-----BEGIN PRIVACY-ENHANCED MESSAGE----- Proc-Type: 2001,MIC-CLEAR Originator-Name: webmaster@www.sec.gov Originator-Key-Asymmetric: MFgwCgYEVQgBAQICAf8DSgAwRwJAW2sNKK9AVtBzYZmr6aGjlWyK3XmZv3dTINen TWSM7vrzLADbmYQaionwg5sDW3P6oaM5D3tdezXMm7z1T+B+twIDAQAB MIC-Info: RSA-MD5,RSA, HsIqK/XFtk0HbAKg6+B590KjN9s4mLu9+MUdsh561e7DCz3QZqULHS2992/DI1mr OK2dHnjl3bASBeOsRmnMhw== 0001378296-10-000063.txt : 20100427 0001378296-10-000063.hdr.sgml : 20100427 20100426180527 ACCESSION NUMBER: 0001378296-10-000063 CONFORMED SUBMISSION TYPE: 6-K PUBLIC DOCUMENT COUNT: 56 CONFORMED PERIOD OF REPORT: 20100426 FILED AS OF DATE: 20100427 DATE AS OF CHANGE: 20100426 FILER: COMPANY DATA: COMPANY CONFORMED NAME: INTERNATIONAL TOWER HILL MINES LTD CENTRAL INDEX KEY: 0001134115 STANDARD INDUSTRIAL CLASSIFICATION: METAL MINING [1000] IRS NUMBER: 000000000 STATE OF INCORPORATION: A1 FISCAL YEAR END: 0531 FILING VALUES: FORM TYPE: 6-K SEC ACT: 1934 Act SEC FILE NUMBER: 001-33638 FILM NUMBER: 10771288 BUSINESS ADDRESS: STREET 1: 1188 WEST GEORGIA STREET STREET 2: SUITE 1920 CITY: VANCOUVER STATE: A1 ZIP: V6E 4A2 BUSINESS PHONE: 604-683-6332 MAIL ADDRESS: STREET 1: 1188 WEST GEORGIA STREET STREET 2: SUITE 1920 CITY: VANCOUVER STATE: A1 ZIP: V6E 4A2 6-K 1 form6k.htm FORM 6K OF THE REGISTRANT DATED APRIL 26, 2010 Form 6-k

FORM 6K


UNITED STATES

SECURITIES AND EXCHANGE COMMISSION

Washington, D.C. 20549


Report of Foreign Issuer


Pursuant to Rule 13a-16 or 16d-16 of

the Securities Exchange Act of 1934


For the month of April 2010


Commission File Number 000-31096



International Tower Hill Mines.

(translation of registrant’s name into English)


#1920 - 1188 West Georgia Street

Vancouver, British Columbia V6E 4A2

(Address of principal executive offices)


Indicate by check mark whether the registrant files or will file annual reports under cover Form 20-F or Form 40-F.


Form 20-F x

Form 40-F ¨



Indicate by check mark if the registrant is submitting the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(1): ¨


Note:

Regulation S-T Rule 101(b)(1) only permits the submission in paper of a Form 6-K  if submitted solely to provide an attached annual report to security holders:


Indicate by check mark if the registrant is submitted the Form 6-K in paper as permitted by Regulation S-T Rule 101(b)(7): ¨


Note:

Regulation S-T Rule 101(b)(7) only permits the submission in paper of a Form 6-K if submitted to furnish a report on other document that the registrant foreign private issuer must furnish and make public under the laws of the jurisdiction in which the registrant is incorporated, domiciled or legally organized (the registrant’s “home country”), or under the rules of the home country exchange on which the registrant’s securities are traded, as long as the report or other document is not a press release, is not required to be and has not been distributed to the registrant’s security holders, and, if discussing a material event, has already been the subject of a Form 6-K submission or other Commission filing on EDGAR.


Indicate by check mark whether by furnishing the information contained in the Form, the registrant is also thereby furnishing the information to the Commission pursuant to rule 12g3-2(b) under the Securities Exchange Act of 1934.


Yes ¨

No x


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Exhibits

1. Summary Report on the Livengood Project, Tolovana District, Alaska
2. Consent of Tim Carew
3. Certificate of Timothy J. Carew
4. Consent of Paul D. Klipfel
5. Certificate of Paul D. Klipfel
6. Consent of William J. Pennstrom, JR.
7. Certificate of William J. Pennstrom, JR.



 





SIGNATURE


Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.


International Tower Hill Mines.


 


/s/ Jeffrey Pontius

Date: April 26, 2010

Name:

Jeffrey Pontius

Title:

President and CEO




  




In connection with the Company’s listing on the American Stock Exchange, LLC, the Company prepared its U.S. GAAP Balance Sheet as at August 3, 2007.



EX-1 3 exhibit1.htm 43-101 TECHNICAL REPORT OF THE REGISTRANT ENTITLED SUMMARY REPORT ON THE LIVENGOOD PROJECT, TOLOVANA DISTRICT, ALASKA Summary Report on the Livengood Project, Tolovana District, Alaska

 


[exhibit1002.jpg]

[exhibit1003.jpg]





MARCH 2010

SUMMARY REPORT ON THE

LIVENGOOD PROJECT,

TOLOVANA DISTRICT,

ALASKA



Prepared by:


Paul Klipfel  Ph.D.  CPG #10821

   Economic Geologist


Tim Carew  P.Geo.

   Mining Engineer


William Pennstrom Jr.  QP-MMSA

   Metallurgical Engineer


For:


International Tower Hill Mines Ltd.

March 16, 2010


AUTHORS:


Paul D. Klipfel Ph.D.   CPG #10821

Mineral Resource Services Inc.


4889 Sierra Pine Dr.

Reno, NV 89519

(775) 742-2237

p.klipfel@sbcglobal.net


Timothy J. Carew P.Geo.

Reserva International LLC

P.O. Box 19848

Reno, NV 89511 USA

(775) 853-2227

timc@reservainternational.com


William Pennstrom, Jr. QP-MMSA

Pennstrom Consulting Inc.

2728 Southshire Rd.

Highlands Ranch, CO 80126

(303) 748-5174

Bill@Pennstrom.com




 




TABLE OF CONTENTS


Section

 



1.0

SUMMARY

 

1.1

Summary

1.2

Description and Location

1.3

History

1.4

Geology and Mineralization

1.5

Exploration Drilling and Sampling

1.6

Quality Assurance, Quality Control, and Data Verification

1.7

Mineral Processing and Metallurgical Testing

1.8

Resource Estimation

1.9

Conclusions

1.10

Recommendations



2.0

INTRODUCTION AND TERMS OF REFERENCE

 

2.1

Introduction

2.2

Terms of Reference

2.3

Glossary of Key Abbreviations

2.4

Purpose of Report

2.5

Sources of Information

2.6

Field Examination


3.0

RELIANCE ON OTHER EXPERTS

 


4.0

PROPERTY DESCRIPTION AND LOCATION

 

4.1

Area and Location

4.2

Claims and Agreements

4.3

Environmental Requirements

4.4

Permits


5.0

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,

INFRASTRUCTURE AND PHYSIOGRAPHY

 

5.1

Access

5.2

Climate

5.3

Local Resources

5.4

Infrastructure and Physiography


6.0

HISTORY


7.0

GEOLOGICAL SETTING

 

7.1

Regional Geology

7.2

Local Geology

7.3

Geological Interpretation


8.0

DEPOSIT TYPES

 


9.0

MINERALIZATION

 

9.1

Mineralization

9.2

Alteration

9.3

Synthesis of Mineralization and Alteration


10.0

EXPLORATION

 

10.1

Past Exploration

10.2

Current Exploration


11.0

DRILLING

 

11.1

Past Drilling

11.2

Current Drilling

11.3

Drill Procedures


12.0

SAMPLING METHOD AND APPROACH

 

12.1

Past Sampling

12.2

Current Sampling


13.0

SAMPLE PREPARATION, ANALYSES AND SECURITY

 

13.1

Past Procedures

13.2

Current Procedures

13.3

Data Handling


14.0

DATA VERIFICATION

 


15.0

ADJACENT PROPERTIES

 


16.0

MINERAL PROCESSING AND METALLURGICAL TESTING

 

16.1

Introduction

16.2

Metallurgical Summary

16.3

Gold Characterization

16.4

Historical Test Work Programs

16.5

Current Test Work Program

16.5.1

Grind Studies and Ball Mill Bond Work Indices Tests

16.5.2

Gravity Centrifugal Concentration Evaluation

16.5.3

Bottle Roll Leach Tests

16.5.4

Bottle Roll CIL Tests

16.6

Future Metallurgical Test Work

16.7

Mineral Processing

16.8

Gold Recovery

16.9

Process Operating Costs

16.10

Process Capital Cost

16.11

Mining Capital Estimate

16.12

Sustaining Capital


17.0

MINERAL RESOURCE ESTIMATE

 

17.1

Data Used

17.1.1

Sample Data

17.1.2

Other Data

17.2

Data Analysis

17.3

Geologic Model

17.4

Composite Statistics

17.4.1

Gold Indicator Statistics

17.4.2

Contact Analysis

17.5

Spatial Statistics

17.5.1

Gold indicator Variograms

17.5.2

Oxide Indicator Variograms

17.5.3

KINT Dike Variograms

17.5.4

Amy Sequence, Lower Sands and Shale Variograms

17.6

Resource Model

17.6.1

Model Extents

17.6.2

Gold Estimation

17.6.3

Oxidation Estimation

17.6.4

KINT Dike Estimation

17.6.5

Amy Sequence, Lower Sands and Shale Estimation

17.7

Model Validation

17.7.1

Global Bias Check

17.7.2

Visual Validation

17.8

Post Processing of MIK Model

17.8.1

Change of Support

17.8.2

Calculation of Metallurgical Recoveries

17.9

Resource Classification

17.10

Economic Considerations

17.10.1

 Metallurgy

17.10.2

 Mining

17.10.3

 General Overhead

17.11

Whittle Pit Optimization and Analysis

17.11.1

 Heap Leach Option

17.12

Discussion

18.0

OTHER RELEVANT DATA AND INFORMATION

 


19.0

INTERPRETATION AND CONCLUSIONS

 


20.0

RECOMMENDATIONS

 

20.1

Recommended Exploration Program

20.2

Budget for 2010


21.0

REFERENCES

 


22.0

DATE AND SIGNATURE PAGE

 


23.0

CERTIFICATES OF AUTHORS

 


24.0

APPENDICES

 



 

LIST OF FIGURES


Figure

 


Figure 4.1  Location map showing the location of the Livengood project.

 


Figure 4.2  Claim map showing land status of the Livengood land position.

 


Figure 4.3  Map of the Money Knob area showing “wetlands” in green

        pattern, initial and expanded archaeological study area.

 


Figure 5.1  Panorama of Money Knob and the project area.

 


Figure 7.1  Terrane map of Alaska showing the location of the Livengood

       Terrane.

 


Figure 7.2  Geologic cross section and map of the Livengood project area.

 


Figure 7.3.  Diagrammatic lithologic column shows the tectonic stacking of

        rock groups in the Livengood area.

 


Figure 7.4.  Generalized geologic map of the Money Knob area based on

        geologic work by ITH.

 


Figure 7.5  Photographs of key rock types at Livengood.

 


Figure 7.6  Photographs of key rock types and mineralization features.

 


Figure 7.7  This cartoon shows an interpretive sequence of N-S sections

        and events to explain the structural relations observed at the

        surface and in drill core

 


Figure 7.8  N-S Section 428625 E illustrates the complexities of

       thrust and normal fault interpretation and shows the

       southerly dip of high grade zones (red).

 


Figure 7.9  N-S Section 428850 illustrates the southerly dip of high

       grade zone (red) along the general stratigraphic pattern.  

 


Figure 7.10 N-S Section 428925 illustrates the general southerly dip of

        mineralization and how it lies along the stratigraphic and

        structural grain.

 


Figure 7.11  N-S Section 429075 illustrates the pattern of mineralization

        reflecting structural and stratigraphic controls.  

 


Figure 7.12 N-S Section 429675 illustrates the pattern of mineralization

        reflecting structural and stratigraphic controls.  

 


Figure 9.1.  Plot of soil samples.

 


Figure 9.2.  Photomicrographs of characteristic alteration among rocks at

        Money Knob.

 


Figure 9.3.  Photomicrographs of characteristic alteration among rocks

        of Money Knob

 

 


Figure 9.4  Interpreted paragenetic sequence of key alteration and

       mineralization stages.  

 


Figure 11.1  Distribution of drilling in the Money Knob area with

          respect to anomalous soil samples.  

 


Figure 11.2  Distribution of drilling in the Money Knob area according

         to year and company.

 


Figure 11.3  Photos of various exploration functions.  

 


Figure 13.1.  These scattergram plots show how different categories

          of sample duplicates compare with original sample results.

 


Figure 13.2.  X-Y scattergrams for 2009 showing the stated value of

          standards vs. the measured value by the lab.

 


Figure 13.3.  X-Y scattergrams for 2009 showing the stated value of

          standards vs. the measured value by the lab.

 


Figure 14.1  X-Y scatter plots of original and duplicate sample data for

         check samples collected by Dr. Klipfel as part of data

         validation procedures.

 


Figure 14.2  X-Y scatter plot of original and check samples for

         June 2009 RC drilling.

 


Figure 16.1 Flow chart and breakdown of Livengood composite sample

        testwork.

.

 


Figure 16.2 Proposed Livengood process block flow diagram showing

         heap leach process streams.

 


Figure 16.3  Alternate Livengood process block flow diagram showing

          both heap leach and mill/CIL process streams.

 


Figure 17.  Gold distribution by lithology unit.

 


Figure 17.  Contact plots.  

 


Figure 17.3  Swath plots of E-type estimate vs. nearest neighbour.

 


Figure 17.4.  Distance to the nearest composite vs. kriging variance.

 


Figure 19.1  Block model for section 428625E.

 


Figure 19.2  Block model for section 428850E.

 


Figure 19.3  Block model for section 428925E.

 


Figure 19.4  Block model for section 429075E.

 


Figure 19.5  Block model for section 429525E.

 


Figure 19.6  Block model for section 429675E.

 



LIST OF TABLES


Table


Table 4.1  Summary of Claim Holdings and Annual Obligations

 


Table 6.1  Exploration History

 


Table 11.1  Summary of AGA and ITH drilling at Livengood

 


Table 16.1 Gold recovery from 2007 cyanide extraction tests

 


Table 16.2  Gold recovery from 2008 Hazen cyanide extraction

tests (-10 mesh)

 


Table 16.3  Gold recovery results from Kappes Cassiday cyanide

        extraction tests (-200mesh)

 


Table 16.4 Cyanide shake leach test procedure parameters

 


Table 16.5  Livengood project – Main Zone summary of cyanide

        shake tests (5 GPL NACN)

 


Table 16.6  Livengood project – Sunshine Zone summary of cyanide

        shake tests (5 GPL NACN)

 


Table 16.7  Livengood project – Main Zone Bond Ball Mill work

        index test results

 


Table 16.8  Livengood project – Sunshine Zone Bond Ball Mill work

        index test results

 


Table 16.9  Livengood project summary of abrasion test results

        Phillips Report 093029_15

 


Table 16.10  Procedures for Knelson Concentrator test work

 


Table 16.11 Livengood Project – Main Zone Knelson Concentrator –

         gravity recoverable summary

 


Table 16.12 Livengood Project – Sunshine Zone Knelson Concentrator –

         gravity recoverable summary

 


Table 16.13 Livengood Project summary of cyanide bottle roll tests

 


Table 16.14 Livengood Project – Main Zone summary of CIL cyanide

         bottle roll tests

 


Table 16.15 Livengood Project – Sunshine Zone summary of CIL

        cyanide bottle roll tests

 


Table 16.16  Summary of batch flotation test results – March 2010

 


Table 16.17  Results of CIL bottle roll tests in gravity concentration

          tests March 2010

 


Table 16.18  Gold recovery estimates by ore type for heap leach and

          mill/CIP process scenarios

 


Table 16.19 Primary Ore Types Estimated Process Operating Costs for

         Heap Leach at 100,000 TPD Throughput Rate

 


Table 16.20 Livengood Capital Cost Estimate 100,000 TPD Heap leach

 


Table 17.1 Summary Mineral Resource

 


Table 17.2 Historical Drilling and Sampling

 


Table 17.3 ITH Drilling and Sampling

 


Table 17.4 Density Determinations

 


Table 17.5 Statistical Summary of Assay Data

 


Table 17.6 Gold Composite Statistics

 


Table 17.7Gold Indicator Statistics


Table 17.8 Average Gold Indicator Variograms

 


Table 17.9 Oxide Indicator Variograms

 


Table 17.10 KINT Dike Variograms

 


Table 17.11 Lower Sands, Shale, and Amy Seq. Variograms

 


Table 17.12 Model Extents

 


Table 17.13 Gold Kriging Plan

 


Table 17.14 Oxidized Kriging Plan

 


Table 17.15 Trace Oxidation Kriging Plan

 


Table 17.16 KINT Dike Indicator Kriging Plan – Southern Domain

 


Table 17.17 KINT Dike Indicator Kriging Plan – Northern Domain

 


Table 17.18 Lower Sands, Shale, and Amy Seq. Indicator Kriging Plan

 


Table 19.1 Comparison of resource estimates October 2009 and March 2010

 


Table 20.1 2010 Exploration Budget

 




LIST OF APPENDICES

Appendix

 


Appendix 1 Claim/Property Information

 


Appendix 2 List of Drill Holes

 




1.0

Summary


1.1

Summary


The Livengood property is the focus of ongoing exploration by International Tower Hill Mines Ltd. (“ITH”).  To date, 406 diamond and reverse circulation holes have been drilled on the property, and provide the basis for reporting a gold resource estimated at approximately 9.3Moz indicated and approximately 3.0Moz inferred at a 0.50 g/t Au cutoff (Carew, 2010).  This report is the seventh in a series of technical reports and the sixth in support of resource estimates regularly updated as new drill information has become available.  This report also provides documentation of the geological and resource estimation procedures that have been undertaken by ITH as they continue to advance this project toward development.  The currently reported estimate includes mineralization in the Core Zone and the new Sunshine and North east Zones.  The Core Zone has comprised the main part of previous resource estimates.


This report updates the October, 2009 technical report with a new resource estimate through addition of data from 64 drill holes received after completion of the October 2009 resource estimate.  This new data and the information provided herein will be used along with data from new metallurgical studies that are currently in progress for a new preliminary economic assessment of the Livengood property.  



1.2

Description and Location


The Livengood property is located approximately 115 km northwest of Fairbanks, Alaska in the Tolovana mining district within the Tintina Gold Belt.  The project area is centered on a local high point named Money Knob.  This feature and the adjoining ridge lines have been considered by many to be the lode gold source for the Livengood placer deposits which lie in the adjacent valley to the north where they have been actively mined since 1914 with production of more than 500,000 ounces of gold.


ITH controls 100% of its ~125 square kilometre Livengood land package, which is made up of Alaska State claims, fee land leased from the Alaska Mental Health Trust, and four leases with private holders of state and federal patented and unpatented mining and placer claims.



1.3

History


The property has been prospected and explored by several companies and private individuals since the 1970’s.  Geochemical surveys by Cambior in 2000 and AngloGold Ashanti (U.S.A.) Exploration Inc. (“AGA”) in 2003 and 2004 outlined a 1.6 x 0.8 km area with anomalous gold in soil.  Scattered anomalous samples continue along strike for an additional 2 km to the northeast and 1.6 km to the southwest.  Eight reverse circulation holes were drilled by AGA in 2003 and a further 4 diamond core holes were drilled in 2004 to evaluate this anomaly.  Favourable results from these holes revealed wide intervals of gold mineralization (BAF-7: 138.7m @ 1.07 g/t Au; MK-04-03: 55.3m @ 0.51 g/t Au) along with lesser intervals over a broad area.  Over the past 4 years, exploration by ITH through its wholly owned Alaskan subsidiary, Talon Gold Alaska, Inc ., has been aimed at assessing this area of mineralization through drilling diamond core and reverse circulation holes.



1.4

Geology and Mineralization


Rocks at Livengood are part of the Livengood Terrane, an east–west belt, approximately 240 km long, consisting of tectonically interleaved assemblages of various ages.  These assemblages include the Amy Creek Assemblage, which is a sequence of latest Proterozoic and early Paleozoic basalt, mudstone, chert, dolomite, and limestone.  In thrust contact above the Amy Creek Assemblage lies an early Cambrian ophiolite sequence of mafic and ultramafic sea floor rocks.  Structurally above these rocks lies a sequence of Devonian shale, siltstone, conglomerate, volcanic, and volcaniclastic rocks which are the dominant host to the mineralization currently under exploration at Livengood.  The Devonian assemblage is overthrust by more Cambrian ophiolite rocks.  All of these rocks are intruded by Cretaceous multiphase monzonite, diorite, and syenite stocks, dikes, and sills.  Gold mineralization is believed to be related to this intrusive event.


Gold mineralization occurs in two styles: as multistage fine quartz veins occurring in all lithologies (commonly in or near intrusive dikes and sills), and as diffuse mineralization within volcanic, intrusive, sedimentary, and mafic-ultramafic rocks without a clear quartz vein association.  Four principal stages of alteration are currently recognized.  These are an early biotite stage followed by albite-black quartz, followed by a sericite-quartz, and finally a carbonate stage.  Arsenopyrite apparently has been introduced during all stages, and gold correlates strongly with arsenopyrite, but it is not clear whether gold was introduced during all four stages or preferentially during one or more stages.


Mineralization is interpreted to be intrusion-related, consistent with other gold deposits of the Tintina Gold Belt, and has a similar As-Sb geochemical association.  Mineralization is controlled partly by lithologic units, but thrust-fold architecture is apparently key to providing pathways for magma (dikes and sills) and hydrothermal fluid.


Local fault and contact limits to mineralization have been identified, but overall the deposit has not been closed off in any direction.  The current resource and area drilled covers the most significant portion of the area with anomalous gold in surface soil samples, but still represents only about 25% of the total anomaly area.



1.5

Exploration, Drilling and Sampling


ITH has conducted drilling campaigns on Livengood property since 2006.  These programs initially identified mineralization in the Core Zone and then identified the Northeast and Sunshine zones through step out drilling and drill testing of areas with anomalous values in surface soil samples.


Nearly all drill holes at Money Knob have been drilled in a northerly direction at an inclination of -50 degrees in order to best intercept the south dipping structures and mineralized zones as close to perpendicular as possible.  A few holes have been drilled in other directions to test other features and aspects of mineralization.  Most holes have been spaced at 75m along lines 75m apart.  A few holes are more closely spaced.


Diamond core holes represent approximately 10% of the total number of holes drilled.  Core is recovered using triple tube techniques to ensure good recovery (>95%) and confidence in core orientation.  The core is oriented using the ACT system and/or the EZ Mark tool.


Reverse circulation holes are bored and cased for the upper 0-30m to prevent downhole contamination and to help keep the hole open for ease of drilling at greater depths.  Recovery of sample material from RC holes is done via a cyclone and dry or wet splitter according to conditions.  Sample chips are collected over the course of each five-foot interval and captured for a primary sample, an equivalent secondary sample (“Met” sample) and a third batch of chips for logging purposes.


Drill hole locations are determined by sub-meter differential GPS surveys at the drill collar.  Initial azimuth of drill hole collars is measured using a tripod mounted transit compass in conjunction with a laser alignment device mounted on the hole collar.


Down hole surveys of core and reverse circulation drill holes are completed using a Gyro-Shot survey instrument manufactured by Icefield Tools Corporation.  Results of surveys and duplicate tests show normal minor deviation in azimuth and inclination for drill holes.


All RC samples are “logged in” on site, analyzed with a field portable XRF (NITON) before being sealed in super sacks, and delivered to ALS Minerals in Fairbanks for preparation.  All core samples are initially logged at the drill rig for recovery, oriented features, RQD, and basic geologic features.  More thorough logging and core mark-up is done at the Livengood camp.  Core is sawed in half and bagged according to geologic intervals up to 1.5m and sealed in super sacks for delivery to ALS Minerals in Fairbanks.


Samples are analyzed by standard 50g fire assay for the gold determinations.  All core samples and select RC drilling samples are also submitted for multi-element ICP-MS analyses using a 4 acid digestion technique.  All RC samples are analyzed on site for trace elements using a Thermo Fisher Scientific NITON portable XRF before shipment to the laboratory.



1.6

Quality Assurance/ Quality Control and Data Verification


The QA/QC program implemented by ITH meets or exceeds industry standards.  A QA/QC program includes insertion of blanks and standards (1/10 samples) and duplicates (1/20 samples).  Blanks help assess the presence of any contamination that might be introduced by analytical equipment and help calibrate the low end of the assay detection limits.  Commercial standards are used to assess the accuracy of the analyses.  Duplicates help assess the homogeneity of the sample material and the overall sample variance.  ITH has undertaken rigorous protocols to assure accurate and precise results.  Among other methods, weights are tracked throughout the various steps performed in the laboratory to minimize and track errors.


Core and RC check samples have been collected during each drilling campaign by the lead author.  Results from these samples, as well as blanks and standards included, are consistent with ITH’s initial results.  This includes a similar increase in variance for samples at higher grades, a pattern consistent with nugget effect.  No systematic high or low bias has been observed.


Data entry and database validation procedures have been checked and found to conform to industry practices.  Procedures are in place to minimize data entry errors.  These include prenumbered, pretagged, bar-coded bags, and bar-coded data entry methods which relate all information to sample and drill interval information.  Likewise, data validation checks are run on all information used in the geologic modeling and resource estimation process.



1.7

Mineral Processing and Metallurgical Testing


ITH envisions a combined heap leach and mill recovery system for optimized gold recovery.  Test work undertaken to date is designed to determine optimal processes using combined methods.  This work involves studies to determine chemical and physical characteristics of the ore and metallurgical response to process treatment parameters according to ore type.  Test work includes assessment of grindability, abrasiveness, optimal particle size for downstream treatment, and response to leach or gravity recoveries as a function of oxidation and lithology.  Previous work completed was sufficient to enable an estimate of heap leach recoverable gold for a portion of the ore as reported in the October 2009 technical report.  Continued work on gold recovery from gravity, carbon in pulp, carbon in leach, and flotation methods is in progress with initial results presented in previous reports and new updated information added to this report.


