit, but may be over-conservative in others.

The design and layout of waste dumps is very important, much more so than the siting of the processing plant and facilities. Early waste hauling in particular must be kept as inexpensive as possible.

18.11.3 Project Risks

The sensitivity studies and the engineering descriptions above allow some project risks to be identified.

These are discussed below .Risk is a matter of opinion at this level of study, as the basic parameters themselves are ill-defined, so no quantification has attempted. The apparently high rate of return is realistic for this small mine, at it indicates robustness to adversity.

Gold Price. Low at $US1000 used. In addition, the Base Case pit shell used is smaller than the theoretical optimum and so is capable of withstanding low prices.

Resource risk. Some risk remains here ,and Almaden have been infill drilling this year in order to convert the inferred part of the resource to indicated or measured .The new holes have not been taken into account in the resource estimate.

Mining Waste. The large waste quantity is the main source of risk in mining. A small increase in quantity or unit cost is the largest source of risk to the project operating costs. There is considered to be ample space for waste dumps, but the suitable locations have as yet not been permitted, proved to be unmineralized, or proved to be geotechnically sound. The waste has recently been tested and to be non-PAG. Slope stability and the angle of 45° used have an important bearing on the quantity of waste that must be mined. The angle is based on a professional assessment of the trial pit by Golder Engineering, but more testing is needed on the larger pit now proposed.

Mining Vein. It is considered that enough has been allowed to cover the method proposed, even for some blasting of the exposed vein if required. Snow will impede the selective mining of vein, and a shorter production period has been adopted. In view of the relatively high importance of waste mining, the ore mining cost risk is low, but mining dilution and recovery are of more importance. If vein mining proves more difficult than allowed for, the result is likely to be increased dilution and a lower process plant feed grade. With a processing plant of limited size, this in turn means that the gold production rate will be decreased. Part of the recommended rock mechanics budget is for study of the ore hardness and its likelihood of parting easily from the wall rocks.

Processing. Like vein mining, the operating cost risk is relatively low. Test work has indicated a high recovery, so upside is unlikely. To mitigate any risk, more test work is recommended on vein scheduled to be produced early in the project s life, using local water. The capital costs for commissioning a plant with used equipment are less certain than for a new one, but the capital cost is in any case low so the model does not exhibit sensitivity to this.

On-site Leaching. The study assumes on-site cyanide leaching of the flotation concentrate. If permission cannot be obtained for this, the high grade and low weight concentrate produced should be saleable, but this product is not as readily marketable as doré bullion. This option would lower capital and operating costs, but there would be freight costs and a lower realised gold price. The uncertainty about concentrate realisation price led to doré production being assumed for this preliminary assessment.

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Permitting. The sovereign risk of permitting this small project is possibly higher in its location than elsewhere in the world, but there do not appear to be any conservation areas, buildings, agriculture, archaeological sites or tourist objects within 1.5 km of the site.

18.12 Other Cases Considered

According to the Optimiser, it would be possible at the $US1000 gold price to mine more mineralization economically, (about 1.7 Mt in all), and achieve a higher overall project NPV in a 10-year project. However, the project actually selected gives a faster payback and a higher IRR, and is overall a lower risk project. Mining the smaller pit as proposed does not preclude expansion later, particularly if the gold price is sustained at current levels.

Table 18-13 shows that higher gold prices would give a higher project NPV, and there are other significant benefits. This table may be compared to the same project at $US1000/tr.oz gold in Table 18-11.

Table 18-13: Base case outcomes with $US1200/tr.oz gold price




Financial indicator		Value	Unit
Pre-tax NPV @	8%	47.6	$CAD M
Pre -tax IRR		77%
Payback, years from start production		1.54	Years
Times covered		7.3
Max exposure		11.82	$CAD M
Tr.oz Au produced		139,198	tr.oz
Op cost/tr.oz, $		528	$US/tr.oz

However, if the project were to be based on an expectation of current gold prices being sustained, at say

$US1200/ tr.oz instead of the conservative $US1000/tr.oz as described above; many other possibilities then arise: The mill throughput could be raised,(say doubled),cutting overhead costs.

The project could sustain the capital cost of a power line, leading to lower plant operating costs. Larger equipment could be used in the pit, lowering waste costs.

Almaden requested SRK to consider such a case. The financial indicators are in Table 18-14. While SRK does not recommend it as an initial case, it provides interesting insight as to what the recommended case could become if current gold prices are sustained. This case takes advantage of a higher base gold price to invest more capital and then recoup at an attractive rate over its nine year life, and mine more of the resource. Like the recommended base case, it is not of ultimate economic size, and could be expanded further depending on the gold price outlook 5 to 8 years into the future.


Table 18-14: Pit Shell at $US1200 Au/tr.oz ($US1200 case) Financial Indicators

Financial indicator			Value	Unit
	Pre-tax NPV @	8%	67.9	$CAD M
	Pre -tax IRR		39%
Payback, years from start production			3.30	Years
	Times covered		66.00
	Max exposure		33.53	$CAD M
Tr.oz Au produced			297,239	tr.oz
	Op cost/tr.oz, $		652	$US/tr.oz

Not surprisingly, in the larger case there would be significantly higher capital costs. Of the ore to be mined in this expanded schedule, 7% is based on Measured Resources, 71% on Indicated and 22% on Inferred Resources; so resource risk is increased.