Key findings to date include the following points:


      • Most Livengood ores can be considered moderately soft to moderately hard with an average Bond Ball Work index of 15.8 ranging from 11.1 to 19.1.
      • The majority of ores are considered non-abrasive with an average abrasion index of 0.0809 and a range of 0.0023 to 0.2872.
      • All Livengood ores respond to cyanide leaching to some degree.
      • Some unoxidized ore with organic carbon has “active” or “preg-robbing” carbon.
      • Leach times and gravity concentration indicate that some ores contain coarse gold.
      • Gold recovery exceeded 90% at 10 mesh for some ores.
      • Gold recovery improved for some ores with finer grinding.
      • Gold recovery for various leach tests suggests that organic carbon is present in varying degrees in some ores, particularly in unoxidized ore.
      • Carbon in Leach bottle roll tests indicate an average 84% recovery for the Sunshine Zone.
      • Gold with sulfide is not classified as refractory ore.
      • ombined gravity and flotation produced on average 88% recovery of gold.
      • Conventional milling using gravity recovery combined with intensive Carbon in Leach (CIL) leaching of gravity recovered gold concentrate achieved gold recoveries averaging 86%.


Continued test work is aimed at optimizing the combination of potential recovery techniques.



1.8

Resource Estimation


This report presents a resource estimate updated from the October 2009 estimate by incorporating data from an additional 64 drill holes.  The resource model was constructed using Gemcom GEMS® and the Stanford GSLIB (Geostatistical Software Library) MIK post processing routine.  The resource was estimated using multiple indicator kriging techniques.


Model parameters include, among others, two oxidation indicators and a single lithology indicator for each lithology.  A three-dimensionally defined lithology model, based on interpretations by ITH geologists, was used to code the rock type block model.  A three-dimensionally defined probability grade shell (0.1 g/t) was used to constrain the gold estimation.  Gold contained within each block was estimated using nine indicator thresholds.  The block model was tagged with the geologic model using a block majority coding method.  Because there are significant grade discontinuities at lithologic contacts, hard boundaries were used between each of the lithologic units so that data for each lithology was used only for that unit.


A summary of the estimated mineral resource is presented below for cutoff grades of 0.3, 0.5, and 0.7 g/t gold.  For comparison, previous estimates from the October 2009 report are included.


Model validation checks include global bias check, visual validation, and swath plots.  In all cases, the model appears to be unbiased and fairly represent the drilling data.


Classification

Au Cutoff (g/t)

Tonnes (millions) 10_2009

Tonnes (millions) 03_2010

Au (g/t)

10_2009

Au (g/t)

03_2010

Million Ounces Au

10_2009

Million Ounces Au

03_2010

Indicated

0.30

525

702

.65

0.60

11

13.5

Inferred

0.30

336

278

.61

0.56

6.6

5.0

Indicated

0.50

297

369

.85

0.78

8.1

9.3

Inferred

0.50

164

112

.84

0.77

4.4

3.0

Indicated

0.70

158

184

1.07

0.98

5.4

5.8

Inferred

0.70

78

56

1.11

0.99

2.8

1.8


It is important to note that compared to the October 2009 resource estimate, the tonnage and total ounces estimated have increased in the Indicated category and have decreased in the Inferred category for cutoff grades of 0.30, 0.50, and 0.70 g/t Au. This change is due, in part, to addition of newly defined lower grade tonnes in the northeastern area of the deposit.  In addition, in this estimate, the composite data were declustered prior to indicator bin selection and data in four octants were required to estimate a block compared to 3 octants in the 2009 estimate.  The net result is that some previously inferred higher grade blocks were downgraded or extinguished through a combination of drilling and use of more rigorous statistical criteria for allowing certain blocks to be included in the resource.



1.9

Conclusions


It is concluded that a substantial gold resource has been identified at Money Knob and the surrounding area.  Dedicated drilling has continuously enlarged the resource over the past several years.  Current metallurgical studies are underway and results indicate that gold is recoverable through heap leach, and combined mill, CIP, CIL, gravity, and flotation techniques.  Continuation of planned and in-progress metallurgical and ore processing studies will enable assessment of the best material processing and gold recovery techniques.  As results for this work are completed, new cost estimates that incorporate optimized gold recovery techniques will be used for a more comprehensive development plan and economic assessment.


1.10

    Recommendations


Exploration of the Livengood project should continue with the aim of advancing the project toward a prefeasibility status.  ITH plans to drill 40,000 m in 2010 to accomplish this goal.  The proposed program is an appropriate amount of drilling for the needs of the project and the time available in the field season.  Activities that will help advance the project in this direction include those listed below.

      • Continue step out drilling to identify the extent of mineralization.
      • Focus infill drilling on areas where Inferred resource blocks can be converted to Indicated resources.
      • Drill close spaced holes to define a variographic cross.
      • Drill core holes to gather sample material for advance metallurgical testing.
      • Continue and advance metallurgical, ore characterization, and mineral processing studies.
      • Assess geotechnical characteristics of the mineralized zone.
      • Begin the sterilization process for land that might be covered by facilities.
      • Continue and expand environmental base line studies including.
      • Complete the combined mill/heap leach preliminary economic analysis that is currently in progress.  This should evaluate the basic economic, logistic, and processing factors for a mining operation at Livengood

ITH has proposed expenditure of approximately $13 million dollars in 2010 for further evaluation of the Livengood project.  This budget will be allocated primarily to drilling and geological analysis of the deposit.  The budget is appropriate for the amount of drilling planned and feasible within the time allocated and the company has sufficient funds to accomplish this goal.  The authors recommend implementation of this program in order to accomplish ITH’s goal of advancing the Livengood project.


 

2.0

Introduction and Terms of Reference


2.1

Introduction


Mineral Resource Services Inc. (“MRS”), Reserva International (“RI”) and Pennstrom Consulting Inc. (“PCI”) have been requested by International Tower Hill Mines Ltd. (“ITH”) to provide an independent technical report on the Livengood gold project in the Tolovana Mining District of Interior Alaska.  The Livengood property is currently being explored by ITH through its wholly-owned subsidiary, Talon Gold Alaska, Inc. (“TGA”).


This report on the Livengood project uses new data from 64 drill holes to update the resource estimate completed in October 2009.  In this update, the resource estimate was prepared by RI.  Initial metallurgical and capital cost findings have been prepared by PCI, and geological and project information described by MRS.  The resource estimate presented in this report is based on drill hole and surface data through February 28, 2010.  Each author is a Qualified Person and is responsible for various sections of this report according to their expertise and contribution.  Dr. Klipfel of MRS is responsible for all sections of this report except Sections 16 and 17 as well as compilation of information.  Mr. William Pennstrom Jr. is solely responsible for section 16.  Mr. Timothy Carew is solely responsible for Section 17.  Each author has contributed figures, tables, and portio ns of Section 1 based on their respective contributions to this report.


The work presented here builds on and revises previous geologic, metallurgical and resource information reported in seven previous technical reports for the project (Klipfel, 2006; Klipfel and Giroux, 2008a; Klipfel, Giroux and Puchner 2008; Klipfel and Giroux, 2008b; Klipfel and Giroux, 2009; Klipfel, et al., 2009a; Klipfel, et al., 2009b).  Gold assays and analyses of other elements along with geological, structural, engineering, and metallurgical data is from 385 holes drilled by ITH and previous explorers, including 195 RC holes and 12 diamond core holes drilled in 2009 as well as data from previous drilling programs.


Information presented in this report is based on data provided to MRS, RI, and PCI by ITH as of February 28, 2010.  Data generated prior to 2006 was provided to ITH by AngloGold Ashanti (U.S.A.) Exploration Inc. (“AGA”).  This report also relies on personal observations made by:


      • Paul Klipfel in the course of seven field visits.
      • Tim Carew in the course of three site visits and generation of modelling data from primary data provided by ITH.
      • Bill Pennstrom, who made one site visit to Livengood and one visit to Fort Knox to identify operating costs at that mine.


and on general geologic information available to the public through peer review journals as well as publications by the U.S. Geological Survey and agencies of the State of Alaska.



2.2

Terms of Reference


Dr. Paul Klipfel of MRS, in Reno, Nevada, Mr. Tim Carew, of RI in Reno Nevada, and Mr. William Pennstrom Jr. of PCI in Denver, were commissioned by ITH to prepare this report on the Livengood project.  This report is based on data generated and results received by ITH through February 28, 2010 and is in support of metallurgical test work and resource information released to the public on March 8 and 10, 2010, respectively.  The 2010 winter drilling program is currently in progress, but no results from that program are utilized in this report.


Dr. Klipfel, Mr. Carew, and Mr. Pennstrom are independent consultants and are Qualified Persons (QP) for the purposes of this report as defined by Canadian Securities Administrators National Instrument 43-101 (“NI 43-101”).



2.3

Glossary of Key Abbreviations


ADEC

Alaska Department of Environmental Conservation

ADFG

Alaska Department of Fish and Game

ADNR

Alaska Department of Natural Resources

AGA

AngloGold Ashanti (U.S.A.) Exploration Inc.

AMHLT

Alaska State Mental Health Land Trust

BES

Barnes Engineering Services, Inc

BLM

U.S. Bureau of Land Management

g/t

grams/tonne

IRGS

Intrusion Related Gold System

ITH

International Tower Hill Mines Ltd.

KWh/T

kilowatt-hours per Ton

M

million

MRS

Mineral Resource Services Inc.

MTpa

million tonnes per annum

MW

megawatts

Opt

troy ounces per Ton

oz(s)

troy ounce(s)

PEA

Preliminary Economic Analysis

PCI

Pennstrom Consulting Inc.

QA/QC

Quality Assurance/Quality Control

QP

qualified person

ROM

run of mine

SHPO

State Historic Preservation Office

t

tonne

TGA

Talon Gold Alaska, Inc.

tpa

tonnes per annum

tpd

tonnes per day

tph

tonnes per hour

USACE

US Army Corps of Engineers

$ or USD

United States dollars

2.4

Purpose of Report


The purpose of this report is to provide an independent evaluation of the Livengood project, its exploration history, resource and mine development potential based on exploration work through February 28, 2010, a resource assessment based on that data, the discovery opportunity based on known geology and current exploration results, and to provide recommendations for future work.  This report conforms to the guidelines set out in NI 43-101.



2.5

Sources of Information


Information for this report was provided to the authors by ITH and consists of data generated by ongoing exploration by ITH and initial data from 2006 and earlier which was provided to ITH by AGA.  In addition, Dr. Klipfel has spent an aggregate of twenty five days on the site during seven visits reviewing core, examining outcrop, and discussing the project with on-site geologic staff and with Mr. Jeffrey Pontius, President of ITH.  In addition, Dr. Klipfel has undertaken independent petrographic evaluation of samples from the project.


Drilling, sampling, QA/QC, logging and sampling, and other exploration activities have been performed by contract geologic staff under the direction of Dr. Russell Meyers, Ph.D. (ITH VP Exploration) and Mr. Chris Puchner M.Sc. (ITH Chief Geologist; AIPG CPG 07048).  Both persons are Qualified Persons as per guidelines set out in NI 43-101.  Support for logistics, surveying, camp management, and digital modeling have been provided by Northern Associates of Alaska Inc. and their geologic, survey, and IT staff.  External consultants and engineering firms have been contracted for numerous functions including Giroux Consultants Ltd. of Vancouver, B.C., (previous resource evaluations), Barnes Engineering Services (previous resource evaluation), Mineral Resource Services, Inc. (petrographic evaluation), Three Parameters Plus, Inc. (environmental studies), Northern Land Use Research Inc. (archaeological surveys),  ABR Inc. (environmental studies), HDR Inc. (environmental studies), SLR Inc. (hydrology studies), Kappes Cassiday and Associates, (metallurgical test work), McClelland Laboratories Inc. (metallurgical test work), Hazen Research Inc. (metallurgical test work), and SRK Consulting (environmental and engineering consulting).


Gold assay and multi-element ICP data from drill hole samples used in the resource evaluation are from ALS Minerals (ALS; formerly known as ALS-Chemex).  ALS operates to international quality standards including compliance with ISO 17025 (www.ALSglobal.com).  The ALS analyses have been validated annually through cross-lab checks using SGS, ACT Labs, and Alaska Assay Laboratories.  Florin Analytical Services LLC. has provided analytical services for test work done by Kappes Cassiday.



2.6

Field Examination


Dr. Klipfel has visited the property seven times, with the most recent visit from February 21-24, 2010.  These visits included sequential updating of data, exploration activities, review of geologic sections, and interpretations of geologic staff.  Visits also included review of the physiographic, geologic and tectonic setting of the property, drill hole collar locations, surface and down-hole survey procedures, core orientation procedures as well as detailed examination of outcrop, drill core and RC chips.  Previous visits were during the following periods: October 5-8, 2009, June 17-24, 2009, September 22 – 26, 2008, June 30 – July 3, 2008, October 4-5, 2007, and June 6-7, 2006.  Independent check samples were collected during each of these visits and are described further in section 14.


Tim Carew has visited Livengood for a total of 26 days on three separate trips in 2009 and 2010.  During the course of these visits, modelling work was conducted collaboratively with ITH geologic staff, database information and contained data were reviewed and validated.


Mr. Pennstrom spent two days on site in May of 2009.  Site characteristics were reviewed with ITH staff.




3.0

Reliance on Other Experts


The preparation of this report has relied upon public and private information gathered independently by the authors and data provided by ITH and AGA regarding the property.  The authors assume and believe that the information provided and relied upon for preparation of this report is accurate and that interpretations and opinions expressed in them are reasonable and based on current understanding of mineralization processes and the host geologic setting.  The authors have used this information to develop their own opinions and interpretations along with external and independent understanding of geologic, metallurgical processing, and resource evaluation concepts and best practices.  The authors have endeavoured to be diligent in their examination of the data provided by ITH and the conclusions derived from review of that information or generated using that information.



4.0

Property Description and Location


4.1

Area and Location


The Livengood project is located approximately 115 km by road (85 km by air) northwest of Fairbanks in the northern part of the Tintina Gold Belt (Figure 4.1).  At this location, the property straddles, but lies predominantly to the north of, the Elliott Highway, the main road connecting Fairbanks with the Alaskan far north.  The property lies in numerous sections of Fairbanks Meridian Township 8N and Ranges 4W and 5W.  Money Knob, the principal geographic feature within the area being explored, lies near the center of the land holding and is located at 65o30’52’’N, 148o27’50’’W (UTM 6W 429600, 7265520; WGS84).


The explored area and current resource footprint reported here lies on the northwest flank of Money Knob and adjacent ridge lines and slopes, the extent of which remains to be determined.  This area lies within, and to the south of, a 1.6 x 0.8 km northeast-trending soil sample anomaly that was the initial target of interest for drill assessment.  The surface geochemical anomaly is situated within in a broader area of less pronounced anomalism that extends a further 2 km to the northeast and 1.6 km to the southwest.  This zone is described further in Section 9.0.  Continued drilling success has lead to several rounds of resource evaluation, the latest of which is the subject of this report.  At this time, mineralization continues to be identified as the area drilled expands outwards from an initial core zone centered over the geochemical soil anomaly.  Identified mineralization has local boundaries such as faults or contacts, but overall, the limits of this mineralized system have not been identified with mineralization effectively open in all directions.  The area with anomalous gold in soil samples has only been partially tested.



4.2

Claims and Agreements


The Livengood Property (Figure 4.2) consists of an aggregate area of approximately 12,499 ha (30,939 acres) controlled through agreements between TGA and the State of Alaska as well as between TGA and various private individuals who hold state and federal patented and unpatented mining and placer claims.  All property and claims controlled through agreements are summarized in Table 4.1 and listed in Appendix 1.  These agreements are with the AMHLT, Richard Hudson and Richard Geraghty, the estate of Ron Tucker, the Griffin heirs, and Karl Hanneman and the Bergelin Family Trust.  The AMHLT Trust Land Office manages approximately 1 million acres of Alaska land through the Department of Natural Resources (www.mhtrust.org) and generates revenue for the AMHLT through land leasing and fees for a range of resources. In February 2010, TGA increased its land position th rough the addition of AMHLT leased ground and Alaska State claims.  


The AMHLT lease (#9400248), signed July 1, 2004 by AGA and assigned to TGA on August 4, 2006 (as amended), includes advance royalty payments of $5/acre/year which escalates to $15/acre in years 4-6 and $25/acre in years 7-9.  The lease has a work commitment of $10/acre in years 1-3, $20/acre in years 4-6, and $30/acre in years 7-9.  The lease carries a sliding scale production royalty of 2.5% @ $300 gold up to 5% for a gold price more than $500.  In addition, an NSR production royalty of 1% is payable to AMHLT with respect to the unpatented federal

[exhibit1004.jpg]


Figure 4.1.  Location map showing the location of the Livengood project and the Tintina Gold Belt (orange dashed lines enclose the belt).


mining claims subject to the Hudson & Geraghty and the Hanneman and Bergelin Family Trust lease.  AMHLT owns both the surface and subsurface rights to the land under lease to TGA.


The Hudson and Geraghty lease, signed April 21, 2003 by AGA and assigned to TGA on August 4, 2006, has a term of 10 years and for so long thereafter as exploration and mining operations continue.  TGA is required to make advance royalty payments of $50,000 per year, which are credited to production royalties.  Production royalties vary from 2% to 3%, depending upon the price of gold.  TGA has the option to buy down 1% of the royalty for $1 million.  The 20 claims under this lease are unpatented federal lode mining claims that have no expiry but require a claim maintenance fee of $140/claim/year to keep them in good standing.


The Tucker mining lease of two unpatented federal lode mining and four federal unpatented placer claims has an initial term of ten years, commencing on March 28, 2007 and for so long thereafter as mining related activities are carried out.  The lease requires payment of advance royalties of $5,000 on or before March 28, 2009, $10,000 on or before March 28, 2010 and an additional $15,000 on or before each subsequent March 28 thereafter during the initial term (all of which minimum royalties are recoverable from production royalties).  ITH is required to pay the lessor the sum of $250,000 upon making a positive production decision.  An NSR production

[exhibit1006.jpg]

A)


[exhibit1007.jpg]

B)



Figure 4.2.  Land holding map showing the Livengood land position.  A)  The AMHL Lease holdings are shown in yellow, Alaska State Claims are shown in light blue, and holdings belonging to other parties shown in respective colors.  B) Detailed map of the individual claims within the AMHL Lease.

TABLE 4.1

SUMMARY OF CLAIM HOLDINGS AND ANNUAL OBLIGATIONS


Holder

Type of Holding

Current Year

2009 Holding Obligation

AMHLT

State Mining Lease

6

$249,250 advance royalty; no work expenditure owing as ITH has banked work commitments to 2013

Hudson and Geraghty,

20 Fed. unpatented lode claims

7

$50K advance royalty payment

Ron Tucker (estate)

2 Fed. unpatented lode claims

3

$5K

 

4 Fed. unpatented placer claims

3

Griffin heirs

3 patented Fed. claims

3

$15K

Karl Hanneman and the Bergelin Family Trust

169 Alaska State mining claims

4

$50K + $200k work expenditure and claim rental fees of $28,730

Alaska State Lands

89 Alaska State mining claims

1

$13,760 claim rental  paid with recording; $8,900 due by Sep. 30, 2010.



royalty of 2% is payable to the lessor.  ITH may purchase all interest of the lessor in the lease property (including the production royalty) for $1million.  The 6 leased claims are federal claims without expiry.  A fee of $140/claim/year or $140 worth of work/claim/year is required to maintain the claims in good standing.


The Griffin lease of three patented federal claims is for an initial term of ten years (commencing January 18, 2007), and for so long thereafter as the Company pays the lessors the minimum royalties required under the lease.  The lease requires minimum royalty payment of $10,000 on or before January 18, 2009, $15,000 on or before January 18, 2010, an additional $20,000 on or before each of January 18, 2011 through January 18 2016 and an additional $25,000 on each subsequent January 18 thereafter during the term (all of which minimum royalties are recoverable from production royalties).  An NSR production royalty of 3% is payable to the lessors.  ITH may purchase all interest of the lessors in the leased property (including production royalty) for $1 million (less all minimum and production royalties paid to the date of purchase), of which $500,000 is payable in cash over 4 years following the clo sing of the purchase and the balance of the $500,000 is payable by way of the 3% NSR production royalty.


The Hanneman/Bergelin Family Trust ground is held via a Mining Lease with Option to Purchase made and effective September 1, 2006.  The lease of 169 Alaska State mining claims is for a fixed term of ten years, commencing on September 1, 2006, without extension.  The lease requires payments to the lessors of $75,000 on execution, $50,000 in each of years 2-5 and $100,000 in each of years 6-10, work expenditures of $100,000 in year 1, $200,000 in each of years 2-5, and $300,000 in each of years 6-10 and a payment to the lessors of $250,000 upon a production decision being made by ITH.  An NSR production royalty of between 2% (gold below $350) and 5% (gold over $500) is payable to the lessors.  ITH may buy all interest in the property subject to the lease (including the retained royalty) for $10 million at any time during the term of the lease.

On Alaska State lands, the state holds both the surface and the subsurface rights.  State of Alaska 40-acre mining claims require an annual rental payment of $35/claim to be paid to the state (by November 20), for the first five years, $70 per year for the second five years, and $170 per year thereafter.  As a consequence, all Alaska State Mining Claims have an expiry date of November 30 each year.  In addition, there is a minimum annual work expenditure requirement of $100 per 40 acre claim (due on or before noon on September 1 in each year) or cash-in-lieu, and an affidavit evidencing that such work has been performed is required to be filed on or before November 30 in each year.  Excess work can be carried forward for up to four years.  If such requirements are met, the claims can be held indefinitely.  The work completed by ITH during the 2008 field season was filed as assessme nt work, and the value of that work was sufficient to meet the assessment work requirements through September 1, 2012 on all unpatented Alaska State mining claims held under lease.  Work completed in 2009 has been filed and the expenditure is sufficient to carry forward through 2013 for claims held prior to 2010.  Claims staked in 2010 will be subject to new work commitments.


Holders of Alaska State mining locations are required to pay a production royalty on all revenue received from minerals produced on state land.  The production royalty requirement applies to all revenues received from minerals produced from a state mining claim or mining lease during each calendar year.  Payment of royalty is in exchange for and to preserve the right to extract and process the minerals produced.  The current rate is three (3%) percent of net income, as determined under the Mining License Tax Law (Alaska).


All of the foregoing agreements and the claims under them are in good standing and are transferable.  Except for the patented claims, none of the properties have been surveyed.


Holders of Federal and Alaska State unpatented mining claims have the right to use the land or water included within mining claims only when necessary for mineral prospecting, development, extraction, or basic processing, or for storage of mining equipment.  However, the exercise of such rights is subject to the appropriate permits being obtained.



4.3

Permits and Environmental Requirements


Project activities are required to operate within all normal Federal, State, and local environmental rules and regulations.  These activities require permits from State and Federal Agencies including the United States Bureau of Land Management (BLM), United States Army Corp of Engineers (USACE), Alaska Department of Natural Resources (ADNR), Alaska Department of Fish and Game (ADFG), Alaska Department of Environmental Conservation (ADEC), and the State Historic Preservation Office (SHPO).


ITH staff and their subcontractors are conscientious in their care and diligence concerning historic features, flora and fauna, water quality, and general good stewardship toward the environment in their exploration activities.  This includes proper and environmentally conscientious protection of operational areas against spills, capture and disposal of any hazardous materials including fuel, drill fluids, and other materials used by equipment that are part of the drilling and exploration process.  Reclamation of disturbed ground and removal of all refuse is part of normal operations.


Operations which cause surface disturbance, such as drilling, are subject to approval and receipt of a permit from the ADNR and the BLM.  The ADNR permit for ground controlled by the State of Alaska was issued on January 26, 2009 and covers calendar years 2009 and 2010.  Exploration on Federal ground is permitted by the BLM under a Plan of Operations covered by EA-AK-024-08-010 (File FF095365) and is effective, without time limit, up until commencement of mining.

One of the USACE permitting requirements is that wetland sites be drilled in winter to minimize surface impact to vegetation and soil.  It also requires that all winter roads and pads in wetlands be fully reclaimed prior to April 15th.  Some slopes are covered in a patchwork of vegetation consistent with a wetlands designation.  These areas have been mapped by Three Parameters Plus, Inc., a natural resource consulting firm (Figure 4.3).  Based on a USACE Preliminary Jurisdictional Determination, a permit was granted on November 13, 2008 and enables ITH to drill in areas of shrub and tundra on and around Money Knob.  In support of this permit, the Alaska Department of Environmental Conservation (ADEC) has issued, on November 4, 2008, their Certificate of Reasonable Assurance for mineral exploration by ITH near Livengood.  These permits require ITH to comply with al l Federal and State regulations that apply to these areas.


The winter 2010 drilling program currently is operating on this permit.  In February, 2010, a new preliminary Jurisdictional Determination was made on the basis of the 3PPI mapping and an amended permit for further drilling outside of this area is currently being processed.


There are no known issues at this time that would hinder ongoing renewal of any permits.


There are no known issues concerning surface waters beyond normal operational obligations which fall under operating permits issued by the state as outlined above.


There are no known native rights issues concerning the project area.


With over 90 years of placer mining activity and sporadic prospecting and exploration in the region, there is moderate to considerable historic disturbance.  Some of the historic placer workings are now overgrown with willow and alder.  The old mining town of Livengood is now abandoned except for more modern road maintenance buildings at the town site.  ITH does not anticipate any obligations for recovery and reclamation of historic disturbance.


ITH commissioned Northern Land Use Research, Inc. (NLUR) to complete a cultural resource survey in 2008 (Figure 4.3).  An initial report was submitted to ITH in January, 2009 (Northern Land Use Research, Inc., 2009).  This Level 1 or Identification Phase survey was commissioned by ITH to locate and document historic sites, cultural features, or artifacts in the project area.  Twelve previously undocumented historic sites or artifacts were identified in 2008.  No prehistoric artifacts and no previously unknown prehistoric cultural resources were located in the 2008 exploration area.



[exhibit1009.jpg]


Figure 4.3  Map of the Money Knob area showing the archaeological study area, the location of cultural features identified in the survey, wetlands as currently assigned under the USACE preliminary wetlands interpretation, and the USACE permit area.  The Elliott Highway runs across the southern portion of the map area.  