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It is also the case that federal permitting would be required, and this may cause delay. Lastly, the pit is now deep enough to mine high grade material, which may actually be more economically mined by underground methods.

More details of the $US1200 case are in Table 18-15. A financial projection is presented in Appendix 3.


Table 18-15:	Details of $US1200 case
$US1200 case summary			Unit
Assumed gold price		1200	$US/tr.oz
Tonnes per day treated		1000	tpd
Life		9	Years
Total tonnes treated		2.6	Mt
Grade		3.89	g/t
Waste: Ore ratio		30.1
Plant recovery		92	%
Ounces Au produced		297,239	tr.oz
Initial capital expense		17.50	$CADM
Working and preproduction capital		9.60	$CADM
Waste mining		1.90	$CAD /tonne waste
Ore mining		5.87	$CAD /tonne ore
Processing		14,74	$CAD /tonne ore
Administration and overheads		1.27	$CAD /tonne ore
Pre-tax NPV @ 8%		67.9	$CADM
Pre-tax IRR		39%
Max exposure		33.53	$CADM
Payback, years from start production		3.30	years
Ratio, gross earnings: max exposure		6.00
Ratio, NPV: max exposure		2.03

From the expanded pit shell it becomes quick and relatively inexpensive to access the high grade underground resource that is already known, pointing to a further extension of project life.

The $US1200 case pit shell is compared to the base case shell in Figure 18-5.

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Figure 18-5: $US1200 Pit Shell and Base Case pit shell with included mineralization (red)

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19.	Interpretation and Conclusions

The following conclusions have been made as a result of the assessment of the exploration data gathered so far.

19.1 Geology and Resource Estimation

LGGC reviewed the drilling and sampling data from the Elk Deposit to obtain a sufficient level of understanding to assess the Mineral Resource estimate. The geology of the deposit is well understood and the deposit is considered to be an example of a gold-bearing mesothermal quartz and sulphide vein deposit.

Diamond drill holes are the principal source of the geological and grade data at the Elk Deposit and were the only data type used for the grade estimations in the block model. LGGC used the geological data from the drill holes to create a new interpretation of the B and WD Vein Complexes that has supported the updated mineral resource estimate that is the subject of this report.

In order to be satisfied that the project data is appropriate for inclusion in the resource estimation, LGGC reviewed the QA/QC data available for the project assay data and completed a database audit. LGGC found the data and QA/QC support to be reasonable for inclusion into Mineral Resource estimation and has included recommendations to optimise the QA/QC protocols for future sampling programs.

19.2 Mining

Waste mining is the most economically important component of the mining operation but should be straightforward. Considerable work on the pit and dump designs and scheduling is required. Rock slope angles need further study, and improvement is possible. Tests have indicated that most of the mine waste will not be acid generating. Geotechnical considerations of environmental areas may limit the areas that can be used for dumping.

Ore mining will be slow and relatively expensive, but SRK are of the opinion that good mining recovery at moderate dilution is feasible.

19.3 Mineral Processing and Metallurgy

Whole ore cyanidation is technically effective but more expensive and would be more difficult to permit.

In the two laboratory testing programs that have been performed within the last 2 years, the various composites have responded very well to flotation at a coarse grind of > P 80 =200 microns, and to cyanidation after fine grinding to P80 = 40 microns or finer. The flotation stage recovers approximately 95 % of the sulphide sulphur, so the flotation tailing, representing 94 % of the plant feed weight, is depleted in sulphide sulphur. That product is non-PAG and can be discarded into an unlined tailing containment. A circuit that includes gravity and flotation concentration is feasible. This route has therefore been adopted.

Beyond that stage, options include selling the flotation concentrate, or cyanide leaching it on site. For purposes of this report, the concentrate cyanidation option is included. Although that option incurs a higher capital cost and operating cost, it does allow the revenue to be forecast more definitely.

The flotation concentrate is only about 5% by weight of the feed. It will be strongly acid generating. The cyanide leaching residue would have to be stored indefinitely in a lined and flooded containment.

The above scenario will result in 95 % gold recovery into the combined gravity + flotation concentrates, and 98 % recovery in the cyanidation stage, for a combined recovery assumed to be 92 %.

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19.4	Environmental Considerations

Almaden has been continuing to collect a wide variety of environmental data from the site since 1992 and is well prepared to integrate the design of all mining, processing, water supply, drainage, and infrastructure works with each other and with the environment.
Mine waste rock samples have been tested and found to be generally not acid generating, refer to Appendix 4. The dumps will be covered with topsoil and reseeded, as has already been trialled on the trial pit dumps. Environmental baseline of potential dump sites and a waste management plan will be drawn up, to include all the normal activities such as condemnation drilling, stripping topsoil, and reclamation procedures.
A tailings dam site must be geologically examined to eliminate the chance of covering resources of value. Geotechnical foundation testing will also be required. A project water balance must be made, involving the development of an acceptable climatic profile. As a result, the dam can then be sized so as to avoid direct discharge. It is assumed that water will be recycled to the plant. The mine drainage will be discharged to the TSF, thus providing some process plant water, and preventing direct discharge.
The plant flotation tailing has been tested and found to be not acid- generating.
The tailings and surface water management studies will also assist in identifying a make-up water source, and obtaining permission to abstract the quantity required for mining and processing.
Permission to build a power line to the site would be desirable before proceeding with the project.
Permission to upgrade and re- route the access road from the adjacent highway will be required.
The site infrastructure must be sited and waste management planned and permitted.
The First Nation Groups will be consulted on an ongoing basis.

19.5 PEA Conclusion

Overall, SRK believes that this preliminary assessment demonstrates that a viable project could be launched on the Elk property, and therefore, that further work is justified. SRK s favoured strategy is to start production using only the currently-owned processing plant on the lines of this document s Base Case. This minimises capital costs and time for procurement. Once production is established, the ongoing outlook for the gold price will determine when an expansion is carried out.

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20.	Recommendations
The execution plan must, of course, be preceded by a PFS study; which is the next step. Specific recommendations for ongoing work are:

20.1	Geology and Mineral Resources Estimates

Infill drilling in the area of the mineral resources should focus on the areas that are likely to be mined within the base case pit shell. LGGC had recommended a drilling program to increase the drill hole spacing in the area of detailed exploration and current mineral resources at the Elk Deposit. The list of recommended drill hole locations and orientations suggested drill hole sites in three categories of priority; 1, 2 and 3.

A category 1 drill hole was typically suggested as it would increase the drill hole spacing in an area of inferred mineral resources contained within the SRK base case and alternate case ($1200 USD Au/oz) pit shell outlines. LGGC had recommended infill drilling 121 drill holes for approximately 15,500 m of drilling in category 1 priority. Almaden completed 12,749 m of drilling in 2010 in 87 drill holes and an updated mineral resource may still benefit from drilling more drill holes in the northern and peripheral areas of the veins where drill density is not ideal and greater than 25 to 35 m between drill holes. There remain about 3000 m of additional drilling in the first priority recommendations. Drilling costs including sampling and other drilling charges are approximately $100/m so the remaining drilling could cost about $300,000.

LGGC had recommended the drilling of 70 further drill holes (13,000 m) to test the mineralization found below the current pit shells produced from SRK s PEA study. These drill holes are a lower priority than the

Category 1 drill holes but should be considered for possible expansion of the mineral resources. This drilling would cost approximately $1,300,000 CAD to complete.

Condemnation drilling of the potential dump sites, plant site, and tailings dam site, will follow an initial design and schedule. It is difficult to estimate the amount of drilling that is going to be required to condemn mineral potential underlying an area of 1 km². This is the area estimated to be required to store all the waste material from the largest pit size included in the PEA study. A budget of $1.5 million CAD should be adequate to condemnation drill a large enough area to start planning the location of the storage facilities along with other potential infrastructure sites.

More bulk density measurements of vein material and waste rock material are required; and should be regularly captured during the course of all drill programs.

As detailed in Section 11.1 there is a survey discrepancy that needs to be resolved. Both the discrepancies in the collar elevations and the 2 to 8 m discrepancies in Northings and Eastings of 4 of the 25 (16% error rate) historical drill holes warrants a complete resurvey of all historical drill holes (surface and underground) by a registered land surveyor before any further resource estimations or a pre-feasibility study is undertaken on the Elk Project A complete resurvey of the underground drill holes may be too difficult to undertake given that the workings are flooded but the currently accessible survey stations in the decline should be included in the resurvey program and a reliable conversion factor be determined to convert the underground drill holes to the UTM coordinate system. The survey must also reconcile the elevation difference in the two separate datums used, as this will allow the integration of the aerial photography and Lidar planimetric mapping.

Based on the review of the QA/QC data available for the Elk Project, LGGC makes the following recommendations to standardise the QA/QC program:

SRM packages should weigh a minimum of 100 g so the laboratory has sufficient material to complete at least two analyses on the same pulp if re-assaying is required.
A wider range of grade values of SRMs should be purchased. LGGC recommends that SRMs with grade ranges around 1.0 g/t Au, 3.0 g/t Au, 5.0 g/t Au, 10 g/t Au and 30 g/t Au be purchased. Two SRMs should be purchased for each of the grade ranges mentioned above with similar expected grades.

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As an example, if CDN-GS-2E with an expected range of 1.52 g/t is purchased then CDN-GS-2F with an expected range of 2.16 would pair well.

SRM results should be reviewed for compliance immediately upon receipt of the assay data from the laboratory. If two consecutive SRM results fail by 2 standard deviations on the same certificate then the corresponding sample batches should be re-assayed immediately by the lab. If a single SRM result fails by 3 standard deviations then the associated assays in the sample batch should be re-assayed.
Core duplicates should continue to be taken and should be the two halves of the core with a note placed in the sample box as to the nature of the missing sample. Almaden photographs the entire drill core so a photographic record will remain.
Coarse reject duplicates should be taken on a one in 20 basis. This is best accomplished by providing the laboratory with two consecutive sample tags to increase the chance that the samples are analysed together.
Pulp duplicates should be taken on a one in 20 basis so that the sample numbers are consecutive and Almaden should not rely on the duplicate results from the laboratory.
Laboratory check samples sent to a secondary laboratory and batches of blind resubmissions back to the primary laboratory need to include SRM samples with the shipments on a 1 in 20 basis to confirm the reliability of the results.
Screen analysis method should be used for all samples containing quartz vein material, especially the B and WD Vein Complex samples. Samples outside of the resource area should be analysed using fire assay method with AA finish for samples that run less than 10 g/t Au. For samples that return higher than 10 g/t Au, fire assay method with a gravimetric finish should be used.
All vein samples should be bracket sampled for over 3 m above and below the vein to properly sample the hanging-wall and footwall material.
The laboratory duplicate data produced by Acme (Pulp and Coarse Reject Duplicates) should be captured in the project database for the assay results from core drilled between 1989 and 1996.
The Elk samples may benefit by increasing the crush and pulverised sample sizes to 500 g from 250 gm. If Acme has the equipment, then a 1 kg sample size should be used.