A second cultural resource survey was conducted by NLUR during the summer of 2009 to cover a larger, expanded exploration area.  The survey documented historic (i.e. archaeological) mining equipment, buildings and linear ditch features, and relocated a previously known prehistoric site within the expanded coverage area (Figure 4.3).  Also, 12 select areas identified during the 2008 and 2009 programs were reviewed at a Phase II level (site documentation).  NLUR has provided recommendations which include a policy of feature avoidance to prevent damage to the condition or integrity of identified features.  All recommendations made by NLUR need to be made official by SHPO who will determine if any identified cultural resources require further action or isolation from disturbance.


Total disturbance associated with ITH’s exploration consists of drill pad access roads and drill pads.  However, as the number of drill holes increases, the local impact does as well.  An ongoing program of reclamation of pads and roads reduces the impacted area to the minimum possible at any given time.  For much of the exploration area, disturbance involves areas covered by secondary growth of alder, willow, and spruce and consequently, the impact is largely not visible from the Elliott Highway or the road into the Livengood town-site.  Visual impact is minimal.  The highest ground is naturally bare broken rock or sparsely covered in small shrubs and mosses.


Three Parameters Plus, Inc. of Fairbanks, AK, has been retained by ITH to: 1) conduct an initial baseline surface water sampling program to evaluate metal and organic content of streams that drain the project area as well as regional streams up-gradient from the project area; and 2) complete a wetlands inventory on and around ITH’s land position.


Water samples have been collected from 14 sites on a near monthly basis from March through October.  This survey is currently in progress but a 2009 report indicates apparent local and seasonal spikes among some analytes (Three Parameters Plus inc. 2009).  These are deemed to be mostly natural and, in part, a reflection of past placer mining activity.  Sampling will continue in order to develop base line trends for each sample location.  One well has been established to monitor the static water table fluctuations on Money Knob and water table measurements are taken on each drill hole upon completion.


ABR Inc. of Anchorage, AK conducted a survey in 2009 to assess quality and biodiversity of fish, benthic invertebrate, and periphyton populations in the streams that drain and are adjacent to the project area.  Surveys of this type are conducted at this early stage to determine the current conditions against which environmental quality metrics can be established should a mine be constructed.  Two separate attempts to identify fish populations that might be suitable for environmental monitoring, including both minnow traps and electrofishing, encountered only grayling, which are unsuitable for monitoring because of their migratory habits.  No other species were identified.


Wildlife in the area consists of moose, bear, and various small mammals.  None were observed in the course of the site visits although moose and bear have been seen in the vicinity.  Hunters can be active in the region and local trap lines may be present.  There are no known wildlife issues.


There are no known existing environmental liabilities.




5.0

Access, Climate, Local Resources, Infrastructure and Physiography


5.1

Access


The Livengood Project area is located approximately 115 km northwest of Fairbanks on the Elliott Highway, which provides paved year-round access to the area.  At present there are no full time residents in the former mining town of Livengood.  A number of unpaved roads have been developed in the area providing excellent access.


A 1400 foot runway is located 6 km to the southwest near the former Alyeska Pipeline Company Livengood Camp and is suitable for light aircraft.  Also, a small airstrip (currently out of service) is in Livengood Creek north of the project area.



5.2

Climate


The climate in this part of Alaska is continental with temperate and mild conditions in summer with average lows and highs in the range of 7 to 22oC.  Winter is cold with average lows and highs for December through March in the range of -27 to -5oC.  Annual precipitation is on the order of 23 cm which arrives mostly in the summer.  Winter snow accumulation ranges up to 66 cm (http://www.wrcc.dri.edu/cgi-bin/cliMAIN.pl?ak5534).



5.3

Local Resources


The project is serviced from Fairbanks, population 87,000.  As central Alaska’s principal center of commerce it is home to many government offices including the Alaska Division of Geological and Geophysical Surveys and the U.S. Geological Survey, as well as the University of Alaska Fairbanks.  The town is serviced by major airlines with numerous daily flights to and from Anchorage and other locations.  Helicopters and fixed wing aircraft are readily available.  Virtually all supplies necessary for the project can be obtained in Fairbanks.


On-site operations are conducted from a refurbished portion of the former Livengood Camp which was installed for the Alaska Pipeline construction.  Current camp facilities can accommodate up to 100 people, sufficient to meet the needs of the on-going exploration program.



5.4

Infrastructure and Physiography


The project is situated in forested hilly countryside with mature, subdued topography partly owing to widespread deposition of Pleistocene loess and gravel in valleys (Figure 5.1).  Elevation ranges from about 150m (~500’) in valley bottoms to 700m (2317’) at Amy Dome along the east side of the property.  Streams meander through wide, flat-bottomed, alluvial-filled valleys.  Ridge lines are generally barren with sparse vegetation.  Hillsides host mixed spruce-birch forest with abundant alder.



[exhibit1010.jpg]


Figure 5.1.  Photos of Money Knob and the project area.  A) Panoramic view looking west and north toward Money Knob.  Dashed yellow line outlines the perimeter of the area under investigation by drilling.  Blue lines to the right outline placer workings to the north in Livengood Creek.  B - D) Aerial view of Money Knob from the west and northwest showing the Lillian Fault (black line), and area under investigation by drilling (yellow dashed line).  The “Core Zone” is outlined with a dotted red line.  The Sunshine Zone is outlined with a pink dotted line.  Arrow indicates north.


The area is drained by Livengood Creek which flows to the southwest into the Tolovana River which then joins the Tanana River and ultimately the Yukon River approximately 190 km to the west.


Existing infrastructure includes a paved highway (the Elliott Highway) which passes through the property and within ~ 1.6 km of Money Knob.  Lesser unpaved roads are developed throughout the property.  A repeater tower has been built on Radio Knob approximately 1.6 km east of Money Knob.


Self generated power currently exists at the Livengood camp.  The nearest Alaskan grid power is approximately 67 km (40 miles) away at its closet point to the Livengood property.  A power line will need to be constructed for power supply to the proposed Livengood facility for operational demands.

 

 

 

6.0

History


Gold was first discovered in the gravels of Livengood Creek in 1914 (Brooks, 1916).  Subsequently, over 500,000 ounces of placer gold were produced and the small town of Livengood was established.  Since then, the primary focus of prospecting activity has been with the placer deposits.  Historically, prospectors have considered Money Knob and the associated ridgeline to be the source of the placer gold.  Prospecting in the form of dozer trenches was carried out for lode type mineralization in the vicinity of Money Knob primarily in the 1950’s.  However, to date no significant production has been derived from lode gold sources.


The geology and mineral potential of the Livengood District has been investigated by state and federal agencies as well as explored by several companies over the past 40+ years.  Modern mapping and sampling investigations were initially carried out by the U.S. Geological Survey in 1967 as part of a heavy metal assessment program (Foster, 1968).  Mapping completed in the course of this program recognized the essential rock relations, thrust faulting, and mineralization associated with Devonian clastic rocks, the thrust system and intrusive rocks.  These relations are summarized in the following insightful comment from the report summary.


“The small lode deposits in the upper plate rocks may represent leakage anomalies above economically significant metal deposits in rocks in or below the thrust fault zones.”


Since then, the Livengood placer deposits and the surrounding geology have featured in numerous investigations and mapping programs at various scales by the U.S. Geological Survey and the Alaska State Division of Geological and Geophysical Surveys.  Principal among these are: Chapman, Weber, and Taylor, 1971; Chapman and Weber, 1972; Cobb, 1972; Albanese, 1983; Robinson, 1983; Smith, 1983; Waythomas, and others, 1984; Arbogast, 1991; Athey and Craw, 2004; and Athey and others, 2004.


In 2003, as part of a larger state-wide program, the Alaska Division of Geological and Geophysical Surveys undertook a district-scale program of mapping and whole rock geochemical sampling in support of the mapping.  They report “one highly anomalous sample that yielded slightly over one ounce per ton gold” (Athey and Craw, 2004).


In addition to individuals prospecting the area, corporate explorers have investigated the potential for lode gold mineralization beneath the Livengood placers and on the adjacent hillsides including at Money Knob.  A summary of these programs is shown in Table 6.1.  Placer Dome’s work appears to have been the most extensive, but it was focused largely on the northern flank of Money Knob and the valley of Livengood Creek.


The most recent exploration history of Money Knob began when AGA acquired the property in 2003 and undertook an 8-hole RC program on the Hudson-Geraghty lease.  The results from this program were encouraging and were followed up with an expanded soil geochemical survey which identified anomalous zones over Money Knob and to the east.  Based on the results of this and prior (Cambior) soil surveys, 4 diamond core holes were drilled in late 2004.  Results from these two AGA drill


TABLE 6.1

EXPLORATION HISTORY


Company / Year

Major Activity

Results

Comment

Homestake / 1976

Geochemistry & 6 boreholes

Significant soil anomaly, low grade gold in drill holes and auger samples

Management decided on other priorities.

Occidental Petroleum / 1981

6 boreholes

Low-grade gold encountered in several holes

Other priorities.

Alaska Placer Development 1981 - 1984

Extensive soil and rock sampling together with mapping, mag, EM, trenching and auger drilling.

Defined soil and rock anomalies; other data not available.

Mostly on flanks of Money Knob.  Changed focus to placer deposits.

Amax / 1991

3 RC holes; surface geochemistry and auger  testing

Good geological mapping, lots of rock sampling, low grade gold in drill holes.

Other priorities.

Placer Dome / 1995 - 97

Surface exploration; / geophysics & 9 diamond core holes

Intersected some moderate grade mineralization.

Work focused to north of Money Knob. Limited land position.

Cambior 1999

Geochemistry

First to identify the extent of gold on Money Knob.

Corporate restructuring – no follow-up.

AGA / 2003-2005

Geochemistry, trenching, geophysics, drill testing;  

Geochemical anomaly, numerous drill intersections

Intersected gold-bearing intervals.

ITH 2006-2007

Surface geochemical sampling; drilling 23 holes

First intersection of extensive zones of > 1g/t Au.

Intersected more gold-bearing intervals; initial resource estimates

ITH 2008

108 reverse circulation, 7 diamond core holes, and 4 trenches through September 27.

Infill and step-out grid drilling of mineralization

Expanded resource estimates.

ITH 2009

195 reverse circulation holes; 12 diamond core holes

Infill drilling in wetland areas; discovery of Sunshine Zone and other areas of mineralization; expanded resource estimate

Early results expanded the resource estimate presented in October, 2009; later results  discussed in this report



programs were deemed favourable but no further work was executed due to financial constraints and a shift in corporate strategy.


In 2006, Livengood and other properties now part of the ITH portfolio were sold to ITH by AGA.  In the same year, ITH drilled a 1227 m, 7-hole program.  The success of this program led to the drilling of an additional 4400 m in 15 diamond core holes in 2007 to test surface anomalies, expand the area of previously intersected mineralization, and advance geologic and structural understanding of subsurface architecture.


Geophysical work in the vicinity includes an airborne magnetic survey by Placer Dome in 1995.  This data has not been recovered.  They also conducted VLF surveys in the northern part of the district in 1996 with only limited success because of the mixed frozen and thawed ground.  This data is only partially preserved.  The state of Alaska flew a 400 meter line spaced DIGHEM survey (an aerial, multi-channel electromagnetic technique) over the Livengood District in 1998 (Burns and Liss, 1999; Rudd, 1999).  AGA ran a series of CSAMT (Controlled-Source Audio-frequency Magneto-Telluric) lines across Money Knob in 2004.  This survey was designed to look for resistive intrusive bodies in the subsurface.  The survey appeared to map the main thrust zone but did not appear to delineate hidden intrusive bodies.



7.0

Geological Setting


7.1

Regional Geology


The Livengood ‘district’ is a portion of the broader Tolovana Mining District.  It is situated in a complex assemblage of rocks known as the Livengood Terrane (Figure 7.1).  This Terrane is an east–west-trending belt, approximately 240 kilometres long, bounded on the north by splays of the dextral Tintina-Kaltag strike-slip fault system and other terranes to the south (Silberling and others, 1994; Goldfarb, 1997).  It is composed of a complex sequence of rocks which do not match assemblages of the adjacent Yukon – Tanana Terrane.  Throughout the Livengood Terrane, individual assemblages of various ages are tectonically interleaved.  Each assemblage, and perhaps the stratigraphy within each



[exhibit1011.jpg]


Figure 7.1. Terrane map of Alaska showing the location of the Livengood Terrane (LG; red arrow).  The heavy black line north of the Livengood Terrane is the Tintina Fault.  The heavy black line to the south of the Livengood and Yukon – Tanana Terrane (YT) is the Denali Fault.  The Tintina Gold Belt lies between these two faults.  After Goldfarb, 1997.

assemblage, is bounded by both low to moderate (?) angle thrust faults and steep faults, of which at least some of the latter type are interpreted to be splays of the Tintina Fault system.  Rocks of the Livengood Terrane are generally highly deformed, but weakly metamorphosed Neoproterozoic to Paleozoic marine sedimentary rocks along with Cambrian ophiolitic sequences, Ordovician Livengood Dome chert, overlying dolomite, volcanic rocks, terrigenous clastic rocks, and minor Devonian limestone (Silberling, et al., 1994; Athey et al., 2004).


The Livengood Terrane is overprinted by later Mesozoic intrusions believed to have originated in the back-arc position above subducting oceanic crust.  These intrusions are quartz monzonite to diorite to syenite in composition and some of them are believed to be responsible for gold mineralization of the Tintina Gold Belt (McCoy, et al., 1997; Goldfarb, et al., 2000).  The Livengood district occurs within the Tintina Gold Belt, an arcuate belt of gold mineralization that extends from the Yukon to south-western Alaska and hosts numerous gold deposits, including Fort Knox and other deposits of the Fairbanks District and the Donlin Creek deposit in the Kuskokwim region (Smith, 2000).



7.2

Local Geology


In the vicinity of the Livengood project, the oldest rocks are Neoproterozoic to early Paleozoic basalt, mudstone, chert, dolomite, and limestone of the Amy Creek Assemblage (IPzZ units on Livengood geology map; Athey et al., 2004) (Figures 7.2 and 7.3).  These units are believed to represent incipient ocean floor basalt in a continental rift system and overlying sediments.  The origin and age are poorly constrained but fossil evidence suggests a depositional age between Neoproterozoic and Silurian time.


Above the Amy Creek Assemblage lies an early Cambrian ophiolite sequence (Plafker and Berg, 1994).  This assemblage consists of structurally interleaved greenstone, pyroxenite, metagabbro, layered metagabbro, ultramafic rocks and serpentinite derived from them (Figures 7.2 and 7.3).  Metamorphic ages suggest that this assemblage was tectonically emplaced over the Amy Creek Assemblage by north-directed thrusting during Permian time.


The Cambrian ophiolite sequence is, in turn, overlain by Devonian rocks which include shale, siltstone, conglomerate, volcanic, and volcaniclastic rocks (Figures 7.3 - 7.6).  This assemblage is the principal host for gold mineralization.  These rocks have been subdivided into “Upper” and “Lower” sedimentary units with volcanic rocks separating them (Figure 7.3).  The Upper Sediments consist of siltstone, sandstone, conglomerate, shale, and minor limestone and dolomite.  The Lower Sediments unit is dominantly shale in the northern portion of the property but includes sandy siltstones and fine sandstones to the south.  Use of trace element ratios has helped discriminate these units from one another.  The volcanics consist of flows and pyroclastic rocks.  Some of these volcanic rocks were previously mapped as Cretaceous intrusive rocks (Athey et al., 2004).  However, geologic observations in drill core and the use of trace element ratios indicate that most of the rocks mapped as the “Ruth Creek” and “Olive Creek” plutons are volcanics and part of the Devonian stratigraphy.


Structurally above the Devonian assemblage is a klippe of the Cambrian ophiolitic mafic and ultramafic rocks with tectonically interleaved wedges of cherty sedimentary rock (Figures 7.3 and 7.4).  The emplacement of this klippe may have taken place in Cretaceous time during closure of the

 

[exhibit1012.jpg]


Figure 7.2.  Geologic cross section and map of the Livengood project area (Athey, et al., 2004).  A) Cross section through Money Knob illustrating the geological components of the Livengood District.  lPZZmc are older siliceous shelf metasediments.  Cs, Cgs and Cmg are Cambrian mafic and ultramafic volcanics and intrusive rocks of oceanic ophiolitic affinity.  Dc represents Devonian siliciclastic sediments.  The thrust imbrication may reflect two deformation events, one in the Permian and one in the Middle Cretaceous.  The thrust package has been intruded by a number of Cretaceous felsic intrusions.  B) Geologic map showing the location of the cross section ‘A-A’.  Pink symbols identify rocks mapped as intrusive and mostly known now to be Devonian volcanics.


[exhibit1013.jpg]



Figure 7.3.  Diagrammatic lithologic column shows the tectonic stacking of rock groups in the Livengood area.



Manley Basin south of the project area.  The thrust contacts between the various rock units indicates that there has been extensive thrust stacking and interleaving of the different assemblages as well as possible local interleaving of some units within the assemblages.


Rocks in each of these assemblages have been folded, but overall, they strike east-west to northwest-southeast and dip shallowly to moderately south, consistent with postulated northward directed thrust transport.


[exhibit1015.jpg]


Figure 7.4.  Generalized geologic map of the Money Knob area based on geologic work by ITH.  



Drill intercept patterns and foliation-bedding relations observed in core (Figures 7.6 d and e) indicate that these rocks define a principal recumbent fold and possible parasitic folds segmented by south-dipping thrust and normal faults.  Later Cretaceous dikes and sills intrude the sequence, some of which are believed to intrude along these faults.


The thrust-stacked Paleozoic sequence described above is intruded by back-arc Cretaceous (91.7 – 93.2 m.y.; Athey and Craw, 2004) multiphase monzonite, diorite, and syenite stocks, dikes, and sills with equigranular to porphyritic textures.  Athey et al. (2004) concluded that the intrusive rocks were the primary host to the gold mineralization.  However, exploration work since then has shown that these rocks are, in part, Devonian volcanics which have undergone extensive alteration along with introduction of mineralization in or associated with quartz and quartz-carbonate veins.  Narrow Cretaceous stocks (?) and large dikes are biotite monzonite.  Narrower, late (?) stage dikes are composed of feldspar porphyry, and aplitic felsic intrusives without biotite (Figure 7.6).  Mineralization is, at least partially, associated spatially and probably genetically with these dikes.

[exhibit1016.jpg]


Figure 7.5.  Photographs of key rock types at Livengood. A) ultramafic rock with carbonate alteration (yellow-brown); MK7-20, 13.5m;  B) siltstone with carbonate and pyrite knots.  Brown color is oxidation front.  MK 07-18, 8.5m  C) sedimentary conglomerate; at least some clasts appear to be rip-up clasts of similar sedimentary rocks; brown color is after introduced carbonate; MK07-18, 41.2m;  D) sedimentary conglomerate; at least some clasts appear to be rip-up clasts of similar sedimentary rocks; brown color is after introduced carbonate; MK07-18, 57.7m;  E) argillite with pyrite; MK07-20, 222m;  F) argillite with siltstone band; MK07-18, 280; G) tuff showing lithic fragments; this unit contains MK07-18, 190m 0.23 – 0.75 g/t Au;  H) fine-graine d tuffaceous sediment; MK07-20, 151.5m.



The structural architecture of the project area is characterized by fold-thrust patterning, apparently overprinted by local, minor normal offset along primary normal faults or reactivated thrust faults (Figure 7.7) and a possible second fold event.  Apparent upright open folds have axes that strike NW and plunge gently in that direction.  Later faults include the Lillian and the Myrtle Creek.


Thrust faults appear to lie in two principle dip orientations; subhorizontal and low to moderately south-dipping.  Undulatory subhorizontal thrust faults appear to define the primary thrust surface separating the Cambrian ophiolite sequence from underlying Devonian sedimentary and volcanic sequence.  These rocks and their low angle thrust contact appear to be segmented and offset by low to moderately south-dipping thrust faults.  In some instances, these south-dipping structures display apparent normal offset.  Details of this patterning are currently being evaluated but possible interpretations include: 1) post-thrusting tectonic relaxation resulting in minor normal offset on reactivated thrust surfaces; 2) the existence of a late-stage extensional tectonic event; or 3) some, as yet, poorly understood complex relation between faults.  Correlation of particular faults from one drill hol e to another is subject to different possible interpretations.  Key points that need to be resolved, if possible, relate to distinguishing low angle and south-dipping structures and the relative timing of these features.


The Lillian Fault is a northwest trending, steeply south-dipping fault that is characterized by a wide zone of sheared sedimentary and dike rocks that separates the property into two domains.  To the south, the structural and stratigraphic sequence is well-defined consisting of gently south-dipping sedimentary and volcanic stratigraphy and thrust faults.  These rocks host the Core Zone and surrounding mineralization.


To the north of the Lillian Fault, the upper Cambrian ophiolite sheet is not preserved and the upper sedimentary sequence is much thicker than the sequence preserved south of the Lillian Fault.  Immediately to the north of the Lillian fault the stratigraphy dips very steeply to the north and strikes parallel to the Lillian Fault suggesting that movement on the fault was reverse at some time.  The mineralized areas north of the Lillian fault is known as the Sunshine Zone where mineralization is related to a dike swarm in the steeply dipping sedimentary and volcanic rocks.


Immediately south of the fault, the axis of a north-vergent, major recumbent fold is subparallel to the strike of the Lillian Fault.  This implies that, during the early history of the fault, there may have been steep reverse movement followed by later collapse and normal offset with down drop to the south.  At present, subhorizontal lineations are common on faults in and around the Lillian Fault suggesting a history with possible late strike-slip movement.  Regional Mesozoic to Cenozoic dextral slip on the Tintina-Kaltag Fault system to the north of Livengood may support an interpretation of late dextral motion on the Lillian Fault.


[exhibit1017.jpg]


Figure 7.6.  Photographs of key rock types and mineralization features.  A) porphyry dike; MK07-18, 41.2 m; 1.01 g/t Au  B) amygdaloidal volcanic, presumably a flow, with possible Na alteration; MK07-18, 152-189  C) silicified volcanic breccia; MK07-18  D) argillite with more silty band and coral hash; note the shearing which is approximately 30o to bedding; MK07-18, 288.4m  E) axial planar cleavage on fold nose in interlayered argillite – silty argillite; MK07-18, 296.11m.  This type of feature supports the fold-thrust interpretations of the cross section shown in Figure 10.  F) fault; broken siltstone fragments in clay gouge/shear zone; this is part of an ~8m interval which contains 2 – 22.4 g/t Au; MK07-18, 77.9 – 86.08m;  G) bro ken rock in shear zone within mineralized interval.  The material in the photo includes portions of sample intervals that contain 15-16.2 g/t Au; MK 07-18, 96.93m  H) narrow mineralized quartz vein in silicified volcanic contains 13 g/t Au and 35,900ppm As from arsenopyrite;  MK07-18, 136.5m.



To the west of the deposit, the approximately north-south Myrtle Creek Fault (Figure 7.2) is mapped as having strike-slip offset by early workers and west-side-down, normal offset by Athey, et al. (2004).  It is believed that offset along this fault influenced the paleo-drainage system of the area.  Based on a number of lines of evidence, it is proposed that Livengood Creek used to flow to the northeast.  Capture of the stream by the Tolovana River, and reversal of flow could have been related, in part, to movement along the Myrtle Creek Fault (Karl, et al., 1987; Athey and Craw, 2004).  The origin and relationship of this fault to other structural elements in the area is not understood.  It lies in an anomalous direction, but also extends for several 10s of kilometres to the south and a lesser distance to the north.  This fault is not known to affect mineralization and is p eripheral to the area of interest at Money Knob.


Immediately to the south of Livengood, the early to middle Cretaceous Manley Basin is preserved as a fold thrust sequence.  Asymmetric overturned folds indicate a northern vergence direction to this deformation event.  The precise age of the deformation is not well constrained but the youngest fossils in the basin are Aptian (125 – 112 m.y.) and the sequence was folded and thrusted prior to the emplacement of the 90Ma monzonitic intrusions in the thrusted sediments (Reifenstuhl et al., 1997).  Because rocks of the Livengood Terrane at Livengood lack structural markers, it is not possible to determine if the fold-thrust deformation and closure of the Manley Basin impacted the older Livengood sequence.  However, given the close spatial proximity of the two sequences and the fact that they are in thrust contact elsewhere, it seems likely that the Cretaceous deformation event affected the L ivengood area.  The extent to which thrust deformation at Livengood is Cretaceous or earlier (Permian), and which rocks were affected at which time is currently being evaluated by ITH geologic staff.  In addition, there is the possibility that multiple thrust events are overprinted by one or more (?) extensional events.  As the Livengood project advances, structural interpretations will continue to mature and some structural interpretations may change as more information becomes available.



7.3

Geological Interpretation


Geologic interpretation at Livengood depends on surface information gained through mapping and examination of outcrops, exposures in road cuts, and trenches.  Subsurface information is acquired from diamond drill core and RC drill chips.  Drill core provides clear macroscopic visual information on rock type and structural features.  RC chips also provide visual information on rock type, but no structural information.  In core, the orientation of structural elements (joints, faults, veins, contacts, etc) are measured and used to help understand the relative relations of structural components.  Visual


[exhibit1018.jpg]




Figure 7.7. This cartoon shows an interpretive sequence of north-south sections and events to explain the structural relations observed at the surface and in drill core.  The details and sequence of the events shown here are partly the interpretations of Dr. Klipfel.  ITH staff geologists are currently developing new hypotheses concerning the relative sequence and suggest that normal faulting has played a role in development of the structural architecture.  One possibility is that the Cambrian ophiolite sequence was thrust in the Cretaceous, possibly contemporaneous with the closure of the Manley Basin to the south of Livengood.

A)

Devonian volcano-sedimentary sequence is deposited.  Pink – volcanics; light gray – upper sediments; dark gray – lower sediments; blue-green – other sediments likely to be present in the Devonian sequence, but not yet identified in outcrop or drill holes.

B)

A compressional event (heavy black arrows) causes initial asymmetric folding typical of early stages in the development of a fold-thrust belt.  Dashed line shows where incipient thrust truncation will develop.