20.2 Mining

At the same time as holes are being drilled for resource upgrading, a rock-mechanics drilling Program should be carried out. The current rock mechanics report is useful, but limited to observations in the current pit, which is far from some of the proposed slopes.

More testing of the current old dumps and the waste from drill holes should be undertaken to determine whether there would be acid drainage from dumps. Almaden have already completed an initial test Program, see Appendix 4, indicating that none of the waste is PAG, but these tests may not be exhaustive. The trial mine dumps have apparently caused no problems. A mine waste management plan should be drawn up, to include all the normal activities such as condemnation drilling, stripping topsoil, and reclamation procedures. A surface water management plan will be drawn up. Mine water, mine waste and the TSF require an integrated design approach.

Once Indicated / Measured resources are available, numerous pit/dump design and scheduling options should be studied. Whittle will play an early part in these assessments, but more flexible and intelligent scheduling packages are available and should be used for detailed work.

Careful thought should be given to road design; in most parts of the pit a footwall ramp would be much more economical for stripping; and also longer-lasting, because of the hanging wall push backs that are likely to be a feature of any schedule. The extra cost of these push backs is currently not allowed for in the simple cost model, and will add to waste removal costs unless minimised and carefully designed.

As part of the pit design, the location of connections to future underground workings should be allowed for.

An outline underground study should be made and the economics compared to the open pit to determine where underground working becomes more economical than the open pit.

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The layout, design and scheduling of waste dumps is important, in order to minimise the cost of waste disposal .It should be given careful consideration in the mine scheduling.

The accuracy of the mining contractor costs will be improved once a contractor has been given a realistic plan and schedule to work to.

At the start of development, the old trial pit and underground workings will have to be dewatered. This water must be stored and if necessary released in a controlled manner, so that the local river water quality remains within allowable limits. A water storage facility for this purpose has been designed (Klohn Crippen).

20.3 Mineral Processing and Metallurgy

Metallurgical samples that are more representative of the head grade of early years of the schedule need to be subjected to off-setting process testing, using the optimum process parameters that are reported in G&T report KM2522.

A tailings dam site must be located, condemnation -drilled, surveyed, and tested for permeability. It is also important to locate a plant site location, in conjunction with the overall mine layout.

A project water balance must be made, involving the development of an acceptable climatic profile. As a result, the dam can then be sized so as to avoid discharge. It is assumed that water will be recycled to the plant.

A careful survey of the existing owned equipment should be carried out to allow its condition and suitability to be determined, and a plant layout and flow sheet should be made.

A definitive capital cost estimate covering all aspects of plant construction must be made.

20.4	Infrastructure
20.4.1	Overall Project Layout

The locations of the various surface facilities that are described in this report have not been subjected to detailed studies. Almaden has just commissioned initial work on this. It seems clear that there is overall ample space for the required facilities; But site soil and foundation studies must be made to determine the optimum locations for waste rock and tailing disposal, plant site location, and other site infrastructure.

20.4.2 Water Supply

The project currently has no identified source of water supply for the required estimated 170 cu m / day of make-up water that will be required for the processing plant at a production rate of 500 tpd. Sources of water will need to be identified in conjunction with other site water management issues, prior to the preparation of an Environmental Impact Statement (EIS).

20.4.3 Electrical Power

At this stage of the project it is planned to supply electrical power at the site with diesel-electric generators for the Base Case. That practice can continue if the plant is expanded, but at larger project sizes the use of grid power may be advantageous. Discussions with BC Hydro should be initiated to determine the feasibility and cost to obtain 3 phase grid power.

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20.4.4 Environmental Matters

Baseline studies on flora, fauna, air quality, and noise must be completed.

Three separate First Nation groups are required to be consulted and potentially accommodated for any work carried out on the property.

Permission to build a power line to the site will be required.

Permission to upgrade the access road from the adjacent highway will be required.

The various mining and metallurgical studies outlined above, have to be completed before an Environmental Impact Statement and Management Plan can be prepared.

20.5 Cost of Recommended Work Program

Specific recommendations for ongoing work are included throughout this report and summarized in Table 20-1. This table provides cost estimates for the more important study work required to advance the project so that a PFS can be completed and permitting requirements satisfied.

The total estimated costs of all proposed study work is $CAD5.44M, which does not include Almaden s overhead costs in the period covered. Seasonal conditions and permitting uncertainties make it difficult to estimate the time required for this Program but SRK estimates it could be in the order of 18 months to 2 years.

Table 20-1 should not be construed as being in time order. The whole study process is iterative and the activities are interdependent, and the scope of each step is usually dependent on the outcome of previous steps.

Table 20-1: Cost of PFS Work Program



Work Item	Cost
	K$CAD
Design a preliminary project layout	15
Complete Environmental baseline studies	150
Consult with first Nation Groups	70
Apply for permission to reroute and upgrade access road	25
Obtain Permission for powerline to site	50
Initiate a drilling Program to:-
Locate ,quantify and permit project water supply	100
Determine slope angles in the proposed pit, define vein strength and
friability and separation from walls	250
Convert Inferred to Indicated Resources by Infill Drilling	1600
Obtain further waste bulk densities and conduct further ABA testing	20
Obtain samples of early production for metallurgical testing	150
Condemnation drill Program for waste dump, TSF, plant site	1500
Update Mineral Resources estimate	150
Make early-years confirmatory metallurgical tests. Update plant	30
parameters, design flow sheet
Survey existing processing equipment, make plant layout and cost	300
estimate for plant
Complete surface rainfall / runoff study	25
Design tailings disposal facility, estimate cost	50

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	Work Item	Cost
		K$CAD
	Obtain initial mining costs	30
Reoptimise pit and schedule pit and dumps, study the integration of
underground mining into the overall mining plan, adjust dumps and		370
	surface layout
Obtain final costs iterate planning		150
Design and estimate site infrastructure		100
Investigate marketing of doré vs. flotation concentrate and estimate costs		50
	Document PFS	250
	Total Costs K$CAD	5435

The PFS has to be finished in order for an EI Report to be documented, but the work to prepare this can be started towards the end of the PFS and the EI report is expected to be produced 3-6 months after the PFS is complete. The EIS may cost K$CAD250.

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21.	References

Almaden Minerals, 2010. Siwash Gold Deposit Elk Gold Project Reclamation Report 2009, prepared for the Ministry of Mines, Similkameen Division.

Almaden Minerals Ltd. 2010. Siwash Gold deposit Elk Gold Project Reclamation Report 2009, Permit M-199 and Amended Mines Act Permit MX-4-387.

Bacon and Donaldson, 1992. Preliminary Metallurgical Testing of Siwash North Gold Deposit, for Cordilleran Engineering Ltd, Vancouver, BC, April 6, 1992.

Bacon Donaldson, 1992b, as outlined in Fairfield Minerals (1993) Siwash North Gold deposit – Compilation of Metallurgical Results, May 15, 1993.

Blythe, M, 2007. Internal Scoping Study Elk Gold Mine, unpublished report, October 2007.

Brenda Process Technology Group, 1992. Letter report on Metallurgical Testing, for Fairfield Minerals Ltd, August 24, 1992.

Conroy, P W, 1994. 1993 Underground Mapping and Mining Report, Siwash North Gold Mine, Elk Property, March 1994.

Cordilleran Engineering Ltd, 1992. Siwash North Gold Deposit, Reverse Circulation Drilling Report - 1992.

G&T Metallurgical Services Ltd, 2006. Report KM1903: Review of Historical Data – Almaden Minerals – Elk Project, September 18, 2006.

G&T Metallurgical Services Ltd, 2008. Report KM2121: Preliminary Metallurgical Evaluation of Samples from the Elk Property, April 29, 2008.

G&T Metallurgical Services Ltd, 2010. Report KM2522: Metallurgical Evaluation of Composites from the Elk Deposit, May 18, 2010.

Geiger, A K, 2000. A Fluid Inclusion Study and Geostatistical Analysis of Vein mineralization in the Siwash North Study Area, South-Central British Columbia, April 2000.

Giroux, G H, 2005. 2004 Update of Resource, Siwash Project, Elk Property for Almaden Minerals Ltd, amended May 28, 2005.

Giroux, G, 2007. 2007 Update of Resource Siwash Project, Elk Property for Almaden Minerals Ltd, November 30, 2007.

Giroux, G, undated. Unpublished documentation in the form of an updated resource and 3D block model for the entire deposit.

Golder Associates, undated. Unpublished report analysing the environmental and First Nations issues which must be addressed to allow the project to progress to the permitting phase.

Golder, 2006. Elk Gold Mine Baseline Fall Fisheries Site Report, technical memorandum dated October 25, 2006.

Golder, 2007. Report on Review and Data Gap Analysis of Environmental Baseline Studies, Elk (Siwash) Gold Mine, prepared for Almaden Minerals.

Golder Associates Ltd, 2008a. Water quality and Benthic Invertebrate Baseline Study, Elk Gold Mine, Siwash Lake, BC, prepared for Almaden Minerals Ltd, July 4, 2008.

Golder Associates Ltd. 2008b. Final Report on Annual Reclamation Report for 2007, Almaden Minerals, Siwash Gold Deposit – Elk Gold Project Permit M-199 and Amended Mines Act Permit MX-4-387.

Prepared for Almaden Minerals Ltd.

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Hawthorn, G, 2003.	Willow Creek Gold Mill, memorandum dated 9 June 2003.
Hylands, J. J, 2007.	2007 Diamond Drilling, Siwash Gold Mine Area, Elk Property, October 2007.

International Corona Corporation, 1992. Letter report on Metallurgical Testing, for Fairfield Minerals Ltd, October 8, 1992.