C)

Cambrian ophiolitic basalt, ultramafic rocks (serpentinite), and gabbro (green) along with tectonic thrust wedges of chert (Amy Creek) and other sediments (pale yellow) are thrust over the folded Devonian volcano-sedimentary sequence.  The thrust surface is undulatory but overall is subhorizontal in orientation.  ITH geologic staff is currently attempting to establish if this event happened in the Cretaceous as part of the deformation event that impacted the Manley Basin to the south or if it is the product of an earlier, possibly Permian deformation event.  Dashed lines show where the next stage of faulting occurs.

D)

Continued (?) thrusting causes thrust stacking along structures that dip 30-45 degrees.  Earlier folds and the Cambrian-Devonian thrust surface are segmented with reverse offset.

E)

Tectonic relaxation after thrusting or a tectonic extensional event following fold-thrust compression allows for normal offset, particularly along some pre-existing faults, particularly the most recent thrust faults shown in D.

F)

Cretaceous dikes (red) of various composition and crystalline character infiltrate the region, particularly along pre-existing faults that dip 30-45 degrees.  Dikes intrude all rock types and generally do not occur along the earliest thrust surface that separates the Cambrian ophiolite sequence from the Devonian volcano-sedimentary sequence.

G)

Erosion to the current topography removes much of the over-thrust Cambrian ophiolitic sequence.  Also, other faults such as the Lillian Fault (steep fault at far right) may have formed during or after extensional tectonism.  This fault separates like rocks but with different orientations.



examination of core is used to assess rock type and alteration.  Petrographic examination of select samples has helped determine alteration mineralogy and relative timing of successive alteration events.


In addition, rock composition is determined for RC samples through use of a portable XRF device (Thermo Fisher Scientific Niton XLT3) which provides a semi-quantitative measure of select elements, which are generally diagnostic of each rock type intersected in the drill hole.  Multielement ICP analyses provide additional data for geochemical evaluation of the rocks by principle component analysis.  This technique utilizes the relative abundance and ratios of various immobile elements and has enabled discrimination of Devonian volcanics from Cretaceous intrusive and dike rocks as well as the upper and lower sedimentary assemblages.  Procedures used by ITH for rock type discrimination rely on consistency between visual and chemical assessment of rock type.  These procedures are described more fully in section 13.2.


At the district scale, thrust stacking of rock assemblages (Amy Creek, Cambrian ophiolite, Devonian sedimentary and volcanic rocks) is reasonably well understood.  Drilling reveals that there are numerous local fold and thrust complications which are only partially understood at this stage.  It is likely that faults and fractures produced during fold-thrust deformation, along with possible overprinting extensional deformation, produced architecture that enabled localization of dikes and auriferous hydrothermal fluid.  Gold mineralization largely appears to be controlled by and is spatially related to the fault architecture.  The gold mineralization envelope encloses and lies parallel to axial planes of thrust-related recumbent folds.  It appears as if mineralization occupies a broad ‘damage zone’ related to the fold-thrust architecture.  Patterning in the resource block m odel is consistent with this interpretation.


The location and density of veins and diffuse mineralization appears to be controlled by lithology.  Mineralization spatially associated with dikes appears to occur within ‘damage zones’ related to the south-dipping faults.  However, the exact relationship and relative orientations of these features is not fully understood.  Structural measurements in drill core indicate that the dominant dike orientation is east-west with dips 30-50 degrees to the south.


Many of the dikes are in faults or are bounded by faults suggesting that they, at least partially, follow thrust faults.  Measured fault orientations in core reveals a broad scatter of attitudes but with clustering generally coincident with dike orientations.  This pattern of partial coincidence between dikes, faults, and mineralization envelopes reinforces the interpretation that the dikes and faults are key controls for mineralization.


Despite these apparent relations, mineralization in sections 428625, 428850, 428925, and 429675 appears to follow, in particular, the Devonian volcanic unit as well as lie oblique to the thrust fault contact between rocks of the Cambrian ophiolite and the Devonian assemblage (Figures 7.8 – 7.11).  Although it is not possible to reliably correlate individual dikes between the drill holes on these sections, it is clear that the 30-50 degree dip of the dikes and associated structures is compatible with the southerly dipping zones of mineralization.  These relationships need further evaluation.  Improved understanding ought to offer predictive information for the location of more mineralization.



[exhibit1020.jpg]


Figure 7.8.  N-S Section 428625 E illustrates the complexities of thrust and normal fault interpretation and shows the southerly dip of high grade zones (red).



[exhibit1022.jpg]


Figure 7.9.   N-S Section 428850 illustrates the southerly dip of high grade zone (red) along the general stratigraphic pattern.  



[exhibit1024.jpg]


Figure 7.10.  N-S Section 428925 illustrates the general southerly dip of mineralization and how it lies along the stratigraphic and structural grain.




[exhibit1026.jpg]


Figure 7.11.  N-S Section 429075 illustrates the pattern of mineralization reflecting structural and stratigraphic controls.



[exhibit1028.jpg]


Figure 7.12.  N-S Section 429675 illustrates the pattern of mineralization reflecting structural and stratigraphic controls.



8.0

Deposit Types


Gold occurs in vein, veinlet, and disseminated styles of mineralization.  It occurs in and adjacent to narrow (≤10 cm) multistage quartz veins dominantly in volcanic rocks, but also in intrusive, sedimentary, and ophiolite rocks, generally in or near intrusive dikes and sills.  Gold also occurs as diffuse mineralization through the same rocks without a clear association with quartz veins.  Many of the dikes appear to fill thrust-related structures and some of the diffuse mineralization occurs in envelopes around these zones.


The structural architecture, host lithologies, styles of alteration, inferred fluid chemistry, and metallogenic association of As, Sb, ±W, Bi, and very minor Cu and Zn at Livengood show similarities to several styles of gold mineralization and deposit types.  Principal among these is the occurrence of Livengood in the Tintina Gold Belt where gold mineralization is hosted in or genetically associated with mid- to late-Cretaceous reduced I-type intrusions (Newberry and others, 1995; McCoy and others, 1997).  Mineralization at Livengood appears to be associated genetically with 91.7 – 93.2 m.y. back-arc Cretaceous dikes (Athey and Craw, 2004).  For this reason, Livengood should be considered most closely aligned with intrusion-related gold system (IRGS) type deposits.


Among deposits of the Tintina Gold Belt, Livengood mineralization appears to be most similar to the dike and sill-hosted mineralization at Donlin Creek deposit where gold occurs in fine quartz veins associated with dikes and sills of similar composition (Ebert, et al., 2000).  However, unlike Donlin Creek, the gold at Money Knob is not tied up in the lattice of arsenopyrite.  Instead, it occurs as native gold grains in and around the pyrite and arsenopyrite grains.


The gold-arsenopyrite-stibnite metal association hosted, in part, by sedimentary rocks with dikes associated with a thrust fault system is also reminiscent of sediment-hosted disseminated deposits (SHD) of the Great Basin (aka Carlin type deposits).  Foster (1968) initially proposed this potential similarity of mineralization types and Poulsen (1996) speculates on the potential of this type of deposit in the Canadian Cordillera which overlaps in its northern portion with the Tintina Gold Belt.  While there are similarities, Livengood lacks prolific decalcification, jasperoid, and a moderate to strong Hg association which are important characteristics of SHD-type deposits.  The association of mineralization with intrusions and possible similar structural preparation for both deposit types may be important.


Vein and diffuse gold mineralization along with the metallogenic association and alteration types are most consistent with IRGS type deposits.  The mineralogy, alteration types, and geochemical association of As-Sb suggests mineralization formed at a crustal level higher than mesothermal depths (~5-10 km) and deeper than shallow epithermal systems (≤3 km).


 

9.0

Mineralization


9.1

Mineralization


Historically, the Livengood district has been known for its >500,000 ounce placer gold production.  The source of this gold is unknown, but the principal drainages which fed the placer gravels are sourced from Money Knob and the associated ridgeline.  Prospecting in this area has revealed numerous gold-bearing quartz veins, generally associated with dikes, sills and stocks of monzonite, diorite, and syenite composition.  The reduced magma type and porphyritic to brecciated textures as well as local zones rich with arsenopyrite, are characteristics common to many deposits of the Tintina Gold Belt (e.g. Brewery Creek, Donlin Creek) (McCoy, et al., 1997; Smith, 2000).


No lode production has taken place at Money Knob.  Exploration of the area by various companies, including soil surveys by Alaska Placer Development, Cambior, AGA and ITH, reveals a 6 x 2 km northeast-trending anomalous area in which a 2.2 x 1.5 km area (~25% of the anomaly area) forms the locus of current exploration interest (Figure 9.1).  Despite drilling of 311 holes to September 25, 2009, this area has been only partially drill tested.  At this time, mineralization shows local fault and contact boundaries such as the Lillian Fault, but overall is open in all directions.


Drilling since 2003 by AGA and ITH has resulted in identification of an indicated and inferred gold resource interpreted to be part of a large IRGS deposit, the details of which are discussed further in section 17.


Mineralization consists of gold in multi-stage quartz, quartz-carbonate, and quartz-carbonate-sulfide veins and veinlets as well as disseminated throughout altered rock with arsenopyrite and Fe-sulfides.  Gold mineralization in the Core Zone preferentially occurs in Devonian volcanics and Cretaceous dikes but also occurs in Upper and Lower Sediments as well as locally in the overthrust ultramafic rocks primarily where dike rocks are present.  Mineralization associated with Cretaceous dikes may also be spatially associated south dipping faults.  Overall, the mineralization envelope appears to dip south along with the dikes and faults.


Better gold values (>1 g/t) tend to be associated with the Devonian volcanics, Cretaceous dikes, dike margins and in broad zones within adjacent volcanic and sedimentary or mafic-ultramafic rocks.  Visible gold occurs locally, particularly in quartz veins and with isolated coarse blebs of arsenopyrite and/or stibnite.  Where gold occurs in sedimentary host rocks, veins are most common in brittle siltstone, sandstone, and pebble conglomerate as opposed to shale.  The diffuse style of mineralization is spatially associated with areas containing vein mineralization, but disseminated mineralization can be present where there is no discernable quartz veining to explain it.


In contrast to the Core Zone, mineralization north of the Lillian Fault within the Sunshine Zone is hosted dominantly in Upper Sediments.  In this zone, mineralization is related spatially to swarms of dikes which appear to dip steeply to the south in a package of sediments that dips steeply to the north.  Disseminated sulphides in the Sunshine Zone as in the Core Zone, but two things distinguish it from other parts of the deposit.  The first is the presence of many thin quartz veins (0.5 to 40 mm) with




[exhibit1030.jpg]


Figure 9.1.  Plot of soil samples.  Color coding shows relative gold content with red indicating gold ≥0.100 g/t Au.  The black line encloses the area investigated by drilling to date.




visible gold and the second is the fact that the rocks are sodium-rich.  These aspects are under evaluation by ITH geologic staff.


Gold is strongly associated with arsenopyrite and locally with stibnite although stibnite is relatively rare.  Other metallic minerals include pyrite, pyrrhotite, and marcasite.  Some pyrite may be arsenian.  Small amounts of chalcopyrite and sphalerite are observed in thin section and locally in core.  Small amounts of molybdenite have been reported by previous workers.


Mineralization appears to be contiguous over a map area approximately 2.5 km2 and the 0.1 g/t grade shell averages 280m thick and ranges up to 510m thick.  On the south side of the Lillian Fault, individual mineralized envelopes are tabular and follow lithologic units, particularly the volcanics, or lie in envelopes that dip up to 45 degrees to the south and follow the structural architecture and dikes.  On the north side of the Lillian fault mineralization is similar in style and orientation, but more widespread and in steeply dipping Upper Sediments.  Interestingly, visible gold has been noted more often in Sunshine Zone mineralization north of the Lillian Fault.



9.2

Alteration


Rocks of Livengood have undergone multiple stages and styles of alteration.  As increased drilling reveals a wider range of subsurface material, complex overprinting and spatial relations for different stages of alteration are becoming apparent.  Four principle alteration styles are currently observed.  These are identified by each stage’s principal alteration mineral; biotite, albite, sericite, and carbonate.  Two other lesser styles of alteration also may be present.  Local zones of smectite-illite alteration and  local presence of possible minor pyrophyllite is curious and may be important, but convincing identification has not been made and it is unclear at this time where and how these minerals fits into the sequence.


Biotite alteration consists of fine-grained remnant patches of secondary biotite in sedimentary, volcanic, and dike rocks or as phlogopite (phlogopitic biotite?) in mafic and ultramafic rocks (Figure 9.2 and 9.3).  Pyrrhotite and quartz accompany the biotite.  Arsenopyrite may be in rocks with this type of alteration, but timing relations are not clear.  Macroscopically, the secondary biotite renders a weak to dark brown hue to the rock or margin to some veinlets.  All rock types have been affected by this stage of alteration, however, secondary biotite and accompanying pyrrhotite are observed only as remnant patches in local intervals in some drill holes where subsequent alteration stages have not obliterated it.


Albite alteration occurs as extensive replacement of volcanic and dike rocks and overprints biotite alteration.  Secondary albite occurs as intergrown radiating plumose to acicular sheaves and rosettes that locally replace all previous rock textures (Figure 9.2 and 9.3).  Albite is accompanied by intergrown fine-grained dark gray to black patches and grains of quartz.  This quartz is cryptocrystalline with an almost cherty character.  The dark color may be from included carbonaceous material (Sillitoe, 2009).  Albite alteration appears to be accompanied by disseminated arsenopyrite and pyrite mineralization.


Sericite alteration consists of pervasive sericitization, sericite veins, and quartz-sericite envelopes around quartz±sulfide veins in all rock types.  Sericite cross-cuts and/or replaces all previous alteration minerals, and locally appears to be developed from destruction of secondary biotite.  Pyrite and arsenopyrite accompany this stage, some of which may result from pyritization of biotite-stage pyrrhotite.  In mafic and ultramafic rocks, tremolite and local fuchsite are the dominant sericite-stage phyllosilicates.  In addition to the previously described black silica that accompanies albite alteration, fine-grained introduced quartz is widespread in many thin sections and replaces primary mineralogy.  However, this form of silica is rarely observed macroscopically due to other alteration minerals which are more readily apparent.  Sericite-stage silica also occurs as the inner zone of centimetre-scale alteration selvages around narrow fractures.


Smectite-illite alteration has been observed in a number of locations, generally in and around brittle fault zones, but is not as widespread as the albite and sericite alterations.  It is characterized by bleaching of the affected rocks and strong swelling and consequent disintegration of core samples from

[exhibit1031.jpg]


Figure 9.2.  Photomicrographs of characteristic alteration among rocks at Money Knob.  A) View of core showing relict patches of secondary biotite (dark color) cut by and overprinted by albite and sericite alteration. 08-33, 190.25  B) rare, relatively weakly altered Cretaceous intrusive dike with abundant interlocking plaigioclase laths and blocks; Weak sericite and carbonate alteration are present.  Some of the plagioclase may be in the early stages of being altered to secondary albite.  09-34, 252.76.  C and D) plane and polarized light examples of a patch of secondary biotite in Devonian volcanics; sericite and carbonate are also present in the lower right portion of the photo; 200x; 8-33; 190.25.  E and F) A quartz-carbonate veinlet crosscuts albitized volcanic rock (MK07-18, 247.5m).  G) Large arsenopyrite grain (A) with an inclusion of pyrrhotite (po), and adjacent to pyrite (py).  Minor chalcopyrite (cp) occurs in the lower right.  200x, 08-33, 230.55.



these zones.  The alteration has been observed most commonly in sediments and dikes.  Pyrite and arsenopyrite are disseminated through the alteration and gold grades of several hundreds of ppb are common.


Carbonate alteration consists of at least three styles of introduced carbonate:  1) clear but fine-grained scaly patches and flakes throughout the rocks; 2) fine-grained cloudy carbonate patches; and 3) clean large euhedral rhombs and clusters of rhombs in and adjacent to carbonate-quartz-sulfide veins.  Some very fine carbonate is brown in color.  It is not clear whether this is a natural color or a product of oxidation or overgrowth and incorporation of very fine secondary biotite.  Macroscopically, some brown carbonate has been mistaken for secondary biotite.  A fourth style of carbonate consists of very late calcite veinlets which crosscut all features.  These could be the product of late-stage cool hydrothermal alteration or supergene.  The vast majority of carbonate appears to overprint previous alteration stages, however, some may accompany the earlier alteration stages .  Carbonate abundance ranges from scattered flakes to complete replacement, particularly in the mafic and ultramafic rocks.  In the sedimentary rocks, it is difficult to determine if some carbonate is redistributed primary carbonate or introduced hydrothermal carbonate.  Local marl and limey beds occur in the Devonian sediments.  Carbonate apparently consists of dolomite and other Fe- Mg species of carbonate such as siderite and ankerite.  Arsenopyrite and pyrite are common in carbonate-quartz veins and veinlets.



9.3

Synthesis of Mineralization and Alteration


The types of alteration stages and their sequence are consistent with other IRGS deposits and prospects of the Tintina Gold Belt (Newberry and others, 1995; McCoy and others, 1997).  This is important as it strongly supports the interpretation that mineralization at Livengood is part of an intrusion-related mineralizing system.  Although it is possible that each alteration stage is the product of independent hydrothermal events, the mineralogy of each alteration type suggests that the various stages formed as part of an evolving, cooling system with initial biotite and pyrrhotite being the highest temperature and subsequent lower temperature assemblages following (Figure 9.4).  This patterning can also be interpreted as consistent with the chemical evolution of hydrothermal fluids emanating from an intrusive source.


Gold shows a strong correlation with arsenopyrite.  However, arsenopyrite has been introduced at least at the biotite alteration stage and significantly at the carbonate stage.  Some amount of arsenopyrite also may have been introduced at the albite and sericite alteration stages.  It is unclear, though, whether


[exhibit1032.jpg]

Figure 9.3.  Photomicrographs of characteristic alteration among rocks at Money Knob.  Plane light on the left; crossed polarized light on the right.  A and B) Sericite and carbonate replace a silty phyllite (MK07-18, 76.0m).  C and D) A quartz-carbonate veinlet crosscuts albitized volcanic rock (MK07-18, 247.5m).  E and F) Carbonate (upper left 2/3rds of section) and tremolite (lower right 1/3 of section) replace mafic rock.  25x; 02-21, 19.35.  G) Core showing a complex sequence of alteration types which generally mimic the larger scale assessment of alteration styles.  Zone 1 = secondary biotite-carbonate±sericite. Zone 2 = Carbonate-sericite with darker color possibly owing to overprinted secondary biotite.  Zone 3 = carbonate-sericite.  Zone 4 = sulfide-rich sericit e-carbonate.  Blue symbol = shear.  Orange dashed lines = bedding.  The yellow lines indicate quartz-carbonate±sulfides veinlets.  Red line indicates quartz-feldspar±carbonate veinlet.  From MK09-43, 388.3.



[exhibit1033.jpg]


Figure 9.4  Interpreted paragenetic sequence of key alteration and mineralization stages.  Gold occurs with arsenopyrite and may have been introduced during all stages or dominantly during a particular stage(s).



gold has been introduced during all of these stages or mostly during a particular stage.  Understanding these relationships is part of ITH’s current exploration program.


10.0

Exploration


10.1

Past Exploration


Several companies have explored the Livengood area as outlined in Section 6 (History).  That work identified a sizeable area of anomalous gold in soil samples and intervals of anomalous gold mineralization in drill holes (described in previous sections).


ITH advanced the soil sampling coverage in 2006 and 2007 by collecting an additional 361 samples.  These samples helped improve definition of anomalous gold in soil on the southwest side of Money Knob and between Money Knob and Radio Knob.


ITH undertook drilling of the surface geochemical anomalies in 2006 with favourable results.  In 2007, the area was drilled sufficiently to produce a resource evaluation (Giroux, 2007; Klipfel and Giroux, 2008a) and a program for 2008 was planned that would further that evaluation.  Drill results through September 27, 2008 were used as part of a revised resource evaluation in October, 2008 (Giroux, 2008; Klipfel and Giroux, 2008b).  Geochemical results received and drilling completed after that date were used for a subsequent resource update (Giroux, 2009; Klipfel and Giroux, 2009).  Results from 34 reverse circulation holes drilled in the winter of 2009 were primarily infill holes.  Data from these holes were applied to a new resource estimate which also incorporated advancements in modeling the deposit (indicator kriging) and resulted in upgrading and enlarging the resource estimate t o 4.04Moz and 3.6Moz in the indicated and inferred categories respectively (Klipfel, et al., 2009a).  Data from drill holes completed through September 2009 were used to complete a new resource estimate along with additional information on possible recovery techniques being contemplated by ITH (Klipfel, et al., 2009b).  The remaining data from drilling completed in 2009 is the subject of a new resource update and reported in this document.



10.2

Current Exploration


ITH has continued to conduct step-out and infill drilling throughout 2009.  This report includes the results from all results for 2009 drilling as received through February 28, 2010.  This data includes results from 195 RC holes and 12 diamond core holes drilled in 2009.  Approximately half of these holes were included in the resource estimate reported in October, 2009.  Results have been used in a new resource estimate reported in Section 17, and include further advances in metallurgical understanding and improved cost estimates which have been incorporated into the estimation process.


 

11.0

Drilling


11.1

Past Drilling


All of the companies that have explored at Livengood in the past, except Cambior, have undertaken drill programs to evaluate the district.  In each case, drill holes targeted different geologic concepts such as veins in bedrock beneath the alluvial gold.  AGA initially, and ITH later, focussed drilling on possible mineralization beneath and down dip from the surface soil anomaly area (Figure 11.1).


Drilling since 2003 by AGA and ITH is summarized in Table 11.1.  Drilling in 2003 by AGA consisted of 1,514 m of vertical and angled reverse circulation (RC) drilling in eight holes.  It identified broad zones of gold mineralization (BAF-7; Table 9.1).  Drilling in 2004 by AGA consisted of 654m of HQ coring in 4 diamond drill holes designed to test for gold beneath the thrust fault at the base of the Cambrian rocks.  These holes were up to 1.7 km to the west of 2003 drill holes.  They identified thick zones of gold mineralization in Devonian rocks beneath relatively barren, thrust-emplaced Cambrian rocks (MK-04-03; 96m@>0.5 g/t in 2 intersections).  These results highlighted the fact that significant mineralization could exist beyond the limits of the main soil anomaly, particularly in blind locations beneath thrust faults.


No drilling took place in 2005.


In 2006, ITH drilled 1,230m of core (HQ) in 8 holes and continued to demonstrate the presence of mineralization over a broader area.  The 2007 campaign consisted of 14 diamond drill holes for a total of 4,400m.  These holes focused on extending and defining the geologic setting of mineralization first recognized in MK-04-03.  This mineralization was originally thought to be hosted primarily in the Devonian volcanic rocks.  However, as drilling has progressed, it has become clear that mineralization is strongest in the volcanic rocks, but occurs in all rock types at Money Knob (Figure 11.2).


Based on favourable results in 2007, the 2008 program consisted of 30,653m of RC and core in 108 RC and 7 core holes.  These holes were designed to improve definition and expand the resource calculated early in 2008 based on 2007 drill data.  The 2008 drill program did not identify limits to mineralization in any direction.  Instead, a thicker mineralized zone was identified (up to 200m; Table 9.1).  In addition, this campaign highlighted the fact that mineralization occurs in all rock types, not just in Devonian volcanic rocks.  This was important as it indicated that there was potential for broader extent of mineralization than envisioned prior to the 2008 drill program.


The winter 2009 program helped fill in gaps in the drilling grid and enabled increased continuity of information for improved resource estimation.  In addition, more rigorous estimation procedures using indicator kriging, improved modeling of the oxidation profile, recoveries of various lithologic types, and cost estimates based on comparable pit mining techniques in this environment.




11. 2

Current Drilling


ITH drilled from June through October 2009.  The resource estimate along with other information presented in this report is based on all data generated through the end of 2009.  This includes the


[exhibit1035.jpg]



Figure 11.1  Distribution of drilling in the Money Knob area with respect to anomalous soil samples.  The majority of the soil geochemical target remains untested.



addition of data from 57 new RC drill holes and 7 diamond drill holes that was received after the end of September, 2009, the data cutoff date for the October, 2009 resource estimate.  Drill holes with incomplete assays and drill holes completed after September 25, 2009 are not included in this work.  These holes were drilled in locations to offer additional fill-in data, but more importantly, they expanded the area of known mineralization through identification and assessment of the new Sunshine Zone north of the Lillian Fault and about 600m east of the Core Zone.  Mineralization between these two zones is contiguous and open in all directions with the exception of a partial local margin formed by the Lillian Fault.

TABLE 11.1

SUMMARY OF AGA AND ITH DRILLING AT LIVENGOOD


Year

DDH

m

RC

m

Results

2003

-

-

8

1,514

Broad zones of Au mineralization

2004

4

654

-

-

Discovered Devonian volcanics as preferential host rock

2005

-

-

-

-

 No drilling

2006

8

1230

-

-

Drilled first >100gram meter intersection in Devonian volcanics

2007

14

4,400

-

-

Defined continuity of volcanics and mineralization. Discovered first sediment-hosted mineralization

2008

7

2,040

108

29,040

Discovered core zone where sericite alteration mineralizes all rock types. Delineated 6.8M oz indicated and inferred resource

2009

12

4572

195

59,757

Expanded the extent of the mineralization and provided input for the resource evaluation reported in this report and previous reports in 2009.



During the summer of 2009, 6 diamond drill holes were drilled across the NNW-trending Core Zone in order to better understand the structural controls and to test the depth continuity of the mineralization.  Three holes were drilled to the northeast and three were drilled to the southwest.  This drilling confirmed that the Core Zone is the locus of a swarm of 0.2 - 1.0m thick southerly dipping dikes.  In addition, a number of larger (≤10m thick) steeply dipping NNW-trending dikes were observed suggesting that ENE extension may have occurred at about the time of dike magmatism.  Quartz-arsenopyrite and quartz-stibnite veins were encountered and display SW dip and N-S, subvertical orientations respectively.  In general, it appears that the broad zones of mineralization in the core zone do not continue to depth, but narrower very high grade structures were encountered below the Core Z one and represent a potential future target.