Jabukowski, V, 2004. Siwash Gold Deposit 2004 Reclamation Report, prepared by Almaden Minerals for the Similkameen Mining Division, British Columbia.

Jakubowski, W J, 2003. 2002 Diamond Drilling Report, Siwash Gold Mine Area, Elk Property, May 2003. Jakubowski, W J, 2004. 2003 Diamond Drilling Report, Siwash Gold Mine Area, Elk Property, May 2004. Jakubowski, W J, 2004. 2004 Diamond Drilling Report, Siwash Gold Mine Area, Elk Property, May 2004. Jakubowski, W J, 2006. 2005 Diamond Drilling, Siwash Gold Mine Area, Elk Property, March 2006. Jakubowski, W J, 2007. 2006 Diamond Drilling, Siwash Gold Mine Area, Elk Property, March 2007.

Kennard, D and Chance, A, 2007. Preliminary pit slope angles for use in pre-feasibility studies of future expansion of the Siwash open pit mine on the Elk Property, Merritt, BC, technical memorandum for Almaden Minerals prepared by Golder Associates, March 29, 2007.

King, H. L., 2001. Geological Report on Elk Property, Similkameen Mining Division for Fairfield Minerals Ltd, August 2001.

Klohn Crippen Berger, 2010. Elk Gold Project – Water Storage Dam Design, prepared for Almaden Minerals Ltd, March 10, 2010.

Knight Piesold Ltd. 2004. Environmental Baseline Studies (Ref: VA102-147/02-1), December 2004.

LaPlante, A, 1993. Report on the Treatment of a Fairfield Sample by Gravity and Flotation, November 20, 1993.

Lewis, P D, 2000. Structural Analysis of the Siwash Mine and Exploration Implications for the Elk Property, December 2000.

Rousseau, 1992. An archaeological Resource Overview Assessment for the Proposed Elk-Siwash Gold Deposit Project Between aspen Grove and Kelowna South Central BC Prepared by Antiquus Archaeological Consultants Ltd. 1992.

Sprott, D (Golder Associates), Blythe, M (Almaden Minerals Ltd) and Allen, J (Golder Associates), undated, unpublished documentation relating to the potential for both open-pit and underground mining.

Taylor, M E and McKee, P, 1999. Wild Ruminant Study at Brenda Mine, prepared by Stantec Consulting Ltd.

Wildstone Resources, 1993. Description of Wildlife Resources on and adjacent to the Siwash North Gold Property, prepared for Fairfield Minerals Ltd and Cordillearan Engineering Ltd, October 28, 1993.

www.	westbankchamber.com and www.merritt-

chamber.comwww.srmwww.gov.bc.ca/cdc/gis/eo_data_fields.htm

http://www.bclaws.ca/EPLibraries/bclaws_new/document/ID/freeside/13_370_2002

Mineral mine” as defined by the Mineral Tenure Act http://www.bclaws.ca/EPLibraries/bclaws_new/document/ID/freeside/00_96292_01

http://laws.justice.gc.ca/eng/F-14/page-5.html#anchorbo-ga:s_34

http://laws.justice.gc.ca/en/N-22/

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http://laws.justice.gc.ca/eng/E-17/page-4.html#anchorbo-ga:s_7

http://laws.justice.gc.ca/en/c-15.2/text.html

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22.	Date and Signature Page
Effective Date of Report: 10 December 2010

R H Pooley BSc, MAusIMM

Senior Mining Consultant, Qualified Person SRK Consulting (Australasia)Pty Ltd

“Signed and Sealed”

S Lomas P.Geo (BC)

Principal Consultant, Qualified Person Lions Gate Geological Consulting Inc.

“Signed and Sealed”

Gary Hawthorn, P Eng.

Metallurgical Consultant, Qualified Person Westcoast Mineral Testing Inc.

“Signed and Sealed”

R. Brian Alexander, P.Geo (BC, Nunavut) Principal Consultant, Alexploration Inc.

14 January 2011

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Level 1
10 Richardson Street
West Perth WA 6005, Australia

PO Box 943
West Perth WA 6872, Australia

Email: perth@srk.com.au
www.srk.com.au

Tel: + 61 (0) 8 9288 2000
Fax: + 61 (0) 8 9288 2001

CERTIFICATE OF QUALIFIED PERSON

Roger Pooley, MAusIMM, CPE Mining

I, Roger Harry Pooley, Level 1, 10 Richardson Street, West Perth, 6005, Western Australia, Australia, do hereby certify that:

1.	I am a Senior Mining Consultant with SRK Consulting (Australasia) Pty Ltd, Level 1, 10 Richardson Street, West Perth 6005, Western Australia, Australia;
2.	I graduated with a BSc degree in Mining Engineering from Royal School of Mines, Imperial College, University of London in 1959, also gaining a RSM diploma concurrently;
3.	I am a Member of The Australasian Institute of Mining and Metallurgy (AusIMM), and a CP in Mining under the AusIMM s criteria;
4.	I have worked as a Mining Engineer for a total of 51 years since graduation from university;
5.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be an independent qualified person for the purposes of NI 43-101;
6.	I am co-responsible for Sections 1 (1.1, 1.1.6 and 1.1.7), 18 (18.1, 18.2, 18.4, 18.5, 18.6, 18.6.1, 18.6.3, 18.7, 18.8, 18.9, 18.10, 18.11 and 18.12), 19 (19.2, 19.4 and 19.5) and 20 (20.2, 20.4 and 20.5) and solely responsible for Sections 2, 3, 4, 5, 21, 22 and 23, part of Table 20-1, Appendices 1-4, and for overall compilation of the Technical Report, Elk Gold Project, Preliminary Economic Assessment, NI 43-101 Technical Report, dated 14 January 2011 (the “Technical Report”).
7.	I have not had prior involvement with the property that is the subject of the Technical Report;
8.	I am independent of the issuer applying all of the tests in Section 1.4 of National Instrument 43-101.
9.	As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Signed Perth, Western Australia, this 18^thday of January 2011.

Roger Pooley


SRK Consulting (Australasia) Pty Ltd	Group Offices:	Australian Offices:
Reg’d No ABN 56 074 271 720	Africa	Brisbane	+ 61 (0) 7 3832 9999
Trading as SRK Consulting	Asia	Melbourne	+ 61 (0) 3 8677 1900
	Australia	Newcastle	+ 61 (0) 2 4922 2100
	Europe	Perth	+ 61 (0) 8 9288 2000
	North America	Sydney	+ 61 (0) 2 8079 1200
	South America

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CERTIFICATE OF QUALIFIED PERSON

Gary Hawthorn, P.Eng. (BC)

I, Gary William Hawthorn, of 2806 Thorncliffe Drive, North Vancouver, BC, do hereby certify that:-

1.	I am a Professional Mineral Processing Engineer and employed as a Metallurgical Consultant with Westcoast Mineral Testing Inc.;
2.	I graduated with a BSc degree in Mining Engineering from Queen s University, Kingston Ontario, in
	1964;
3.	I am a registered professional engineer in the Province of British Columbia, since 1972;
4.	I am an independent qualified person;
5.	Since 2003 I have intermittently provided Consulting Engineering services to Almaden Minerals Limited to investigate and review the mineral processing characteristics of the ELK deposit;
6.	I have worked as a mineral processing engineer for 46 years and as a professional engineer for 38 years;
7.	My operating experience includes 18 years, mainly as a mill superintendent for Cominco Ltd and Placer Development Ltd;
8.	Starting in 1982 I have been self employed as a consulting Mineral Processing Engineer, and since 1988 have operated as Westcoast Mineral Testing Inc.;
9.	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 fulfil the requirements to be an independent qualified person for the purposes of NI 43-101;
10.	I am co-responsible for Sections 18 (18.3 and 18.6.2), 19 (19.3) and 20 (20.3) and solely responsible for Section 16 of the Technical Report, Elk Gold Project, Preliminary Economic Assessment, NI 43-101 Technical Report, dated 14 January 2011.
11.	I have not visited the subject property;
12.	I am not aware of any material fact or material change with respect to the subject matter of the Technical Report that is not reflected in the Technical report, the omission to disclose which makes the Technical Report misleading;
13.	I have read NI 43-101, Form 43-101F1, and 43-101CP. The Technical report has been prepared in compliance with those regulations.

Signed Vancouver, BC, this 14^thday of January 2011.

_____________________________________

Gary Hawthorn

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CERTIFICATE OF QUALIFIED PERSON

R. BRIAN ALEXANDER, P.GEO.
Alexploration Inc.
503-2759 Carousel Crescent
Ottawa, Ontario, Canada, K1T 2N5

I, Robert Brian Alexander am a Professional Geoscientist and Principal Consultant for Alexploration Inc.

This certificate applies to the technical report titled NI43-101 Technical Report for a Preliminary Economic Assessment on the Elk Gold Project, Merrit, British Columbia, Canada, dated January 14, 2011 (the Technical Report).

I graduated from the University of New Brunswick in Fredericton, NB, Canada, in 1979 with a B.Sc. in Geological Sciences and I have practiced my profession continuously since then for a total of 31 years in Canada and internationally. I worked exclusively on gold exploration projects in the Abitibi area of Ontario and in the Woodburn Group in Nunavut, Canada during 1985-1993, and 1997-2003. In the period of 1999 to 2004, I provided QP services on gold projects in Nunavut.

I am a Practicing Member of the Association of Professional Engineers, Geologists and Geophysicists of the Northwest Territories and Nunavut (Registration #L1093). I am also a Practicing Member of the Association of Professional Engineers and Geoscientists of the Province of British Columbia (Registration #34713). I am also a member of the Prospectors and Developers Association of Canada.

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

I am independent of the issuer applying all of the tests in section 1.5 of National Instrument 43-101.

I was on the project site and supervised the 2010 Drill Program for Almaden Minerals Ltd, Elk Project in southeastern BC during the period July to November, 2010, during which I reviewed data, drill site locations, and drill core.

I am responsible for section 1.1.4.1, 11.1.1 and 13.1.3 and of the Technical Report.

I am independent of Almaden Minerals as independence is described by Section 1.4 of NI 43-101 and have no prior involvement or interest in the Elk Project prior to commencing the 2010 drill program.

I have read National Instrument 43-101 and this Technical Report has been prepared in compliance with that Instrument.

As of the date of the certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

“Signed and Sealed”

Robert Brian Alexander, P.Geo.
18 January, 2011

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23.	Additional Requirements for Technical Reports on Development and Production Properties

The Elk Gold Property is not currently either under development, or in production, so this section is not applicable.