As part of an overall step-out program, a NE-trending fence of 11 holes spaced at approximately 150 meters trending N45E and inclined -55 degrees was drilled.  The holes confirmed the stratigraphic continuity of the area but failed to delineate major zones of mineralization.  However, the farthest NE holes, encountered multi-gram mineralization associated with the faulted contact between the Devonian and the Cambrian mafic rocks.  This find is a step out of 500m from the nearest grid drilling.


In early 2010, ITH will be undertaking a drill program within areas that are sensitive to disturbance during the summer season.  This drilling will allow the infill of information in key areas not accessible for drilling in the summer.  The goal is to test areas further to the southwest, southeast, and northeast.  If successful, this drilling will expand known mineralization further in three of four directions.




[exhibit1037.jpg]


Figure 11.2  Distribution of drilling in the Money Knob area according to year and company.



11.3

Drill Procedures


To date, virtually all drill holes at Money Knob have been drilled in a northerly direction at an inclination of -50 degrees in order to best intercept the south dipping structures and mineralized zones as close to perpendicular as possible.  A few holes have been drilled in other directions as described above.  Most holes have been spaced at 75m along lines 75m apart.  A few holes are more closely spaced.  Surveys of the holes show that with depth, holes steepen 10-20 degrees depending upon the length.  Most holes have been drilled to depths of 250-300m.


Diamond drill core is recovered using triple tube techniques to ensure good recovery and confidence in core orientation.  Recovery is excellent being greater than 95% over the course of the entire program.  The core is oriented using the ACT system and/or the EZ Mark tool.  Core is marked so that a continuous line is located along the base of the core as long as core pieces can be matched continuously from the marked top of the run.  Subsequent runs are matched also.  Oriented core is important for recovery of structural, vein, and contact orientation information and is essential for interpreting fault and dike orientations on sections.


Currently, core is reviewed, marked for orientation, and ‘quick-logged’ at the drill rig prior to transporting it to ITH’s core shed for logging.  This is a relatively new procedure and has been in place since August, 2009.  In the past, core would be marked for orientation and then placed as an entire run in a case of prepared PVC pipe and sealed until opened by core loggers at ITH’s core shed.  This custom procedure was implemented to assure minimal breakage or crumbling of core between retrieval from the hole and transfer to boxes by the logging geologist.  Core is cleaned, measured, marked, labelled, and logged by contract geologists from Northern Associates, Inc.


Reverse circulation holes are bored and cased for the upper 0-30m to prevent downhole contamination and to help keep the hole open for ease of drilling at greater depths.  Recovery of sample material from RC holes is done via a cyclone and dry or wet splitter according to conditions.  Sample chips are split into 3 recovery points (Figure 11.3): one is the interval sample, the second is an equivalent split “met” sample, and a third smaller split is used to collect chips for logging purposes.  These chips are placed in standard chip trays.  Samples are collected in porous polybags that allow retention of sample material and evaporative seepage of water from the sample.


Drill hole locations are determined by sub-meter differential GPS surveys at the drill collar.  Initial azimuth of drill hole collars are measured using a tripod mounted transit compass in conjunction with a laser alignment device mounted on the hole collar (Figure 11.3).


Down hole surveys of core and reverse circulation drill holes are completed using the Gyro-Shot survey instrument manufactured by Icefield Tools Corporation.  Precision and accuracy of this method was assessed in 2008 through a series of duplicate surveys using this instrument and by comparison in holes surveyed by the EZ-Shot (magnetic) borehole surveying device.  Results of surveys and duplicate tests show normal minor deviation in azimuth and inclination with reproducibility within a close margin of error.  In 2009, a duplicate survey performed by the Gyro-Shot instrument measuring the same hole twice (MK-RC-0195 to 985 feet) and a tandem survey performed by running two Gyro-

Shot instruments simultaneously on the same probe assembly (MK-RC-0178 to 900 feet), demonstrated close replication and agreement between the surveys.  The 3-D coordinates at the maximum depth of the paired surveys plot to within 1% of the coordinates in the corresponding survey relative to length of hole surveyed.  Drill hole surveys were completed by Northern Associates, Inc. and were observed in the field by Dr. Klipfel.


The RC drilling in 2003 was conducted by Layne Christiansen Company using an MPD 1500 Track RC drill.  Drilling in 2004 was also by Layne using a CS1000 core drill.  No drilling took place in 2005.  In 2006, 2007, 2008, and 2009, diamond core drilling was conducted by AK Drilling Inc, and Layne Christensen.  RC drilling was by AK Drilling, Inc., and T and J Enterprises.



[exhibit1038.jpg]

Figure 11.3  Photos of various exploration functions.  A) ITH geologist logging RC chips with a binocular microscope.  B) View of ITH’S core shed and core boxes in the foreground.  C) Driller taping core securely in PVC holder/carrier.  Core barrel parts are on the left.  D) RC drilling chips are split into 3 collection points, the sample (foreground bucket), the met sample (background bucket), and the visual chip sieve for logging purposes (left).  E) A representative sample of RC chips is retained in chip trays with individual compartments for each 5’ interval.  F) Drill hole collars are surveyed with a differential GPS instrument.  G) The driller marks the core to indicate its oriented position with respect to the core barrel.  H) Drill core is sawed in half with a diamond saw at the core s hed.  I) The driller marks a line along the base of the core to indicate its oriented position.  J) Niton portable XRF instrument records trace-element abundances prior to shipment of samples to the lab.  K) Trace elements are measured by two NITON portable XRF instruments for all RC samples prior to shipment to the lab for assay and multi-element ICP analyses.  L) Example of porous polybag which allows the escape of water, but not sample material.  Pre-printed labels indicate drill hole, depth interval, sample number, and bar-coded sample ID information.


 

12.0

Sampling Method and Approach


12.1

Past Sampling


The sampling procedures of previous companies are not known but the major companies that did the work are known for their conscientious QA/QC protocols.  Sample data from past programs are consistent with more recent data generated by AGA and ITH.  On this basis, there is no reason to doubt the validity or credibility of samples from Occidental, AMAX, Homestake, or Placer Dome.  The similarity of results for each program suggests that sample collection and analytical procedures are sufficiently similar to allow use of their data by ITH in current exploration efforts.


For samples collected by AGA, all soil, stream sediment, rock, and drill sampling was done according to AGA in-house protocols for geochemical sampling.  These protocols specified technical procedures for collection and documentation of samples.  In general, -80 and -200 mesh material was analyzed for soils and stream sediment respectively.  Dr. Klipfel reviewed these protocols in 2006 as well as AGA’s security procedures and verified that they met or exceeded standard industry practices.  Sampling procedures remained the same through the course of the 2003 and 2004 exploration programs.


All AGA geochemical samples were secured and shipped to Fairbanks according to AGA protocols for sample preparation (drying, crushing, sieving, and pulverizing) at ALS-Chemex in 2003 and Alaska Assay in 2004.  Sample splits (300-500g for rock material; -80 mesh for soil samples) were sent to ALS Chemex in Vancouver for analysis.  Analytical methods used were standard 50g fire assay with AA finish and four-acid digestion, multi-element ICP-MS.  These are standard analytical packages for the exploration industry and are performed to a high standard.  Analytical accuracy and precision were monitored by the analysis of reagent blanks, reference material and replicate samples.  Quality control was further assured by the use of international and in-house standards.  ALS Chemex is accredited by the Standards Council of Canada, NATA (Australia) and also has ISO 17025 and 9001 accreditation .


AGA reverse circulation drill samples were collected at five foot intervals as measured by the driller.  Pulverized material from the hole was passed though a cyclone to separate the solids from the drilling fluid and then over a spinning conical splitter.  The splitter was set to collect two identical splits each of which weighed 2-5 kg.  Representative material was also collected and saved in chip trays for later visual inspection.  The split material was put into pre-numbered bags by the drillers’ helpers on site.  One of the splits was sent for analysis while the other was retained for future reference.  Samples were secured and transported to the sample preparation facility of ALS Chemex in Fairbanks for drying, crushing, pulverization, and splitting.  120 gram splits were sent to Vancouver for analysis by standard 50 gm fire assay with AA finish and multi-element IC P-MS.  The RC chips were logged by project geologists by recording basic information on the lithology, alteration, and mineralization for each interval.


AGA’s core material was collected at the drill site and placed in core boxes under the supervision of an experienced geologist and Qualified Person for the purposes of NI 43-101.  It was logged for rock type, alteration, structure, and with detailed descriptions.  Dr. Klipfel examined the core logs and core from the four 2004 holes and can verify the reliability of the logging.  Sample intervals were determined on the basis of the distribution of veining and alteration with a minimum sample width of 30 cm and the maximum width of 1.5m.  Samples were collected to isolate different components of the alteration and mineralization to characterize them.


After the samples were marked, the core was sawed in half, and one half sent for analysis.  The other half was either kept on site or at AGA’s core storage facility in Fairbanks.  The average recovery in the core program was in excess of 90% and there is no indication that poor recovery is an issue in the interpretation of the assay data.  Sampling was selective but barren samples were always collected to bracket zones of mineralization so that reliable boundaries could be defined in the intercepts.  Dr. Klipfel examined this core in the course of the site visits.



12.2

Current Sampling


ITH has adopted and continued the sampling protocols used by AGA and described in the previous section, with the exception that all drill holes are sampled from surface to total depth.  In addition, ITH has implemented a number of customized steps in their procedures to minimize errors and assure the integrity of sample material.  This assures a high level of reliability in the sample data set and assures continuity of methodology, laboratory standards and conventions as well as confidence in the data generated.  All core samples are weighed prior to shipping to the ALS-Chemex facility in Fairbanks.  These weights are compared to the laboratory received weights to confirm that the samples were logged in correctly.  RC samples are collected in pre- numbered, bar-coded bags (Figure 11.3).  They are logged-in on-site by ITH using the barcodes to prepare the shipments and ALS Ch emex uses the same barcodes to log the samples into their system.  The sample weights are recorded at various stages in the preparation process.  These procedures minimize labelling and other potential errors and add an extra level of assurance that the sample is tracked correctly and matched with the data generated by that sample.


Currently, core is examined by the logging geologist in the original split tube before being boxed and transported to the core shed for logging, mark-up and sampling.  For the 2009 program, core was slid from the core barrel into a half-section of PVC pipe, covered with the other half of PVC pipe, and sealed for transport to the logging shed at ITH’s camp (Figure 11.3).  This procedure is effective and minimizes disturbance to the core, prevents unnecessary breakage, and minimizes crumbling of core prior to logging by a geologist.



13.0

Sample Preparation, Analyses and Security


13.1

Past Procedures


Soil and drill samples obtained in 2003 and 2004 exploration programs were subject to AGA’s in-house methodology and Quality Assurance/Quality Control (QA/QC) protocols.  Samples were analyzed by various methods by different laboratories.


The QA/QC program implemented by AGA met or exceeded industry standards.  The program involved analysis of blanks, standards and duplicates.  Blanks help assess the presence of any contamination that might be introduced by analytical equipment.  Standards are used to assess the accuracy of the analyses, and duplicates help assess the reproducibility or precision of the analytical methods and equipment used.


All sampling campaigns were subject to insertion of blanks and standards at a rate of 1 blank and 1 standard for every 23 samples (total = 2QA/QC samples per 25 submitted samples).  Blank samples consist of material known to contain below detection amounts of the metal for which the sample is being tested.  Standards consist of sealed sachets of material with a certified abundance of the metal for which the sample is being tested.  Standards were purchased from RockLabs and GeoStats.


Duplicate core and rock samples were run from pulp and coarse reject splits along with sample repeats approximately every 20 samples.  Duplicate samples were also collected at the drill rig for 2003 RC drilling.  Results of AGA’s QA/QC program were reviewed by Dr. Klipfel in 2006 and in subsequent visits and reports.  Overall, the QA/QC samples indicate that sampling and analytical work is accurate and reliable.  In 2004, there were two instances of issues with blanks and standards out of compliance with AGA protocols, but these were satisfactorily resolved by AGA.  The sample database did not appear to be compromised.



13.2

Current Procedures


ITH has continued with the QA/QC protocol of AGA as described above and increased the number of control samples (blanks and standards) to 1 in 10.  Duplicate splits of drill samples are prepared for every 20 samples.  ITH has undertaken rigorous protocols to assure accurate and precise results.  Among other efforts, weights are tracked throughout the various steps performed in the laboratory to assure accurate assignment of results to the appropriate sample (Figure 13.1).  ITH weighs all core samples before shipping.  They are then reweighed by the laboratory when received and logged in.  RC samples are dried and then weighed at the laboratory.  Sample reject material is weighed again by the laboratory after the sample aliquot has been removed for pulverization.  This tracking of sample weights enables constant verification of quality throughout the preparation pro cess.  Key results of this protocol include minimization of sample switches and transcription errors.


All core and RC samples are taken from the drill rig directly to ITH’s core shed.  RC and core samples are placed in super sacks, sealed, and palleted for shipment to ALS Minerals’ preparation facility in Fairbanks.


Samples are analyzed by standard 50g fire assay for the gold determinations.  All core samples and select RC drilling samples are also submitted for multi-element ICP-MS analyses using a 4 acid digestion technique.  All RC samples are analyzed on site for trace elements using a Thermo Fisher Scientific NITON portable XRF before shipment to the laboratory (Figure 11.3).


ITH geologic staff has developed a set of decision criteria that compare the NITON-measured abundance of Cr, Ni, Th, Zr, Mo, and V for determination of ultramafic, volcanic, Cretaceous intrusive (dikes), Upper Sediment, and Lower Sediment rocks.  These results are cross checked with visual logging and ICP data before a final lithologic determination is entered in the database.  The advantage of this type of procedure is that rock types can be more readily and more consistently identified in spite of significant alteration and replacement of original rock textures and minerals.  Also, because arsenic correlates strongly with gold, an XRF determination of arsenic abundance has helped ITH anticipate gold-bearing zones before assays are returned.  This information has proved constructive for drill planning and execution.



13.3

Data Handling


A project database is maintained by ITH with all drill hole location, survey, logging, sample, and assay information contained therein.  As drill holes are completed, data is entered either manually, or through data downloads directly from instruments to the database.  Assay information is received electronically from the laboratory and downloaded into the database.  Subroutines check for errors and data format consistency.


The creation of sample data for RC drilling begins with pre-numbered sample bags that have drill hole number, sample interval, and sample number printed and bar-coded on a label attached to the bag (Figure 11.3).  These bags are used at the drill rig for collection of RC chips into a primary sample, a secondary duplicate sample, and a chip sample for logging purposes (Figure 11.3).  Drill core is sawed in half with a diamond saw with half the core going in a sample bag together with a tear off sample ticket preprinted with the sample number, and the other half retained in core boxes and stored on site.


NITON data collected by the instrument is keyed to the sample number so that data transferred from the NITON “gun” to the database remains matched with the sample number.  Chip loggers similarly enter information into the logging database while reviewing chips under a binocular microscope with all intervals keyed to the sample interval and sample number (Figure 11.3).  These are checked regularly by loggers and rechecked by the senior geologist.  Database check and validation tools are also used to detect errors.  Core logs are created manually and then the information is entered into a digital format for the database.


Dr. Klipfel has reviewed these procedures and watched the data entry process at various steps at different times on each of the visits.  He is satisfied that ITH is diligent in their data management procedures and have check procedures in place that should identify any issues.  He has not completed a thorough check or validation of the database but is not aware of any issues.




[exhibit1039.jpg]


Figure 13.1.  This diagram shows the flow path and steps involved for RC samples from the drill rig to analytical results.

13.4

Quality Assurance and Quality Control


The QA/QC data from ITH sampling program has been reviewed by Dr. Klipfel.  Analyses of blanks and standards that fall outside of an acceptable range, such as 3x detection limits for blanks or 10% for standards, are flagged for investigation.  Unless a suitable explanation, such as a sample switch, can be found, the error is reported to the laboratory and the sample intervals around the questionable sample are rerun.  A new certificate is issued by the lab for the reanalysis if the correct values for the standards and blanks are determined.  Errors are generally attributable to sample switches, weighing errors and contamination of the first sample in a batch.  Multi-element QA/QC is monitored using the compositions of the blank and standard materials.


Duplicate samples are used to assess reproducibility of the laboratory procedures and to ensure that the sampling procedure is representative.  Field duplicates (63 in 2009) represent equivalent samples collected at the drill rig during the original sampling process and confirm that the sampling process is representative (Figure 13.2a).  Prep duplicates (1527 in 2009) are prepared by splitting the whole sample in half at the laboratory and subjecting each half to the full sample preparation routine and subsequent analysis (Figure 13.1).  These duplicates are designed to assess sample homogeneity and confirm that no bias is created during the sample preparation process (Figure 13.2b and c).  Pulp duplicates (388 in 2009), representing multiple assays of the same pulverized material show that the laboratory procedures are precise and that the pulp material is uniform with errors of mostly less than 10% (Figure 13.3).  Errors greater than 10% are believed to be due to normal nugget effect typical of gold deposits.


As the number of samples increases with each drilling campaign, it appears that there are local variations in the scale of nugget effect.  The result is that some duplicates at higher values of gold (e.g. >3 g/t Au) show higher variance in reproducibility.  Dr. Klipfel has evaluated this issue carefully and believes it is the result of normal nugget effect where a grain of relatively coarse gold ends up in one split and not the other, thus producing a high value in one run and a lower value in another.  This can be tested by comparing the blanks and standards for that range of samples and verify that these values are accurate and precise (Figure 13.3).  Also, reproducibility tends to improve as gold values decrease except as the detection limit is approached (e.g. 0.005 vs 0.01 g/t = 100% error, but is at the detection limit and normal error envelope).  This is most likely due to more even distribution of smaller gold grains so that an equal number of fine grains end up in each sample split.  This level of variation due to nugget effect is deemed unlikely to impact the data set or the resource evaluation, because for each instance of a value in one sample being higher than in its paired duplicate, there should be an equal number of lower values recorded which missed the higher value split.


Prep duplicates (1527 in 2009, 736 in 2008, 187 in 2007), created by splitting either core samples after coarse crushing or splitting raw RC chips, show a somewhat higher degree of variability but demonstrate no bias to either high or low grade (r=0.91, Mean original samples = 0.43g/t, Mean of duplicates=0.45g/t).  The reproducibility of most pulp duplicates also indicates that most of the gold is not so coarse that it causes major nugget effects.  The variability in the coarse duplicates indicates that gold grains are not uniformly distributed within the sample material.  This is consistent with the interpretation that gold is, at least partially, hosted in narrow veins and veinlets, which when crushed produce a small number of gold-bearing fragments in the overall sample, thereby causing nugget effect during the coarse sample splitting.  In recognition of this effect sample preparation pro cedures were

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Figure 13.2.  These scattergram plots show how different categories of sample duplicates compare with original sample results.  The diagonal line has a slope of 1.  Perfect duplication of results would plot on this line.  Variation and scatter is interpreted to be the product of normal nugget effect.  A) 2009 field duplicate vs. original samples; n= 27.  The envelope of points flares with increasing grade.  This is typical of nugget effect which becomes more pronounced at higher grades.  B) 2009 prep duplicates compared to original sample values.  The scatter indicates no particular bias with a good overall correlation between the two sets.  The scatter is believed to reflect normal nugget effect in these samples..  C) 2009 pulp duplicates vs. original sample.  Scatter is similar to that in B.



modified so that 1kg of sample material is now pulverized rather than 350g aliquot previously used.  Dr. Klipfel considers these results to be appropriate for Livengood mineralization and indicative of sound QA/QC procedures.


The ALS Minerals preparation facility in Fairbanks has been visited twice by Dr. Klipfel to verify sample handling procedures.  The lab follows sound log-in, weighing, drying, and splitting procedures.  Sample crushing, splitting, and pulverization is done by modern equipment with diligent air cleaning

[exhibit1041.jpg]


Figure 13.3.  X-Y scattergrams for 2009 showing the stated value of standards vs. the measured value by the lab.  A) values are plotted according to measured value vs. stated values of a standard placed in the sample stream. B) values plotted as a function of time to check for drift in results over time.  The horizontal nature of the points for each value indicates that drift is minimal to non-existent.


between samples and cleaning with blank material between runs and at the beginning of the day.  Handling techniques demonstrate care in assuring that bags and samples are not mixed up.  All pulps are sealed in paper envelopes and placed in boxes, packaged and sealed for transport to the Vancouver or Reno labs for analysis.



14.0

Data Verification


Field and drill core observations made by Dr. Klipfel during site visits are consistent with the style of mineralization and alteration interpreted and reported in ITH documents.  Outcrop exposures in drainages, trench faces, road cuts, and along the ridge lines were examined and found to be consistent with existing geological maps.


Drill logs, sections and maps were reviewed and are to a high quality.  Provided information is consistent with observations of core and surface exposures.


In 2006, Dr. Klipfel collected a single sample along 3 m of a trench face where intrusive material with quartz veins is exposed.  This sample was crushed, split, pulverized and assayed with a 50 g fire-assay AA finish method by ALS Chemex in Reno, Nevada.  The sample contains 1.31 g/t Au, a value consistent with results from AGA sampling and expectations for material of that type and location.  In addition, Dr. Klipfel witnessed the sluicing and panning of concentrated “clean up” material shovelled from a trench face.  The material contained a significant amount of fine colors as seen in the panning dish verifying the presence of free gold at a range of sizes in that part of the trench face.


In 2007, Dr. Klipfel collected seven samples from portions of two different drill holes, MK-07-18 and MK-07-20, from the remaining half of drill core previously sampled by ITH.  Samples were selected for a range of gold content and rock type.  The range of gold content in these samples is from below detection to 16.8 g/t Au.  The core was quartered for the same sample interval as previously collected by ITH.  Core material was bagged, labelled and information recorded by Dr. Klipfel and by ITH staff.  Sample bags were sealed and transported to the ALS-Chemex laboratory in Fairbanks for sample preparation.  Pulverized material was split into 300 gram master pulps and 120 gram analytical pulps before being sent to ALS Chemex in Vancouver for analysis.  All samples except one returned results reasonably consistent with results from the ITH original sampling.  The single samp le that is different contains 0.61 g/t Au compared to 6.92 g/t Au in the original ITH analysis.  This discrepancy is similar to the few discrepancies that occur in ITH’s QA/QC sample duplication procedures.  For this reason, the discrepancy is interpreted to reflect normal variation attributable to nugget effect as described in section 13.2.  To the extent that this type of error is throughout the database, it is equally likely that a corresponding number of samples report low when the other half of core might report higher.


In 2008, 31 samples (26 RC and 5 core) were collected by Dr. Klipfel for verification analyses.  These samples came from 5 different RC holes and 1 core hole.  Samples were selected at random and specifically for a range of gold content from near detection limits (0.005 g/t Au) to high grade (20.9 g/t Au).  Half-core that remains after a first sample was quartered and analyzed.  Two standard and two duplicate samples demonstrated good reproducibility.  RC samples demonstrated reasonable reproducibility, and core samples showed a range.  No systematic bias was observed.  Dr. Klipfel interprets these results to show normal scatter and nugget effect typical of mineralization at Livengood and for gold in general.


As a check of the data generated during 2009, Dr. Klipfel selected two batches of samples.  The first batch consisted of 28 samples selected from the duplicates collected by ITH from the winter program.  The second batch consists of 13 duplicate RC chip samples randomly selected at each of the three RC drill rigs.  Samples of the first batch were selected to be representative of a range of rock type and gold values from different holes.


Results for the first batch show very good accuracy and precision for the standard and blank samples included with the sample set.  The duplicate sample shows variation (2.13 vs. 2.89) of about 25%.  Five other samples within this batch show significant variation between the original and duplicate analysis.  For this reason, both the original and duplicate samples were re-analyzed.  The values from these four runs show consistent variation among samples with higher gold values (e.g. 1 or more runs with higher values) for at least one run out of the four runs (Figure 14.1).  It also shows minimal variation among samples with very low gold content.  Importantly, samples with minimal or no gold (≤0.1 g/t Au) show consistency and repeatability.  When plotted in log-log format, the envelope of variation becomes smooth, again suggesting a natural nugget effect.  Th is assumes that the gold at Money Knob is consistent with the concept that natural systems follow logarithmic abundance patterns (Levinson, 1974; Rose and others, 1979).


Results for the second batch show good correlation and do not display any discernible bias (Figure 14.2).  Deviation from an ideal 1:1 correlation is consistent with past sampling and the degree of nugget effect observed throughout the course of ITH’s drilling program.


Data from duplicates for drilling have been reviewed by Dr. Klipfel and conform to previous QA/QC assessments.


Dr. Klipfel has not verified all sample types or material reported.  To the best of his knowledge, ITH has been diligent in their sampling procedures and efforts to maintain accurate and reliable results.



[exhibit1042.jpg]


Figure 14.1  X-Y scatter plots of 2008 and early 2009 original and duplicate sample data for check samples collected by Dr. Klipfel as part of data validation procedures.  The diagrams on the left are plotted with numeric scales.  The diagrams on the right are plotted with log-log scales.  The scatter increases with grade on diagrams with numeric scales while the envelope of points remains approximately parallel to the “unity” line.  This is consistent with data following lognormal abundance pattern typical of natural elemental abundance patterns.  A and B) original vs. “met” splits.  C and D) original vs. duplicate original splits.  E and F) original vs. duplicate “met” sample.  G and H) met and duplicate met samples.  These diagrams collectivel y indicate a lack of consistent bias and show that different splits show variation consistent with nugget effect at all grades, but more pronounced at higher grades.






[exhibit1043.jpg]


Figure 14.2  X-Y scatter plot of original and check samples for June 2009 RC drilling.  The correlation line shows a slope of 1.  Samples with identical results will plot on the line.  Deviation of results from the line is interpreted to be the result of normal variation and nugget effect.


 

15.0

Adjacent Properties


Another claim block called the Shorty Creek claims is controlled by Select Resources Corporation, Inc. and is located approximately 10 km to the SW of the Livengood project area.  This area is actively being explored for gold mineralization by Select Resources.


The Alaska Pipeline, the main means of transporting crude oil from Alaska’s North Slope to the south coast of Alaska, runs northwest-southeast about 6 km to the west.  This feature is not expected to have any impact on the project.


 

16.0

Mineral Processing and Metallurgical Testing


16.1 Introduction


ITH has undertaken metallurgical and processing test work to determine optimal recoveries using some combination of heap leach, mill with Carbon in Leach (CIL) and gravity or flotation separation techniques.  Current test work focuses on determining the best means of optimizing these combined recovery methods.  This work involves studies that evaluate how the ore is characterized and how the ores vary in their physical and metallurgical response to process treatment parameters by ore type according to the various lithologic units that host ore.  The characteristics under review include grindability, abrasiveness, optimal particle size for downstream treatment, and response to leach or gravity process parameters as a function of oxidation and lithology.  In addition, the right combination of these techniques for different ore types is being evaluated.


The information presented here derives from on-going studies which are in progress.  In the previous October, 2009 report (Klipfel, et al., 2009b), results from leach tests were applied to ore that is amenable to heap leach processing.  Although ITH has envisioned that Livengood gold would be recovered through a combination of processes, test work for a mill with Carbon in Leach (CIL) and gravity or flotation techniques had not been completed at that time and was not used in the Whittle Pit estimation.  Test work continues and is still in progress.  Results received since then are presented here but are not final.  On-going work will support future evaluations.  

Important metallurgical findings include:


      • variable metallurgy (chemical and physical properties), depending upon ore type; degree of oxidation, amount of organic carbon, etc.;
      • identification of ore types that are highly amenable to simple cyanide leaching process techniques like heap leaching with a carbon in column adsorption plant (CIC), particularly oxidized and partially oxidized ore types;
      • identification of some sediment-hosted ore that contains organic “preg-robbing” carbon that will require CIL techniques, gravity or flotation techniques;
      • high recoveries for some ore types using gravity separation techniques.


Test work completed or currently in progress includes grindability, abrasiveness, optimal particle size for downstream treatment, and response to flotation or gravity concentration followed by cyanide leaching of the concentrates as a function of oxidation and lithology.  Power requirements and reagent consumptions for each ore type for each process scenario are also being developed from test work data.  This information will be used as inputs into process operating costs for each ore type in future estimates.  



16.2 Metallurgy Summary


Metallurgical test work programs on the Livengood ores began in 2004 and continue as of the preparation of this report.  The ore types at Livengood are variable in their chemistry, in their physical properties, and in their metallurgical characteristics.  The following statements best describe the observed results of the test work performed to date:

      • Most Livengood ores could be considered moderately soft to medium hard in hardness with an average Bond Ball Work index of 15.8.  The ores varied significantly in hardness, with Bond Ball Work indices varying from a minimum of 11.1 to a maximum of 19.1.
      • The majority of the ores would be considered non-abrasive, with an average Abrasion index of 0.0809.  The ore types abrasion characteristics varied significantly from 0.0023 to 0.2872.
      • All of the Livengood ore types respond to cyanide leaching to some degree.
      • Some of the unoxidized ore with organic carbon has “active” or “preg-robbing” carbon.
      • The effect of leach times on gold recovery and gravity concentration results indicates some of the ores contain coarse gold.
      • Gold recovery at 10 mesh particle sizes on some of the ore types exceeded 90 percent.
      • Gold recovery on some of the ore types, but not all, is improved with finer grinding.  A grind size where 80 percent of the particles are smaller than (p80) 200 mesh (74 microns) has been tested to date.
      • The degree of oxidation of the ore, as observed by the geologists, has a marginal impact on the gold recovery.
      • Differences in gold recovery between cyanide shake leach tests, bottle roll leach tests, and Carbon-in-Leach tests suggest organic carbon in the ores is active to varying degrees in some of the ore types, particularly the un-oxidized version of those ore types.
      • The gold in the ore is often associated with sulfides, but the ore would not be classified as a sulfide refractory ore.

These results indicate that some of the ores are amenable to conventional heap leaching and gravity separation recovery processes, while others present more challenging metallurgical issues.


It became evident early in the test work that the oxidized to partially oxidized ores responded well to cyanide leaching while other un-oxidized ore types performed moderately to poorly depending on the method used to perform the analysis, i.e. cyanide shake leach tests versus bottle roll tests. However, it was found that all of the ores do respond to cyanidation to some degree.


The most significant metallurgical parameter observed to date with the Livengood un-oxidized ores is the presence of organic carbon and the indication that some, but not all, of the organic carbon is “active” or “preg robbing” in nature.  Metallurgical test work began to focus on process methods that could be used to counter the preg-robbing effects of the ores.  The simplest of these methods, the Carbon-in-Leach (CIL) process, has been the focus of test work since October, 2009, and is currently being used in test work performed in the latest round of tests.  The CIL test work, although incomplete, is showing positive results in counteracting the effects of preg-robbing carbon, providing an average increase of gold recovery compared to standard cyanide leaching for all ore types of approximately 18 percent and as high as a 49.5 percent increase in gold recovery for the more d ifficult un-oxidized ores.


In addition, gravity concentration testing of the Livengood ores continues to show encouraging results with 58% of the gold reporting to the gravity concentrates.  The results show a 69-1 concentration ratio (gravity concentrate weight percent of 1.43%) provides an average concentrate grade of 46.1 g/t Au.  Test work on the gravity concentrate to establish the viability of ultra-fine grinding and high intensity cyanide leaching of the concentrate indicate that total recoveries of gold in the main rock units range between 83% and 92%.  Lower total recoveries were observed in the unoxidized KINT units (59%-62%), however, these rocks form a minor portion of the total ore.


From the data currently at hand, the oxide and partially oxidized ores will respond well to heap leaching.  Ongoing test work indicates higher gold recoveries can be obtained from all ore types and particularly the weakly to un-oxidized ores with the use of standard milling that utilizes an initial gravity circuit followed by a Carbon-In-Leach (CIL) process for the gravity circuit tails.  The ability to increase recoveries from the higher grade mineralized zones as well as effectively process the weakly-oxidized to un-oxidized ores has the potential to significantly improve the Livengood project in both its size and economic performance.


Metallurgical test work currently underway and/or planned and scheduled for the future will continue to focus on utilizing CIL as a primary metallurgical process.  Enhancing the CIL test work with tests that attempt to render preg-robbing organic carbon inactive will also be performed.


Initial batch flotation test work has been performed to determine the potential for concentrating the gold and depressing gold preg robbers prior to downstream cyanidation.  In these tests, flotation was followed by gravity recoverable gold tests.  The test results indicated that flotation would recover between 52%-74% of the total gold, and that gravity recovery on the flotation tails would recover an additional 15%-59% of the gold.  Total gold recovery was relatively high and ranged between 83%-93%.  Further testing is ongoing to evaluate the CIL recovery from the flotation and gravity concentrates.


Column leach test work is being performed at a ½ inch crush on ores that do not show preg robbing tendencies to establish the effectiveness of heap leaching as a process option. Column tests are still in process.


Other future test work will include enhancing the ability of utilizing the gravity susceptible component of the ores for improving overall gold recovery.


This test work will be performed in conjunction with enhancing the ability to identify the ores that are subject to having preg robbing issues.  Understanding the geology of the ore types with respect to preg robbing organic carbon will be an important task moving forward in the Livengood project.



16.3 Gold Characterization


Hazen Research, Inc. performed gold characterization work on products they prepared from a heavy liquid separation test program performed on Livengood samples during late 2006 and early 2007 (Hazen Research Inc. letter report dated February 7, 2007, Subject: Characterization of Livengood Gold Ore, Hazen Project 10504).


The samples were ground to minus 35 mesh for gravity separation.  The minus 35-mesh material was first wet-screened at 500 mesh (25 μm).  The minus 35- plus 500-mesh product was split in half, and each half was separated with heavy liquid at a density of 2.96 to upgrade the heavy minerals plus the gold to enhance detection of the gold.  The float (tailings), sink (concentrate), and the unseparated minus 500-mesh slimes from one set of heavy-liquid separation were fire assayed for gold and silver.  The products from the other set were used for the mineralogical examination.  To concentrate the gold even further, the sink product and the minus 500-mesh slimes were panned.


The test showed 4% to 10% of the sample mass reported to the heavy mineral concentrate, which contained between 44% and 77% of the gold.  Another 13% to 33% of the gold reported to the minus 500 mesh slime fraction with the balance reporting to the +35 mesh float fraction.  Silver values in the ore were essentially negligible and the silver did not report to the heavy mineral concentrate with the gold.  Microprobe analysis of one gold grain indicated that the silver content was 7.4%.  The balance of the silver was probably held in other sulphide phases.


The main sulphide minerals in the heavy mineral concentrates were pyrite and arsenopyrite in ratios ranging from 2:1 to 6:1.  Pyrrhotite and chalcopyrite were commonly observed as inclusions in both pyrite and arsenopyrite.  Pyrite may be euhedral or anhedral and was frequently porous, enclosing abundant inclusions of gangue and rutile.  Sphalerite tended to occur as liberated grains or intergrowths with pyrite and arsenopyrite rather than as inclusions.  Trace amounts of several other sulphide minerals and gold were also present.  Hematite was observed in the only partially oxidized sample examined.  Marcasite was reported in some samples also, and in one of these it occurred as distinct clusters of acicular crystals and was possibly a product of oxidation.


Gold occurrences were scarce.  The size of the gold varied between less than 5 and 23 μm.  The particles observed were mostly associated with arsenopyrite as small attachments or inclusions, and one liberated particle was found in the minus 500-mesh product of the partially oxidized volcanic-hosted sample.



16.4 Historical Test Work Programs


In 2004, AGA attempted to test the cyanide solubility of gold in drill sample material by analyzing samples containing more than 200 ppb Au.  Samples were sent to ALS Chemex for a 30g cold cyanide leach assay (Au-AA24).  A total of 198 samples were analyzed in this manner and they showed consistent CN soluble assays, on average about 60% of the fire assay value (AGA in house memorandum to files).  The significance of this result was unclear at the time because there were many variables which could affect this outcome.  These included small sample size, nugget effect, host rock type, sulphide content, other mineral content, encapsulation, and possible inappropriate testing method.  Of these, nugget effect is expected when there is coarse free gold which was witnessed by Dr. Klipfel in the sluice sample of trench face material and has been seen in drill core.  Sulphide and organic ca rbon are present and also could be significant factors.  In an effort to determine which minerals might impact the cyanide test, AGA used principle component analysis for four sets of ‘factors’.  They concluded that As and Sb had little impact, but that sulphide content and coarse gold were the leading contenders for lowering recovery in the CN leach samples.


The AGA test work was deemed inconclusive due to small sample size and nugget effect.  However, it should be an indicator of processing and recovery possibilities and issues.  It also showed that gold and sulphide characterization studies are needed for metallurgical and process planning.  Any such study should address sample size, coarse free gold content, distribution and location of gold in host rock, material type (shale, volcanic, intrusive), sulphide species, and organic carbon content.  At this stage, the results were only considered as a preliminary indicator of potential issues for a cyanide leach process.


In 2006, ITH submitted a single sample of unoxidized vein-related mineralization to Hazen Research for a gold characterization study.  The sample showed that the bulk of the gold occurs as micron-scale native gold grains in and adjacent to pyrite and arsenopyrite grains with a smaller number of grains associated with silicate gangue.  Cyanide recovery in a bottle roll test was 61% (Table 16.1, Sample 1A).


In 2007 six more samples were submitted to Hazen Research for additional gold characterization studies.  These samples represented both high and low grade mineralization from oxidized, partially oxidized and unoxidized material.  Cyanidation of the samples shows that the cyanide extraction of gold is very high on the oxide and partially oxidized samples (Table 16.1) and somewhat less in the sulphide material.  Two of the sulphide samples (Table 16.1, samples 3 and 1A) were from rock with albitic alteration and they each returned 60% cyanide recovery.  The 3rd sulphide sample (Table 16.1, sample 5) came from rock with sericite alteration and had only a 42% recovery.


A very important result of this work is the observation that, for all the samples tested in 2007, the bulk of the gold recovered by cyanide extraction is released in the first 16 hours.  This implies that the gold is readily available to the cyanide solution.  Further studies will address the cyanide extraction on both fine and coarse material as a first step in the determination of the optimal recovery process.



TABLE 16.1
GOLD RECOVERY FROM 2007 CYANIDE EXTRACTION TESTS



Sample #


Ore Type

Average Grade (g/t)

% Cyanide Extraction*

1

Oxide Sediments

1.52

99.9%

2

Oxide Sediments High-grade

10.80

96.9%

3

Un-Oxidized Volcanic

1.52

59.7%

4

Oxide Sediments

1.39

99.9%

5

Un-Oxidized Volcanic

1.38

42.3%

6

Weakly Oxidized Volcanic  

1.06

90.2%

1A

Volcanic Un-Oxidized

2.30

60.9%

* Samples were 300 gram bottle rolls with sample material crushed to ~200 mesh and sampled every 8-10 hours for a total of 48 hours.


In 2008 an additional 24 samples were submitted to Hazen research for bottle roll testing on coarse material from a variety of lithologies and oxidation states (Table 16.2).  This was undertaken as a separate study from a previous one with Chemex.  Results indicate that overall average cyanide extraction was approximately 70% with 15 of the 24 samples showing greater than 70% recovery.  Interestingly many of the unoxidized samples showed better recovery than some of the partially oxidized samples.  These data also show that the majority of the gold is released to solution within the first 16 hours.  The same sample materials have been submitted to Kappes Cassiday in Reno for fine grinding and tests of gravity recovery and cyanide extraction at a -200 mesh grind.  The results are presented in Table 16.3.

TABLE 16.2
GOLD RECOVERY FROM 2008 HAZEN CYANIDE

EXTRACTION TESTS (-10 MESH)


Sample ID

Ore Type

Hazen Head   Au g/t

Chemex Head    Au g/t

Calculated Head       Au g/t

Residue Assay     Au g/t

Hazen Gold Extraction

Chemex Gold Extraction

Calculated Head Extraction

100112113

Partial Oxide Um

0.48

1.26

0.81

0.17

64%

87%

79%

100123124

Trace Oxide Um

0.83

0.83

0.81

0.33

60%

60%

59%

100588589

Partial Oxide Um

0.88

1.03

1.13

0.47

47%

54%

58%

100772773

Partial Oxide Intr

0.77

0.74

0.96

0.23

70%

69%

76%

100829830

Unoxidized Lower Seds

1.18

1.04

1.33

0.31

74%

70%

77%

101024026

Unox Volc

1.30

0.85

1.04

0.31

76%

64%

70%

101273274

Unox Volc

1.00

0.92

1.11

0.25

75%

73%

78%

101291292

Partial Oxide Volc

1.24

0.71

1.51

0.21

83%

70%

86%

101437438

Partial Oxide Volc

0.60

1.44

1.12

0.46

23%

68%

59%

101548549

Partial Oxide Volc

2.47

1.17

3.22

0.16

94%

86%

95%

101604605

Partial Oxide Volc

1.70

0.80

1.36

0.35

79%

56%

74%

101618619

Partial Oxide Volc

1.15

0.96

1.14

0.47

59%

51%

59%

101774775

Partial Oxide Volc

1.13

0.82

1.06

0.16

86%

80%

85%

101827829

Partial Oxide Volc

0.72

0.84

0.59

0.12

83%

86%

80%

101847849

Partial Oxide Volc

0.80

0.81

1.05

0.44

45%

46%

58%

101896897

Partial Oxide Volc

3.36

1.16

1.17

0.89

74%

23%

24%

102070071

Trace Oxide Volc

0.44

0.49

0.74

0.06

86%

88%

92%

102096097

Trace Oxide Volc

1.35

1.03

0.94

0.28

79%

73%

70%

102536537

Comp Ox Upper Seds

1.67

1.09

0.69

0.07

96%

94%

90%

102575576

Part Oxide Upper Seds

0.77

1.96

1.16

0.05

94%

97%

96%

102642643

Part Oxide Upper Seds

0.58

0.71

0.81

0.25

57%

65%

69%

102886887

Part Oxide Upper Seds

0.96

0.95

1.05

0.69

28%

27%

34%

102925926

Part Oxide Upper Seds

1.46

1.16

1.49

0.77

47%

34%

48%

103110111

Part Oxide Upper Seds

0.63

0.91

0.87

0.22

65%

76%

75%

*Samples were 1400 gram bottle rolls with sample material crushed to -10 mesh and sampled in multiples of 4 hours for a total of 72 hours.



TABLE 16.3
GOLD RECOVERY RESULTS FROM KAPPES CASSIDAY CYANIDE EXTRACTION TESTS (-200 MESH)


Sample ID

Calculated Head,

Extracted,  

Avg. Tails,

Au Extracted,

Leach Time,

Consumption NaCN,

Addition Ca(OH)2,

Au g/t

Au g/t

Au g/t

%

days

kg/t

kg/t

100112113

0.459

0.39

0.073

84.1%

3

1.10

2.75

100123124

0.609

0.47

0.144

76.4%

3

0.45

1.00

100588589

1.686

1.23

0.461

72.7%

3

0.53

2.00

100772773

0.728

0.51

0.221

69.6%

3

2.01

2.75

100829830

1.278

1.06

0.221

82.7%

3

0.55

2.50

101024026

0.620

0.54

0.077

87.6%

3

0.66

2.25

101273274

0.787

0.68

0.105

86.7%

3

0.51

1.50

101291292

1.333

1.21

0.125

90.6%

3

0.81

1.00

101437438

0.819

0.57

0.247

69.8%

3

0.48

1.50

101548549

2.670

2.51

0.162

93.9%

3

0.22

1.50

101604605

0.992

0.83

0.166

83.2%

3

0.37

1.50

101618619

1.434

1.15

0.280

80.5%

3

0.82

2.50

101774775

1.069

1.00

0.068

93.7%

3

0.56

1.50

101827829

2.733

2.67

0.063

97.7%

3

0.66

1.50

101847849

1.279

0.75

0.525

59.0%

3

0.48

1.50

101896897

1.269

0.52

0.747

41.1%

3

0.79

1.50

101925926

1.552

1.00

0.555

64.2%

3

0.12

1.50

102070071

0.594

0.52

0.077

87.0%

3

0.72

2.00

102096097

1.074

0.96

0.117

89.1%

3

0.57

1.50

102536537

0.875

0.84

0.034

96.1%

3

0.69

2.00

102575576

0.927

0.87

0.053

94.3%

3

0.71

1.50

102642643

0.596

0.48

0.120

79.9%

3

2.49

4.00

102886887

0.873

0.36

0.510

41.6%

3

1.28

4.00

103110111

0.711

0.60

0.110

84.6%

3

0.94

2.50

Average

1.124

0.90

0.219

79.4%

--

0.77

1.99

*Samples were 1000 gram bottle rolls with sample material crushed to -200 mesh and sampled in multiples of 4 hours for a total of 72 hours.



Comparing the results of the two test series, indications were that finer grinding improved the overall gold recovery, in some cases as much as 18 percent.  These results indicated that the gold was not refractory, but is tightly held in the ore matrix.  The gold recovery averaged 79.4 percent on an average head grade of 1.12 g/t.  Lime and cyanide consumption data were also gathered during this series of tests and are presented in Table 16.3.


Additional test work is currently underway on 35 composites made up of 1195 individual samples from the Livengood drilling campaign.  The composites are of eight different stratigraphic units further delineated by the degree of oxidation and gold grade.  The test work is being performed to further investigate ore chemical and physical characteristics, and the effectiveness of gravity and cyanidation for gold recovery.


Other test work planned for these composites includes flotation, gravity and flotation concentrate fine grinding and high intensity leaching, and aeration and lead nitrate addition.



16.5 Current Test Work Program


A test work program is currently underway (February 2010) at Kappes Cassidy and Associates (KCA) in Reno, Nevada, on Livengood ore samples.  KCA has also contracted with ALS Minerals to perform ICP analyses of the composites, and Phillips Enterprises LLC to perform grinding and abrasion studies.  This program is nearly complete and most of the test work results have been compiled and are included in this report.


Initially, thirty-five test composites were sorted and provided to the laboratory for testing.  The samples represent eight different stratigraphic units with distinct silicate mineral assemblages.  Samples from each stratigraphic unit were selected to represent variations in grade and degree of surficial oxidation.  Samples that make up the composites were sorted on site into 35 bins with an average weight of 200 kilograms.  These bins were shipped directly to the KCA laboratory in Reno, NV.


More recently, an additional 8 composites were sorted from recent drilling of the Sunshine Zone.  These composites were similar to two of the stratigraphic units previously supplied to KCA, Upper Sediments and Kint, but were from a new ore zone.


When the samples arrived at the lab, they were identified by composite, logged in, and weighed.  The lab blended the samples to insure the composites were thoroughly mixed and homogenous prior to removing any sample splits.  Samples were handled and stored in a manner which prevented the possibility of cross contamination with other clients’ samples and other Livengood composites.


The primary focus of the test work campaign was to identify the chemistry of each of the composites, identify the potential for utilizing gravity separation and cyanidation as a metallurgical processes for gold extraction, and establishing preliminary grinding parameters for the various Livengood ore types.  The lab conducted grind studies to develop laboratory stage ball mill grind times and developed Bond Ball Work indices.  Gravity concentration test work has been performed in a stage grinding test that identified the total gravity recoverable gold (GRG).  Cyanide shake leach tests and cyanidation bottle roll tests were performed in duplicate and at a target 80% passing 10 mesh, 100 mesh, and 200 mesh grind sizes.

The following diagram, Figure 16.1, presents the breakdown of sample requirements by composite for the proposed test work program.  A list of proposed tests and a test work outline for Livengood Gold ore follows the diagram.



[exhibit1044.jpg]

Figure 16.1.  Flow chart and breakdown of Livengood composite sample testwork.


The Livengood Samples were initially separated by the following Stratigraphic Units:

      • Overburden
      • Upper Sediments
      • Main Volcanics
      • Lower Sediments
      • Lower Sands
      • Kint
      • Cambrian
      • Amy Sequence

Each Stratigraphic Unit was then separated by degree of Oxidation:

 

      • None
      • Trace
      • Partial and Complete

Each Stratigraphic Unit by degree of Oxidation was composited by grade ;

 

      • 0.5 ppm Au to 1.0 ppm
      • >1.0 ppm to 5.0 ppm

Using this methodology the total number of composite samples comes to 54.  However, some of the composites selected were volumetrically insignificant in the deposit and therefore the total number of composites submitted totalled 41.


The composites were blended in order to ensure each composite was homogeneous prior to removing any sample splits.  Most of the composites weighed approximately 200 kg each, with 5 composites weighing about 40 to 50 kg.


Each composite has had a multi-element analysis performed by ALS Minerals (4-acid digest ICP-MS method ME-MS61m).  Gold was determined by triplicate 1 kilogram screen fire assays and silver was determined by triplicate fire assays with an AA finish.  Composites were also analyzed for sulfate, sulfide and total sulfur, as well as carbonate, organic carbon and total carbon.


All of the composites had a comparative cyanide leach assay using a hot cyanide leach and a cold cyanide leach. The tests were performed under conditions listed in Table 16.4.


After leaching the samples they were centrifuged and the solution removed for Au assay by atomic absorption spectrometry. Assays were performed in triplicate.


TABLE 16.4
CYANIDE SHAKE LEACH TEST PROCEDURE PARAMETERS


Procedure

Sample wt.

Soln. Temp.

Soln. NaCN Conc.

Soln. Amount

Leach Time

Hot Cyanide Leach

30 g

60°C

0.50 %

60 mL

1 hour

Cold Cyanide Leach

30 g

Ambient

0.50 %

60 mL

1 hour



TABLE 16.5

LIVENGOOD PROJECT – MAIN ZONE

SUMMARY OF CYANIDE SHAKE TESTS (5 GPL NACN)

           

Description

Average Head Assay,
Au g/t

Average Met Screen,
Au gs/t

Overall Average Head,
Au g/t

Average
Cyanide Sol.
(22 °C),
Au g/t

Average
Cyanide Sol.
(60 °C),
Au g/t

Overburden: Partial Ox (L)

0.82

0.59

0.71

0.27

0.30

Cambrian: Partial Ox (L)

0.28

1.21

0.75

0.19

0.21

Cambrian: Partial Ox (H)

2.17

1.78

1.97

0.21

0.21

Cambrian: Trace Ox (L)

0.69

0.66

0.67

0.15

0.15

Cambrian: Trace Ox (H)

1.79

1.79

1.79

0.23

0.36

Kint: Partial Ox (L)

0.80

0.72

0.76

0.21

0.25

Kint: Partial Ox (H)

0.68

2.18

1.43

0.41

0.47

Kint: Trace Ox (L)

0.73

0.76

0.75

0.02

0.02

Kint: Trace Ox (H)

0.68

1.43

1.06

0.01

0.03

Kint: No Ox (L)

0.66

0.89

0.77

0.02

0.01

Kint: No Ox (H)

0.93

0.95

0.94

0.01

0.05

Lower Seds: Trace Ox (L)

0.74

1.00

0.87

0.01

0.01

Lower Seds: Trace Ox (H)

1.81

0.85

1.33

0.01

0.02

Lower Seds: No Ox (L)

0.54

0.73

0.63

0.01

0.02

Lower Seds: No Ox (H)

0.78

1.10

0.94

0.01

0.02

Main Volcanics: Partial Ox (L)

0.53

0.77

0.65

0.16

0.25

Main Volcanics: Partial Ox (H)

1.79

1.75

1.77

0.19

0.39

Main Volcanics: Trace Ox (L)

0.73

0.74

0.73

0.03

0.10

Main Volcanics: Trace Ox (H)

1.12

1.55

1.33

0.05

0.07

Main Volcanics: No Ox (L)

0.96

1.02

0.99

0.05

0.08

Main Volcanics: No Ox (H)

3.01

1.88

2.45

0.03

0.06

Upper Seds: Partial Ox (L)

1.84

0.89

1.36

0.23

0.20

Upper Seds: Partial Ox (H)

1.30

1.40

1.35

0.38

0.41

Upper Seds: Trace Ox (L)

1.25

1.11

1.18

0.06

0.03

Upper Seds: Trace Ox (H)

0.94

1.53

1.24

0.09

0.08

Upper Seds: No Ox (L)

0.77

1.14

0.95

0.05

0.01

Upper Seds: No Ox (H)

2.77

0.99

1.88

0.06

0.03

Lower Sand: Partial Ox (L)

0.80

0.98

0.89

0.01

0.03

Lower Sand: Partial Ox (H)

1.52

2.01

1.76

0.04

0.05

Lower Sand: Trace Ox (L)

1.29

0.70

0.99

0.02

0.01

Lower Sand: Trace Ox (H)

0.82

1.33

1.08

0.03

0.01

Lower Sand: No Ox (L)

1.05

0.59

0.82

0.03

0.06

Lower Sand: No Ox (H)

0.75

1.25

1.00

0.05

0.02

Amy Sequence: Partial Ox (L)

1.34

0.29

0.81

0.09

0.09

Amy Sequence: No Ox (L)

0.49

0.44

0.46

0.03

0.06

 

1.12

0.10

0.12

 Descriptions from documentation provided by Talon Gold:  (L) - 0.5 ≤ Au g/t ≤ 1.0;   (H) - 1.0 ≤ Au g/t ≤ 5.0




TABLE 16.6

LIVENGOOD PROJECT – SUNSHINE ZONE

SUMMARY OF CYANIDE SHAKE TESTS (5 GPL NACN)


Description

Average Head Assay,
Au g/t

Average Met Screen,
Au gs/t

Overall Average Head,
Au g/t

Average
Cyanide Sol.
(22 °C),
Au g/t

Average
Cyanide Sol.
(60 °C),
Au g/t

Kint: Ox_high

2.25

1.51

1.88

0.35

0.51

Kint: Ox_Low

0.59

1.03

0.81

0.17

0.27

Kint:TraceOx_High

1.34

1.44

1.39

0.22

0.21

Kint: TraceOx_Low

1.24

0.81

1.02

0.13

0.22

Upper Seds: Ox_High

0.77

1.50

1.13

0.24

0.41

Upper Seds: Ox_Low

2.38

0.99

1.68

0.15

0.25

Upper Seds: Trace_High

1.32

1.60

1.46

0.18

0.25

Upper Seds: Trace_Low

0.63

0.84

0.74

0.09

0.19

     

1.26

0.19

0.29



As indicated by the test results, (Table 16.5 and 16.6) the response of the Livengood ores to the CN shake leach test procedure for determining gold leachability was poor.  The poor results were later found to be linked to “active” organic carbon in some of the ores, slow leaching gold ores, and large gold particle sizes.