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Appendix 1: Base Case Financial Projection (Gold at $US1000/tr.oz)

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Appendix 2: Base Case Financial Projection (Gold at $US1200/tr.oz)

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Appendix 3:	$US1200 Case Financial Projection
(Gold at $US1200/tr.oz)

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Appendix 4: Summary of ABA testing on Mine Waste Rock

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Acid Mine Drainage

To investigate the potential for acid mine drainage (AMD) from the site, samples of waste rock from diamond drill core have been assayed using the acid- base accounting (ABA) procedure.

In performing an ABA analysis, data is included to identify both the condition of the submitted sample and to determine the potential for contained sulphide sulphur to produce sulphuric acid. The sulphur assay, as % sulphur, and the calculated acid potential (AP) from sulphide sulphur, expressed as kg/ tonne, are reported.

The analyses also includes a measurement of the contained carbonate to determine the extent to which any generated acid will be consumed within the rock mass on disposal as either waste rock or tailing. The carbonate content is expressed as neutralization potential (NP) also as kg / tonne.

From a chemical perspective, if the AP equals the NP, acid will either not be produced or will not escape the disposal site since it will be neutralized by the contained carbonate. In practice, to provide greater assurance that that acid generation will not occur, disposal rock is typically considered to be not potentially acid generating (non-PAG) if the ratio of NP to AP (expressed as neutralization potential ratio, NPR) exceeds 2.

Any material in which the NPR is less than 1 is PAG.

The NPR in the range of 1 – 2 is neither non-PAG nor PAG, so the data needs to be reviewed on a case by case basis to determine its impact upon the overall site water rock disposal.

In the case of the ELK project, all of the submitted samples of mine waste have NPR of over 2, and therefore are non-PAG. Because the current testing is not exhaustive (see table and map below), the report does address the possibility that maybe 5% of the waste would be PAG, which is easily managed during disposal. Further testing is recommended as part of the next phase of development.

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Waste Rock Test Results

	OA-	OA-		OA-	OA-	OA-		S-	S-	S-	C-	C-
			OA-VOL08				S-IR08
Sample	VOL08	VOL08		VOL08	ELE07	VOL08		GRA06	GRA06a	CAL06	GAS05	GAS05
description			Ratio	FIZZ		Paste
	AP	NP			NNP		S	S	S	S	C	CO2
			(NP:MPA)	RATING		pH
	kg/t	kg/t	Unity	Unity	kg/t		%	%	%	%	%	%
ARD10-1	0.6	14	22	2	13	9.0	0.02	<0.01	0.03	0.02	0.13	0.5
ARD10-2	0.6	23	37	2	22	9.0	0.02	<0.01	0.02	0.02	0.17	0.6
ARD10-3	0.3	5	16	2	5	8.8	0.01	<0.01	0.01	0.01	0.06	0.2
ARD10-4	0.6	25	40	2	24	8.8	0.02	0.01	0.01	0.01	0.25	0.9
ARD10-5	0.3	16	51	2	16	9.2	0.01	<0.01	0.02	0.01	0.13	0.5
ARD10-6	0.3	15	48	2	15	8.8	0.01	<0.01	0.01	0.01	0.14	0.5
ARD10-7	0.3	21	67	2	21	9.4	0.01	<0.01	<0.01	0.01	0.16	0.6
ARD10-8	0.3	8	26	1	8	9.0	0.01	<0.01	0.01	0.01	0.08	0.3
ARD10-9	0.3	7	22	1	7	8.9	0.01	<0.01	0.02	0.01	0.1	0.4
ARD10-10	0.3	6	19	2	6	8.7	0.01	<0.01	0.03	0.01	0.09	0.3
ARD10-11	<0.3	18	115	2	18	9.0	<0.01	<0.01	0.02	<0.01	0.13	0.5
ARD10-12	0.3	25	80	1	25	9.4	0.01	<0.01	0.01	0.01	0.12	0.4
ARD10-13	0.9	19	20	2	18	9.0	0.03	<0.01	<0.01	0.03	0.21	0.8
ARD10-14	1.6	10	6	2	8	8.5	0.05	<0.01	0.01	0.05	0.21	0.8
ARD10-15	3.1	26	8	1	23	9.4	0.1	<0.01	0.02	0.1	0.08	0.3

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Sample Location Plan

Waste rock ABA test locations

Sample #	Hole #	From	To
		(m)	(m)
ARD10-001	SND10-075	33.0	34.0
ARD10-002	SND10-060	33.0	34.0
ARD10-003	SND10-010	72.5	73.5
ARD10-004	SND10-065	35.5	36.5
ARD10-005	SND10-068	37.1	38.1
ARD10-006	SND10-079	52.5	53.5
ARD10-007	SND10-069	48.0	49.0
ARD10-008	SND10-081	46.4	47.4
ARD10-009	SND10-087	95.3	96.3
ARD10-010	SND10-086	34.2	35.2
ARD10-011	SND10-083	37.3	38.3
ARD10-012	SND10-059	113.3	114.3
ARD10-013	SND10-072	186.3	187.3
ARD10-014	SND10-072	239.5	240.5
ARD10-015	SND10-059	169.8	170.8

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