16.5.1

Grind Studies and Ball Mill Bond Work Indices Tests


Grind studies were performed on each of the composites in order to establish grind time versus grind size relationships.  This information was used to prepare samples for future studies at varying grind sizes.


In addition to the above grinding tests, Bond Ball Work Index tests were performed.  The results of these tests will be used to obtain preliminary grinding operating costs and to perform preliminary mill sizing calculations.


A total of 43 composites were tested to achieve a work index for each of the ore types.  Tables 16.7 and 16.8 provide the results of the Bond Ball Work Index tests for rock from the Main Zone and the Sunshine Zone respectively.  Since the samples used for performing the tests were finer than typically received for bond testing, a conservative factor of 1.2 has been applied to the test work results.


Fifteen core samples from the Livengood property were obtained for abrasion tests. The results are shown in Table 16.9.  The abrasion data indicates that the Livengood ores vary from being medium abrasive (Ai of 0.30) to relatively non-abrasive (Ai less than 0.10).

 

TABLE 16.7

LIVENGOOD PROJECT – MAIN ZONE

BOND BALL MILL WORK INDEX TEST RESULTS

         

Description

BWI
kW-hr/st

BWI
kW-hr/MT

BWI x 1.2
kW-hr/st

BWI x 1.2
kW-hr/MT

Overburden: Partial Ox (L)

9.81

10.82

11.78

12.98

Cambrian: Partial Ox (L)

11.21

12.36

13.45

14.83

Cambrian: Partial Ox (H)

9.76

10.76

11.71

12.91

Cambrian: Trace Ox (L)

12.66

13.96

15.19

16.75

Cambrian: Trace Ox (H)

11.00

12.12

13.20

14.55

Kint: Partial Ox (L)

11.25

12.41

13.50

14.89

Kint: Partial Ox (H)

11.80

13.01

14.16

15.61

Kint: Trace Ox (L)

13.20

14.55

15.83

17.46

Kint: Trace Ox (H)

13.06

14.40

15.67

17.28

Kint: No Ox (L)

13.44

14.82

16.13

17.78

Kint: No Ox (H)

13.16

14.51

15.79

17.41

Lower Seds: Trace Ox (L)

13.33

14.70

16.00

17.64

Lower Seds: Trace Ox (H)

13.09

14.43

15.70

17.31

Lower Seds: No Ox (L)

13.26

14.62

15.92

17.55

Lower Seds: No Ox (H)

13.55

14.94

16.26

17.93

Main Volcanics: Partial Ox (L)

13.07

14.41

15.68

17.29

Main Volcanics: Partial Ox (H)

12.75

14.06

15.31

16.87

Main Volcanics: Trace Ox (L)

14.81

16.32

17.77

19.59

Main Volcanics: Trace Ox (H)

13.26

14.61

15.91

17.54

Main Volcanics: No Ox (L)

13.65

15.05

16.38

18.06

Main Volcanics: No Ox (H)

13.49

14.87

16.18

17.84

Upper Seds: Partial Ox (L)

13.53

14.91

16.23

17.89

Upper Seds: Partial Ox (H)

13.20

14.56

15.84

17.47

Upper Seds: Trace Ox (L)

13.21

14.57

15.85

17.48

Upper Seds: Trace Ox (H)

13.29

14.66

15.95

17.59

Upper Seds: No Ox (L)

13.69

15.09

16.42

18.11

Upper Seds: No Ox (H)

14.18

15.63

17.02

18.76

Lower Sand: Partial Ox (L)

15.36

16.93

18.43

20.32

Lower Sand: Partial Ox (H)

15.53

17.12

18.63

20.54

Lower Sand: Trace Ox (L)

15.92

17.55

19.11

21.06

Lower Sand: Trace Ox (H)

15.23

16.79

18.27

20.14

Lower Sand: No Ox (L)

15.18

16.73

18.21

20.08

Lower Sand: No Ox (H)

15.36

16.93

18.43

20.32

Amy Sequence: Partial Ox (L)

12.51

13.80

15.02

16.56

Amy Sequence: No Ox (L)

9.23

10.18

11.08

12.21

Average

13.14

14.49

15.77

17.39


 


TABLE 16.8

LIVENGOOD PROJECT – SUNSHINE ZONE

BOND BALL MILL WORK INDEX TEST RESULTS

Description

BWI
kW-hr/st

BWI
kW-hr/MT

BWI x 1.2
kW-hr/st

BWI x 1.2
kW-hr/MT

Kint_Ox_high

11.89

13.11

14.26

15.73

Kint_Ox_Low

12.12

13.36

14.54

16.03

Kint_TraceOx_High

12.92

14.24

15.50

17.09

Kint_TraceOx_Low

12.69

13.99

15.23

16.79

US_Ox_High

12.05

13.28

14.46

15.94

US_Ox_Low

12.67

13.97

15.20

16.76

US_Trace_High

12.96

14.29

15.55

17.15

US_Trace_Low

13.00

14.33

15.59

17.19

Average

12.54

13.82

15.04

16.59



TABLE 16.9

LIVENGOOD PROJECT

SUMMARY OF RESULTS

ABRASION TEST RESULTS

PHILLIPS REPORT 093029_15 OCTOBER 2009

       

Description

Rock Type

Alteration Type

Ai

Upper Seds: Partial Ox

Siltstone

Sericite

0.0023

Upper Seds: No Ox

Siltstone

Sericite

0.1497

Upper Seds: Partial Ox

Sandstone

Sericite

0.0120

Upper Seds: Partial Ox

Shale

Albite Mica

0.0848

Lower Seds: No Ox

Shale

Sericite

0.0189

Main Volcanics: No Ox

Andesite

Mixed Albite Mica Kspar

0.0391

Main Volcanics: Partial Ox

Volcanic Breccia

Albite

0.2872

Main Volcanics: Partial Ox

Volcanic Breccia

Clay Mica

0.1151

Main Volcanics: Partial Ox

Tuff

Sericite

0.1627

Main Volcanics: Partial Ox

Tuff

Albite Mica

0.0643

Amy Sequence: Trace Ox

Chert

Albite Mica

0.2040

Cambrian: Partial Ox

Serpentinite

No K or Na

0.0111

Cambrian: Partial Ox

Listwanite

Dolomite Clay Mica

0.0161

Cambrian: Trace Ox

Serpentinite

No K or Na

0.0343

Cambrian: Trace Ox

Gabbro

Clay Mica

0.0118



16.5.2   Gravity Centrifugal Concentration Evaluation


The Knelson® Gravity Recoverable Gold (GRG) tests were performed.  The test consists of three sequential liberation and recovery stages.  The progressive grinding was necessary in order to obtain an accurate GRG value, an indication of the size distribution of the GRG and a measure of progressive liberation.  It also limits any smearing of coarse gold particles that may be present in the as-crushed sample.


The GRG test is based on the treatment of a sample mass of typically 20 Kg using a laboratory Knelson Concentrator (KC-MD3).  Table 16.10 summarizes the test procedure.



TABLE 16.10

PROCEDURES FOR KNELSON CONCENTRATOR TESTWORK


Sample Require-ments

30 Kg of sample is required to perform a standard GRG test. 20 Kg of sample is required for the GRG test and the other 10 Kg sample is used for a grinding test prior to running the GRG.

 

Particle Size Requirements

Operating

Variables

Sample collection

Stage 1

90 - 100% -850 µm

Feed Rate:

800-1000 g/min

Fluid’n Water (FW):  3.5 l/min

Total Knelson concentrate for fire assay to extinction*

300 gr. tail sample for fire assay

Bulk tails to stage 2

Stage 2

45 - 60%  -75 µm

Feed Rate:

600-900 g/min

F.W: 3.5 l/min

Total Knelson concentrate for fire assay to extinction*

300 gr. tails sample for fire assay

Bulk tails to stage 3

Stage 3

75 - 80%  -75 µm

Feed Rate:

400-800 g/min

F.W: 3.5 l/min

Total Knelson concentrate for fire assay to extinction*

300 gr. of tails for fire assay

* the concentrate can  be panned for a visual observation of the concentrate - the panned products should then be assayed to extinction.



Note that it is not necessary to perform the test grind with 10 kg as this step has been previously performed in the grind studies portion of the test work.


Stage recoveries were based on the concentrate and tail assay of each stage.  However, overall recovery is based on the assays of the three concentrates produced and the tails product of the third recovery stage, whose assays are more reliable than those of the first two, which still contain some of the GRG.  Gold assays on the products will be by fire assay and in duplicate when sufficient sample exists.


Results from this test work for the Main and Sunshine Zones is shown in Table 16.11 and 16.12 respectively.  The gold in the Livengood ores appears to respond well to gravity separation.


TABLE 16.11

LIVENGOOD PROJECT – MAIN ZONE

KNELSON CONCENTRATOR - GRAVITY RECOVERABLE SUMMARY

             

Description

Calculated Head,
Au g/t

Conc + Mid Wt. %

Conc + Mid Assay,
Au g/t

Conc + Mid  Rec % Au

Conc + Mid Assay,
Ag g/t

Conc + Mid Rec % Ag

Overburden: Partial Ox (L)

0.55

1.3%

20.79

49.6%

14.3

7.3%

Cambrian: Partial Ox (L)

0.62

1.5%

28.16

66.4%

13.8

9.1%

Cambrian: Partial Ox (H)

1.34

1.3%

76.95

76.2%

15.6

10.9%

Cambrian: Trace Ox (L)

0.63

1.5%

29.88

69.8%

9.1

6.2%

Cambrian: Trace Ox (H)

1.59

1.5%

89.01

82.0%

13.8

10.7%

Kint: Partial Ox (L)

0.80

1.5%

19.07

35.6%

5.5

3.4%

Kint: Partial Ox (H)

1.67

1.5%

36.90

33.1%

10.2

5.4%

Kint: Trace Ox (L)

0.96

1.5%

25.72

40.6%

7.6

4.1%

Kint: Trace Ox (H)

1.41

1.5%

41.93

45.1%

7.2

5.1%

Kint: No Ox (L)

0.77

1.5%

15.77

31.1%

4.5

2.8%

Kint: No Ox (H)

1.40

1.5%

37.31

40.9%

5.6

2.8%

Lower Seds: Trace Ox (L)

1.12

1.5%

38.88

52.3%

5.8

3.2%

Lower Seds: Trace Ox (H)

1.21

1.5%

40.88

51.7%

12.0

8.3%

Lower Seds: No Ox (L)

0.75

1.5%

32.79

63.9%

6.3

4.4%

Lower Seds: No Ox (H)

1.21

1.5%

55.36

67.2%

9.5

6.5%

Main Volcanics: Partial Ox (L)

0.79

1.4%

22.37

40.1%

7.2

4.8%

Main Volcanics: Partial Ox (H)

1.80

1.4%

87.75

70.2%

12.9

9.9%

Main Volcanics: Trace Ox (L)

0.90

1.5%

23.83

38.7%

5.5

2.9%

Main Volcanics: Trace Ox (H)

1.65

1.5%

54.10

48.3%

7.9

4.2%

Main Volcanics: No Ox (L)

0.86

1.5%

23.20

40.1%

4.1

3.0%

Main Volcanics: No Ox (H)

1.84

1.5%

52.81

43.5%

6.2

3.4%

Upper Seds: Partial Ox (L)

0.84

1.4%

30.70

50.4%

6.6

6.3%

Upper Seds: Partial Ox (H)

1.42

1.4%

57.79

58.9%

8.9

7.0%

Upper Seds: Trace Ox (L)

0.80

1.4%

36.33

63.5%

8.2

4.6%

Upper Seds: Trace Ox (H)

1.42

1.4%

73.57

72.9%

10.1

6.6%

Upper Seds: No Ox (L)

0.84

1.4%

39.56

65.3%

8.3

4.7%

Upper Seds: No Ox (H)

1.11

1.4%

58.55

73.8%

8.1

5.3%

Lower Sand: Partial Ox (L)

1.09

1.5%

42.34

57.6%

8.0

6.6%

Lower Sand: Partial Ox (H)

1.42

1.4%

63.66

65.0%

11.8

6.0%

Lower Sand: Trace Ox (L)

0.99

1.4%

44.22

63.7%

9.2

5.3%

Lower Sand: Trace Ox (H)

1.34

1.5%

58.08

64.6%

9.4

7.6%

Lower Sand: No Ox (L)

0.72

1.4%

28.13

56.5%

6.0

3.1%

Lower Sand: No Ox (H)

1.48

1.4%

74.67

71.8%

11.5

5.8%

Amy Sequence: Partial Ox (L)

0.40

1.3%

15.00

49.3%

4.2

2.6%

Amy Sequence: No Ox (L)

0.57

1.4%

24.19

60.2%

7.1

3.6%

Averages

 

1.45%

42.9

56.0%

8.6

5.5%

(L) - 0.5 ≤ Au g/t ≤ 1.0; (H) - 1.0 ≤ Au g/t ≤ 5.0


TABLE 16.12

LIVENGOOD PROJECT – SUNSHINE ZONE

KNELSON CONCENTRATOR - GRAVITY RECOVERABLE SUMMARY

             

Description

Calculated Head,
Au g/t

Conc + Mid Wt. %

Conc + Mid Assay,
Au g/t

Conc + Mid  Rec % Au

Conc + Mid Assay,
Ag g/t

Conc + Mid Rec % Ag

Kint: Partial Ox (L)

1.87

1.27%

92.21

62.7%

8.3

5.9%

Kint: Partial Ox (H)

0.94

1.37%

39.20

57.5%

6.7

5.1%

Kint: Trace Ox (L)

1.43

1.27%

79.50

70.9%

14.5

8.3%

Kint: Trace Ox (H)

0.96

1.41%

43.32

63.5%

8.7

5.7%

Upper Seds: Partial Ox (L)

1.07

1.37%

57.54

73.6%

8.1

6.1%

Upper Seds: Partial Ox (H)

0.84

1.34%

42.32

67.6%

7.8

7.0%

Upper Seds: Trace Ox (L)

1.72

1.45%

95.31

80.5%

10.9

7.8%

Upper Seds: Trace Ox (H)

0.69

1.39%

34.94

70.6%

6.1

4.4%

Averages

 

1.36%

60.54

68.4%

8.9

6.3%

(L) - 0.5 ≤ Au g/t ≤ 1.0; (H) - 1.0 ≤ Au g/t ≤ 5.0



16.5.3

 Bottle Roll Leach Tests


Composite samples were be used to run 72 hour bottle roll tests.  For those composites with adequate amount of sample, bottle roll tests were run at 10 mesh, 100 mesh, and 200 mesh grinds.  Each bottle roll test had solution removed for Au assay at 2, 4, 8, 12, 24, 36, 48, and 72 hour intervals.  Cyanide and pH levels were also be checked as often as necessary to maintain reagents at adequate leach conditions.  Reagent consumptions were monitored and lime and cyanide consumptions were calculated.  Composites with insufficient amounts had only 72 hour bottle rolls run on them at the -200 mesh grind size.


Table 16.13 provides a summary of the bottle roll test results.  They indicate that most of the Livengood ores respond positively to cyanide leaching.  The bottle roll results were considerably better than the cyanide shake leach results.  However, gold leach recoveries appear to be highly variable by ore type.  The degree of oxidation also appears to have an effect on the gold cyanide leachability.



16.5.4

Bottle Roll CIL Tests


After reviewing the data from the bottle roll leach tests and the cyanide shake leach tests, it was determined that bottle roll Carbon-in-Leach (CIL) tests should be performed to establish if the poor response to cyanide leaching by some of the ore types was due to ore “preg robbing” issues.  Thus, the same composite samples were used to run 92 hour bottle roll CIL tests.  All of the bottle roll CIL tests were run at 200 mesh grinds.  Cyanide and pH levels were checked as often as necessary to maintain reagents at adequate leach conditions.  Reagent consumptions were monitored and lime and cyanide consumptions were calculated.


Recoveries improved significantly, with some ore types showing as high as a 49.5% increase in overall gold recovery, with the addition of carbon in the cyanide leach process.  It appears that some of the ores have “preg robbing” characteristics, which explain the poor response observed in the cyanide


TABLE 16.13

LIVENGOOD PROJECT

SUMMARY OF CYANIDE BOTTLE ROLL TESTS

Description

Calculated Head,
Au g/t

Overall Average Assay,
Au g/t

Au Extracted, %

Consumption NaCN, kg/t

Addition Ca(OH)2, kg/t

Overburden: Partial Ox (L)

0.63

0.66

87%

0.43

2.80

Cambrian: Partial Ox (L)

0.80

0.75

80%

0.26

3.00

Cambrian: Partial Ox (H)

1.35

1.82

83%

0.35

2.08

Cambrian: Trace Ox (L)

0.61

0.66

87%

0.32

2.00

Cambrian: Trace Ox (H)

1.48

1.82

90%

0.33

2.50

Kint: Partial Ox (L)

0.69

0.73

54%

0.51

4.50

Kint: Partial Ox (H)

1.67

1.55

60%

0.99

3.33

Kint: Trace Ox (L)

0.82

0.76

24%

0.38

2.50

Kint: Trace Ox (H)

1.24

1.23

22%

0.41

2.50

Kint: No Ox (L)

0.84

0.76

32%

0.29

2.80

Kint: No Ox (H)

2.42

1.51

32%

0.81

3.00

Lower Seds: Trace Ox (L)

1.05

0.88

0%

0.28

2.33

Lower Seds: Trace Ox (H)

1.20

1.36

1%

0.38

2.00

Lower Seds: No Ox (L)

0.62

0.65

0%

1.78

2.00

Lower Seds: No Ox (H)

1.36

1.28

0%

0.35

2.00

Main Volcanics: Partial Ox (L)

0.75

0.68

56%

0.30

2.67

Main Volcanics: Partial Ox (H)

1.43

1.68

77%

0.36

3.17

Main Volcanics: Trace Ox (L)

0.82

0.75

36%

0.51

2.00

Main Volcanics: Trace Ox (H)

1.45

1.42

42%

2.21

2.00

Main Volcanics: No Ox (L)

0.97

0.91

49%

0.20

2.00

Main Volcanics: No Ox (H)

1.66

2.10

39%

2.13

2.00

Upper Seds: Partial Ox (L)

0.71

1.05

64%

0.35

2.00

Upper Seds: Partial Ox (H)

1.45

1.50

80%

0.37

2.00

Upper Seds: Trace Ox (L)

0.89

0.97

37%

0.22

2.00

Upper Seds: Trace Ox (H)

1.67

1.58

73%

0.42

2.00

Upper Seds: No Ox (L)

0.76

0.85

26%

0.31

2.00

Upper Seds: No Ox (H)

1.28

1.68

55%

0.32

2.00

Lower Sand: Partial Ox (L)

0.91

0.86

49%

1.98

2.00

Lower Sand: Partial Ox (H)

1.10

1.52

61%

0.63

2.00

Lower Sand: Trace Ox (L)

1.09

0.94

48%

0.39

2.50

Lower Sand: Trace Ox (H)

1.35

1.32

67%

0.40

2.33

Lower Sand: No Ox (L)

0.68

0.75

21%

0.45

2.00

Lower Sand: No Ox (H)

1.32

1.28

55%

0.55

2.33

Amy Sequence: Partial Ox (L)

0.39

0.70

49%

0.24

2.60

Amy Sequence: No Ox (L)

0.52

0.51

4%

0.22

2.50

Overall Average

1.09

1.13

47%

0.59

2.38

TABLE 16.14

LIVENGOOD PROJECT – MAIN ZONE

SUMMARY OF CIL CYANIDE BOTTLE ROLL TESTS

 

Description

Calculated Head,
Au g/t

Overall Average Assay,
Au g/t

CIL Au Rec, %

BRT Au Rec, %

Difference between CIL and BRT Au Rec, %

Overburden: Partial Ox (L)

0.63

0.66

86.7%

87.2%

-0.5%

Cambrian: Partial Ox (L)

0.44

0.44

89.0%

80.0%

9.0%

Cambrian: Partial Ox (H)

1.33

1.33

94.0%

83.3%

10.7%

Cambrian: Trace Ox (L)

0.58

0.58

95.0%

87.0%

8.0%

Cambrian: Trace Ox (H)

1.64

1.64

95.0%

89.8%

5.2%

Kint: Partial Ox (L)

0.68

0.68

59.0%

54.0%

5.0%

Kint: Partial Ox (H)

1.52

1.52

59.0%

60.3%

-1.3%

Kint: Trace Ox (L)

0.78

0.78

42.0%

24.0%

18.0%

Kint: Trace Ox (H)

1.33

1.33

40.0%

21.5%

18.5%

Kint: No Ox (L)

0.76

0.76

49.0%

31.6%

17.4%

Kint: No Ox (H)

1.21

1.21

43.0%

32.2%

10.8%

Lower Seds: Trace Ox (L)

0.84

0.84

40.0%

0.0%

40.0%

Lower Seds: Trace Ox (H)

5.18

5.18

79.0%

0.5%

78.5%

Lower Seds: No Ox (L)

0.51

0.51

41.0%

0.0%

41.0%

Lower Seds: No Ox (H)

1.20

1.20

63.0%

0.0%

63.0%

Main Volcanics: Partial Ox (L)

0.76

0.76

71.0%

55.7%

15.3%

Main Volcanics: Partial Ox (H)

2.14

2.14

85.0%

76.7%

8.3%

Main Volcanics: Trace Ox (L)

0.92

0.92

63.0%

36.2%

26.8%

Main Volcanics: Trace Ox (H)

1.24

1.24

39.0%

41.7%

-2.7%

Main Volcanics: No Ox (L)

1.11

1.11

65.0%

49.0%

16.0%

Main Volcanics: No Ox (H)

2.31

2.31

23.0%

38.7%

-15.7%

Upper Seds: Partial Ox (L)

0.74

0.74

72.0%

64.0%

8.0%

Upper Seds: Partial Ox (H)

1.37

1.37

87.0%

80.3%

6.7%

Upper Seds: Trace Ox (L)

0.63

0.63

67.0%

36.5%

30.5%

Upper Seds: Trace Ox (H)

1.46

1.46

83.0%

73.0%

10.0%

Upper Seds: No Ox (L)

0.78

0.78

73.0%

26.2%

46.8%

Upper Seds: No Ox (H)

1.25

1.25

82.0%

54.8%

27.2%

Lower Sand: Partial Ox (L)

0.93

0.93

57.0%

49.2%

7.8%

Lower Sand: Partial Ox (H)

1.99

1.99

53.0%

61.4%

-8.4%

Lower Sand: Trace Ox (L)

1.01

1.01

60.0%

47.7%

12.3%

Lower Sand: Trace Ox (H)

1.32

1.32

62.0%

66.5%

-4.5%

Lower Sand: No Ox (L)

0.80

0.80

70.0%

20.5%

49.5%

Lower Sand: No Ox (H)

1.03

1.03

76.0%

54.5%

21.5%

Amy Sequence: Partial Ox (L)

0.47

0.47

79.0%

48.8%

30.2%

Overall Average

1.20

1.20

65.9%

48.0%

17.9%


shake leach tests.  Fortunately, the presence of activated carbon offsets, to a major degree, the “preg robbing” nature of the ores.


Similar tests were run on Sunshine Zone ores.  Tables 16.14 and 16.15 illustrate the results of these tests from the Main and Sunshine Zone ores respectively.


TABLE 16.15

LIVENGOOD PROJECT – SUNSHINE ZONE

SUMMARY OF CIL CYANIDE BOTTLE ROLL TESTS

       

Description

Calculated Head,
Au g/t

Overall Average Assay,
Au g/t

CIL Au Rec, %

Kint_Ox_(H)

1.97

2.25

86.5%

Kint_Ox_(L)

0.69

0.59

73.0%

Kint_TraceOx_(H)

0.87

1.34

79.1%

Kint_TraceOx_(L)

0.73

1.24

72.4%

Upper Seds_Ox_(H)

0.98

0.77

90.9%

Upper Seds_Ox_(L)

1.02

2.38

94.5%

Upper Seds_Trace_(H)

1.72

1.32

87.6%

Upper Seds_Trace_(L)

0.54

0.63

88.6%

Averages

 

1.31

84.1%



16.5.5 Flotation Concentration Tests


In order to understand how Livengood ores respond to sulfide flotation, a test program was developed for KCA to perform on their existing Livengood ore composites.  Knowing that the Livengood ores have a substantial amount of coarse gold, a test protocol was developed that would first subject the ore to sulfide flotation followed by performing a GRG test on the flotation tailings.  From this test scenario, a better understanding is gained of the ability to float the coarse gold while understanding the ability to collect gold in a pre- or post-flotation gravity circuit.


Batch flotation tests were performed by Kappes, Cassiday and Associates on samples drawn for the composites prepared for metallurgical testing from the reverse circulation drilling samples as described earlier in this report (Section 16.5).  Duplicate tests were conducted for each of the samples, with the sample material being ground to nominally 80% passing 0.075 mm.  The samples were then conditioned for 5 minutes with 5 g/t of CuSO4 and 25 g/t of PAX.  A float concentrate was then produced in 20 minutes with rougher flotation parameters of 25% solids and AF 70-20 g/t.  The flotation tails were then run through a Knelson Concentrator to collect the remaining gravity recoverable gold.  The middlings portion was recovered by hand panning the gravity concentrate.  All concentrate fractions and the gravity tails were assayed for gold and silver.


Results of the duplicate tests have been averaged and the proportion of total gold recovered by flotation and gravity are listed in Table 16.16.


TABLE 16.16

SUMMARY OF BATCH FLOTATION TEST RESULTS – MARCH 2010


Test Sample

% Gold Recovered by Flotation

% Gold Recovered by Gravity from Flotation Tails

Total Gold Recovered (%)

Volcanics, Partial Ox - Low Grade

76%

8%

84%

Volcanics, Partial Ox - High Grade

72%

15%

87%

Volcanics, Trace Ox - Low Grade

47%

49%

96%

Volcanics, Trace Ox - High Grade

66%

25%

91%

Volcanics, No Ox - Low Grade

79%

8%

87%

Volcanics, No Ox - High Grade

74%

20%

94%

Average

69%

21%

90%

Upper Seds, Core Z, Part Ox - Low Grade

78%

4%

82%

Upper Seds, Core Z, Part Ox - High Grade

63%

19%

81%

Upper Seds, Core Z, Trace Ox - Low Grade

39%

39%

78%

Upper Seds, Core Z, Trace Ox - High Grade

27%

69%

96%

Upper Seds, Core Z, No Ox - Low Grade

49%

32%

81%

Upper Seds, Core Z, No Ox - High Grade

53%

42%

95%

Upper Seds, Sunshine, Part Ox - Low Grade

61%

26%

87%

Upper Seds, Sunshine, Trace Ox - High Grade

73%

15%

88%

Upper Seds, Sunshine, Partial Ox - High Grade

69%

21%

90%

Upper Seds, Sunshine, Trace Ox - Low Grade

84%

7%

91%

Average

60%

27%

87%

Lower Seds, Trace Ox - Low Grade

34%

57%

91%

Lower Seds, Trace Ox - High Grade

24%

59%

83%

Lower Seds, No Ox - Low Grade

23%

54%

77%

Lower Seds, No Ox - High Grade

19%

66%

85%

Average

25%

59%

84%

Cambrian, Partial Ox - Low Grade

54%

35%

90%

Cambrian, Partial Ox - High Grade

41%

51%

92%

Cambrian, Trace Ox - Low Grade

74%

22%

95%

Cambrian, Trace Ox - High Grade

52%

44%

96%

Average

55%

38%

93%

Lower Sand, Partial Ox - Low Grade

78%

14%

93%

Lower Sand, Partial Ox - High Grade

80%

14%

94%

Lower Sand, Trace Ox - Low Grade

83%

11%

93%

Lower Sand, Trace Ox - High Grade

69%

26%

95%

Lower Sand, No Ox - Low Grade

73%

14%

86%

Lower Sand, No Ox - High Grade

61%

35%

95%

Average

74%

19%

93%

Kint, Partial Ox - Low Grade

74%

4%

79%

Kint, Partial Ox - High Grade

59%

21%

80%

Kint, Trace Ox - Low Grade

64%

27%

91%

Kint, Trace Ox - High Grade

72%

17%

89%

Kint, No Ox - Low Grade

78%

9%

87%

Kint, No Ox - High Grade

76%

20%

96%

Kint, Trace Ox - Low Grade

66%

22%

88%

Kint, No Ox - Low Grade

91%

3%

93%

Average

73%

15%

88%

       

Amy Sequence, Partial Ox - Low Grade

52%

30%

83%

Average

52%

30%

83%



16.5.6

CIL Recovery on Gravity Concentrates


Carbon in Leach (CIL) bottle roll tests (BRT) were performed on samples used to produce a gravity recoverable gold concentrate.  Twenty kilogram (20 Kg) samples were split from the composites discussed earlier in this section of the report, and then ground to 90% passing 0.85 mm.  The material was slurried in water and then fed into a Knelson Concentrator in 3 stages:


      • Stage 1: A gravity concentrate and tails was produced for the 90% passing 0.85 mm;
      • Stage 2: The tails from Stage 1 were milled to 50% passing 0.075mm and fed into the Knelson Concentrator, producing a Stage 2 concentrate and Stage 2 tails;
      • Stage 3: the tails from Stage 2 were milled to 80% passing 0.075mm and fed into the Knelson Concentrator, producing a Stage 3 concentrate and Stage 3 tails.

At each of the three stages, middlings were separated by hand panning the concentrate.  The middlings products and concentrate products were combined for each of the 3 stages, and CIL bottle roll tests were performed for the Stage 3 tails, the combined middlings, and combined concentrates.


The results of the CIL bottle roll tests on gravity recoverable gold concentrates and tails are summarized in Table 16.17.



TABLE 16.17

RESULTS OF CIL BOTTLE ROLL TESTS IN GRAVITY CONCENTRATION TESTS

MARCH 2010


Description

Product

% Gold Recovery in Product

% Total Gold Recovered

NaCN Consumption, (kg/MT)

Ca(OH)2 Addition, (kg/MT)

Cambrian: Partial Ox (L)

Con

97%

58%

9.04

2.54

Mid

91%

4%

6.08

0.72

Tail

77%

26%

2.21

0.50

Overall

90%

89%

2.27

0.51

 

Cambrian: Trace Ox (H)

Con

97%

67%

9.78

2.10

Mid

93%

6%

7.18

0.72

Tail

82%

19%

1.96

0.50

Overall

93%

92%

2.05

0.51

 

Kint: Partial Ox (H)

Con

89%

31%

9.33

0.88

Mid

76%

1%

7.86

0.79

Tail

48%

29%

2.21

0.50

Overall

63%

62%

2.31

0.51

 

Kint: No Ox (H)

Con

92%

45%

7.81

0.86

Mid

81%

1%

5.79

0.79

Tail

26%

12%

2.23

0.50

Overall

59%

59%

2.30

0.51

 

Main Volcanics: Partial Ox (H)

Con

98%

71%

13.12

0.95

Mid

95%

2%

7.05

0.74

Tail

68%

16%

2.38

0.50

Overall

90%

89%

2.49

0.51

 

Upper Seds: Trace Ox (L)

Con

78%

50%

12.34

1.97

Mid

71%

2%

6.30

0.75

Tail

36%

11%

1.82

0.50

Overall

64%

63%

1.90

0.51

 

Upper Seds: Trace Ox (H)

Con

96%

74%

9.63

0.95

Mid

88%

3%

7.03

0.81

Tail

48%

9%

1.80

0.50

Overall

87%

86%

1.89

0.51

 

Lower Sand: No Ox (H)

Con

95%

71%

5.85

1.43

Mid

87%

2%

6.91

0.80

Tail

44%

10%

2.10

0.50

Overall

83%

83%

2.17

0.51

 

Kint_Ox_high

Con

97%

66%

9.32

1.43

Mid

93%

2%

7.47

0.81

Tail

59%

17%

2.36

0.50

Overall

86%

85%

2.45

0.51

 

Kint_TraceOx_High

Con

97%

68%

15.13

1.47

Mid

64%

2%

6.15

0.77

Tail

54%

14%

1.98

0.50

Overall

84%

84%

2.09

0.51

 

US_Ox_Low-Sunshine

Con

96%

61%

17.21

2.01

Mid

92%

4%

8.33

0.77

Tail

69%

21%

2.04

0.50

Overall

87%

86%

2.16

0.51

 

US_Trace_Low –Sunshine

Con

96%

70%

12.31

1.72

Mid

93%

4%

7.90

0.78

Tail

69%

16%

1.86

0.50

Overall

90%

89%

1.96

0.51



16.6 Future Metallurgical Test Work


ITH has undertaken or is planning further metallurgical test work.  The following list is a brief outline of the test work currently envisioned for the Livengood project.


CIL Tests

      • Grind Size Effects
      • Leach Time Effects
      • Cyanide Strength Tests
      • Carbon Addition Concentration Tests
      • Lead Nitrate Tests

Organic Carbon Chemical Oxidation Tests

      • Oxidizer Tests

Gravity Tests

      • Grind Size Effects
      • Gravity Concentrate Fine Grind and Cyanidation Tests
      • Gravity Tails Leach and CIL Tests<

Flotation Tests

      • Collector and Depressant Tests
      • Grind Size Effects
      • Flotation Time Effects
      • Additional Flotation Concentrate and Tails Leach Tests

Gravity and Flotation Concentrate Tests

      • Additional Fine Grinding
      • Pre-aeration
      • Chemical Oxidation
      • Additional High Intensity Cyanide Leach Tests
      • Column Tests

      • Crush Size Effects
      • Leach Time Effects
      • Cyanide Strength Tests


Metallurgical test work program costs should be budgeted for Livengood at $750,000 for the next six months to cover the test work detailed above.

16.7 Mineral Processing


Based on the test work discussed previously and on the current estimated ore resource, process options were investigated.  The process envisioned for this document involves crushing the run of mine ore in a three stage crushing circuit to less than 3/4 inch, and placing the crushed ore on a lined heap leach pad and utilizing conventional heap leaching technologies.  Figure 16.2 presents a simple block flow diagram of the proposed circuit.  The placement of ore on the heap would likely not be performed during the coldest three months of the year due to arctic winter temperatures.  However, leaching would continue twelve months of the year since all piping, pregnant ponds, and solution application drip irrigators would be buried within the heap to prevent freezing.


Future mineral processing investigations will likely focus on a combination of a Mill/CIL process facility in conjunction with the heap leach facility.  Figure 16.3 presents a simple block flow diagram of the alternate circuit.



16.8 Gold Recovery


Utilizing existing test work data and industry experience, and applying the process scenarios described previously, an estimation of the gold recovery by ore type has been performed.  Table 16.18 provides the gold recoveries as currently estimated.  As shown in the table, the Lower Sediments are carrying a zero percent gold recovery and for the purposes of this study are not included as ore.  These ore types will be brought back into future studies when the Mill/CIL circuit is included as one of the proposed circuits.




[exhibit1045.jpg]

Figure 16.2  Proposed Livengood process block flow diagram showing heap leach process streams.


[exhibit1046.jpg]

Figure 16.3  Alternate Livengood process block flow diagram showing both heap leach and mill/CIL process streams


 

Current test work that may have a significant impact on the reported gold recoveries are those involving gravity concentration.  The estimated Mill/CIL recoveries may be improved with extraction of coarse gold in a gravity circuit followed by CIL processing.  The gravity circuit could address the two key recovery factors effecting the un-oxidized ores, coarse gold and preg-robbing carbon.


TABLE 16.18

GOLD RECOVERY ESTIMATES BY ORE TYPE FOR HEAP LEACH AND MILL/CIP PROCESS SCENARIOS


Ore Type – Main Zone

Heap Leach

% Au Rec

Mill / CIL
% Au Rec

Overburden

75.0%

87.0%

Cambrian Oxidized

80.0%

92.0%

Cambrian Trace

82.0%

95.0%

Cambrian Unoxidized

70.0%

75.0%

Upper Seds Oxidized

75.0%

87.0%

Upper Seds Trace

45.0%

82.0%

Upper Seds Unoxidized

35.0%

77.0%

Kint Oxidized

50.0%

59.0%

Kint Trace

22.0%

46.0%

Kint Unoxidized

22.0%

46.0%

Main Volcanics Oxidized

60.0%

79.0%

Main Volcanics Trace

40.0%

65.0%

Main Volcanics Unoxidized

35.0%

65.0%

Lower Seds Oxidized

0.0%

72.0%

Lower Seds Trace

0.0%

60.0%

Lower Seds Unoxidized

0.0%

60.0%

Lower Sand Oxidized

50.0%

63.0%

Lower Sand Trace

50.0%

63.0%

Lower Sand Unoxidized

15.0%

63.0%

Amy Sequence Oxidized

35.0%

79.0%

Amy Sequence Trace

35.0%

79.0%

Amy Sequence Unoxidized

27.0%

35.0%

     

Ore Type – Sunshine Zone

   

Sunshine Upper Seds Oxidized

78%

93%

Sunshine Upper Seds Trace

75%

88%

Sunshine Kint Oxidized

68%

80%

Sunshine Kint Trace

64%

76%

     
     



16.9 Process Operating Costs


ITH developed and presented the following operating costs for a 100,000 tpd heap leach operation in October, 2009 (Klipfel, et al., 2009).  Since then, further work has been done on cost estimation and assessing multiple recovery methods.  This work is incomplete at this time and is not presented here.  However, ITH views the currently presented heap leach scenario as a first step in assessing that component of the recovery process.


Process operating costs have been developed on each of the identified Livengood ore stratigraphic units (Table 16.19).  The test work performed on each of the ore types has provided preliminary data for calculating reagent and grinding media consumptions, and power consumptions by ore type.  The size of the facility envisioned and the unit processes involved also enabled preliminary maintenance and manpower requirements to be created for a 100,000 tpd process facility at Livengood.  Alaskan wage rates were applied to the various staff and operating and maintenance positions.  Unit costs for reagents and power were obtained from a survey of properties operating in the region.


The operating costs provided include power costs at $0.135 per kWh.


The 100 ktpd Stand Alone Heap Leach operating costs have been used in the financial calculations provided later in this report.


TABLE 16.19

PRIMARY ORE TYPES ESTIMATED PROCESS OPERATING COSTS

FOR HEAP LEACH AT 100,000 TPD THROUGHPUT RATE


Ore Type

100 ktpd Stand Alone Heap Leach Op. Cost, $/t

Main Zone Deposit

 

Overburden

 $           2.79

Cambrian Oxidized

 $           3.05

Cambrian Trace

 $           3.38

Cambrian Unoxidized

 $           3.38

Upper Seds Oxidized

 $           3.23

Upper Seds Trace

 $           3.36

Upper Seds Unoxidized

 $           3.36

Kint Oxidized

 $           3.97

Kint Trace

 $           3.58

Kint Unoxidized

 $           3.58

Main Volcanics Oxidized

 $           3.62

Main Volcanics Trace

 $           3.17

Main Volcanics Unoxidized

 $           2.70

Lower Seds Oxidized

 $           3.23

Lower Seds Trace

 $           3.36

Lower Seds Unoxidized

 $           3.36

Lower Sand Oxidized

 $           3.23

Lower Sand Trace

 $           3.36

Lower Sand Unoxidized

 $           3.36

Amy Sequence Oxidized

 $           3.05

Amy Sequence Trace

 $           3.38

Amy Sequence Unoxidized

 $           3.58

Sunshine Zone Deposit

 

Sunshine Upper Seds Oxidized

$            3.23

Sunshine Upper Seds Trace

$            3.36

Sunshine Kint Oxidized

$            3.97

Sunshine Kint Trace

$            3.58



16.10 Process Capital Cost


Preliminary capital costs were prepared for this study (Table 16.20). The capital costs that were developed are based on the following assumptions:

      • Strip Ratio of 1:1, waste to ore
      • 100,000 tpd ore production rate
      • Heap Leach processing scenario
      • Process facility would be relatively close to the Livengood ore deposits
      • Power to the site would be via a new power line included in the capital cost
      • The existing access to the site would be sufficient to support construction and operating requirements for Livengood


Capital costs for the different areas shown in the table were obtained from recent projects of similar size and scope, located in north western Canada and Alaska.  In an area where the size may have been different, the area was scaled using the rule of thumb factors.



TABLE 16.20

LIVENGOOD CAPITAL COST ESTIMATE
100,000 TPD HEAP LEACH – STAND ALONE


Area

Total  

(US$ Millions)

Crushing

105.3

Carbon Stripping and Regeneration

18.6

Refining

4.9

Reagents

22.0

Plant Mobile Equipment

50.8

Leach Pad and CIC Plant

98.8

Site Roads

10.5

Power Line

32.0

   

Administrative Building

4.9

Mine Truckshop and Warehouse

25.7

   

Subtotal

$373.6

   

Indirect Costs @ 50%

$186.8

   

Total Installed Costs

$560.4


 

17.0

Mineral Resource Estimate


The October 2009 mineral resource estimate for the Livengood deposit was updated using information available through February 28th, 2010.  The drill data was maintained in a Gemcom® GEMS database, and the basic statistical and geostatistical analysis was performed using SAGE2001® and WinGSLib®.  The resource model was constructed using Gemcom GEMS® and the Stanford GSLIB (Geostatistical Software Library) MIK post processing routine.  The mineral resource model was estimated using multiple indicator kriging (MIK) for gold.  Two oxidation indicators were used to estimate the oxidation and a single indicator was used to estimate the distribution of Kint dikes, Lower Sands, Amy Sequence and Shale.  A three-dimensionally defined lithology model, based on interpretations by ITH geologists, was used to code the r ock type block model.  A three-dimensionally defined probability grade shell (0.1 g/t) was used to constrain the gold estimation.  A summary mineral resource at cutoff grades of 0.3, 0.5, and 0.7 g/t gold is shown in Table 17.1.


TABLE 17.1

SUMMARY MINERAL RESOURCE


Classification

Au Cutoff (g/t)

Tonnes (millions)

Au (g/t)

Million Ounces Au

Indicated

0.30

702

0.60

13.5

Inferred

0.30

278

0.56

5.0

Indicated

0.50

369

0.78

9.3

Inferred

0.50

112

0.77

3.0

Indicated

0.70

184

0.98

5.8

Inferred

0.70

56

0.99

1.8



It is important to note that, compared to the October 2009 resource estimate, the tonnage and total ounces estimated has increased in the Indicated category and has decreased in the Inferred category for cutoff grades of 0.30, 0.50, and 0.70 g/t Au.  This change is due, in part, to addition of newly defined lower grade tonnes in the northeastern area of the deposit.  In addition, in this estimate, the composite data were declustered prior to indicator bin selection and data in four octants were required to estimate a block compared to 3 octants in the 2009 estimate.  The net result is that some previously inferred higher grade blocks were downgraded or extinguished through a combination of drilling and use of more rigorous statistical criteria for allowing certain blocks to be included in the resource.  The reason for changing these estimation parameters was to produce a more tightly constra ined model with greater reliability for the purposes of the preliminary economic assessment.



17.1

Data Used


17.1.1

Sample Data


The data available for this model comprised 83,200 metres of core and RC drilling, plus trench data.  Historical drilling and sampling is shown in Table 17.2.  Drilling performed by TGA is shown in Table 17.3.  It can be seen that the historical data represents about 6% of the total information used.

TABLE 17.2

HISTORICAL DRILLING AND SAMPLING


Year

Company

Drill Type

Number of Holes

Metres

1976

Homestake

Percussion

4

153

1981

Occidental

Percussion

6

310

1989

AMAX

Trench

2

160

1990

AMAX

RC

3

320

1997

Placer Dome

Core

9

1,100

2003

AngloGold

RC

8

1,514

2004

AngloGold

Trench

8

276

2004

AngloGold

Core

4

762

Total

   

47

4,746



TABLE 17.3

ITH DRILING AND SAMPLING


Year

Drill Type

Number of Holes

Meters

2006

Core

7

1,227

2007

Core

15

4,411

2008

Core

7

2,040

2008

Trench

4

80

2008

RC

108

28,619

2009

Core

12

4572

2009

RC

195

59757

Total

 

348

100,706



17.1.2

Other Data


Topography

The topographic surface used is based on a 4m DEM derived from 2008 aerial photography.


Density

Densities used in the resource are based on 98 determinations from core and RC chip samples and are shown in Table 17.4.


TABLE 17.4

DENSITY DETERMINATIONS


Lithology Unit

N

Mean

StdDev

Max

Min

Amy Sequence

4

2.67

0.04

2.72

2.65

Cambrian

12

2.82

0.07

2.95

2.69

Combined Cambrian-Amy

 

2.78

     

Kint

3

2.56

0.18

2.76

2.44

Lower Sediments

21

2.74

0.05

2.84

2.62

Main Volcanics

36

2.72

0.13

2.86

2.11

Upper Sediments

22

2.68

0.13

2.79

2.23

Average of all readings

98

2.72

     



17.2

Data Analysis


Multi-element assay information is available for less than 50% of the samples.  A statistical summary of this data from a previous report (July 09) is shown in Table 17.5.  The only element of economic significance is gold, which was the only element modeled in the resource model.  No significant correlations were found between the various elements.  There were numerous weak to moderate correlations, but nothing that could be exploited to improve the gold estimate.  Based on the lack of significant correlations previously determined, the exercise was not updated for this estimate (Figure 17.1).  It is still a matter of geological debate as to exactly why this is so, but the volcanic unit is preferentially mineralized relative to the units above and below it.  Also, the Kint dikes, which appear to be the conduits for much of the mineralization, are also well miner alized.  Not only are the volcanics and Kint dikes higher grade, they are uniformly well mineralized as shown by the relatively low coefficient of variation (C.V.) of each unit.


TABLE 17.5

STATISTICAL SUMMARY OF ASSAY DATA


Element

Units

N

Mean

Maximum

Std.Dev.

C.V.

Au

ppm

34786

0.40

56.2

1.22

3.0

Ag

ppm

12969

0.41

440

4.07

10.0

Cu

ppm

12969

42

1120

34

0.8

Pb

ppm

12969

19

9240

128

6.7

As

ppm

12971

2169

137000

4181

1.9

Sb

ppm

12969

221

138000

2394

10.8

Zn

ppm

12969

186

3440

221

1.2

Fe

%

12708

4.3

21.3

1.4

0.3

Mo

ppm

12969

5.5

74.0

6.9

1.3

S

%

12081

1.4

18.4

1.4

1.0

Te

ppm

12063

0.16

25.1

0.5

3.0


[exhibit1048.jpg]

Figure 17.  Gold distribution by lithology unit.



Each of the assay intervals were also logged for lithology, alteration and mineralization.  Of all of the available qualitative data, the lithology appears to exert the most influence on the gold mineralization



17.3

Geological Model


ITH geologists provided sectional interpretations of the major lithologic units - these were used to generate a three dimensional wire framed geological model of these units and major fault structures.  South of the Lillian Fault, the rock units modeled were the Cambrian, Upper Sediments, Main Volcanics, and the Lower Sediments.  North of the Lillian fault most of the material is undifferentiated Upper Sediments, with a small amount of Volcanics and Lower Sediments modeled.  These represent the major lithologic units that host the mineralization.  No other controls were modeled.



17.4

Composite Statistics


All of the available drilling was composited into fixed length 10m composites.  Composite residuals <4m in length were added to the previous composite.  These composites were back-tagged with the lithology using rock type block model developed from the defined geological three-dimensional wire frames.


The composite data was declustered by estimating a nearest-neighbour value into each block.  The declustered composite statistics are tabulated below, (Table 17.6).


TABLE 17.6

GOLD COMPOSITE STATISTICS


Mean:

0.37

Variance:

0.24

C. of V.:

1.32

Min:

0.005

Q1:

0.11

Median:

0.23

Q3:

0.47

Max:

10.99



17.4.1

Gold Indicator Statistics


The composite data was used to set the gold indicator thresholds.  Since the coefficient of variation of the composite data is relatively low, only nine indicator thresholds were needed to fully define the gold distributions.  The indicator thresholds were chosen at the low end to have approximately 20% of the data per class and at the high end to have 10 to 11% of the metal per class (Table 17.7).  With MIK, top cutting of the assays is not necessary.  In this case all composite values greater than 1.8 g/t Au (the highest threshold) are treated the same as “high grade” and the mean value of 3.14 g/t Au is used to evaluate the highest class.


TABLE 17.7

GOLD INDICATOR STATISTICS


   

Data

Metal

 
 

Threshold

%

Cum%

%

Cum%

Mean

1

0.09

19.4

19.4

2.8

2.8

0.053

2

0.18

20.5

39.9

7.4

10.2

0.133

3

0.28

17.6

57.5

10.9

21.1

0.228

4

0.40

12.7

70.2

11.5

32.6

0.335

5

0.60

12.6

82.9

16.8

49.4

0.493

6

0.75

5.2

88.1

9.3

58.8

0.658

7

0.90

4.1

92.2

9.0

67 .8

0.818

8

1.20

3.8

96.0

10.7

78.4

1.034

9

1.80

2.7

98.6

10.1

88.5

1.405

Max

10.99

1.4

100.0

11.5

100.0

3.137



17.4.2

Contact Analysis


Because significant grade contrasts were noted between the different rock types from the assay statistics, contact analysis was performed in the previous study (October 2009) using the composite data to evaluate grade discontinuities at the lithology contacts.  Wherever a contact was crossed with a drill hole, the grade profile was examined on either side of the contact.  Contacts were evaluated from the Cambrian to the Upper Sediments, from the Upper Sediments into the Main Volcanics, and from the Main Volcanics into the Lower Sediments.


Between the Cambrian and Upper Sediments the grade contrast is fairly significant.  In the vicinity of the contact, the average grade of the Cambrian is 0.30 g/t Au while the Upper Sediments is 0.45 g/t Au (Figure 17.2).


Between the Upper Sediments and the Main Volcanics the grade contrast is also fairly significant.  The contact between the Main Volcanics and the Lower Sediments is the most significant with the grade in the Main Volcanics being 0.63 g/t Au and the Lower Sediments 0.43 g/t Au. The additional data available for this update did not appear to alter these relationships, and the contact analysis was not repeated.


[exhibit1050.jpg]

[exhibit1052.jpg]

Figure 17.  Contact plots.  



Because of the sharp contrasts in gold grade between the different units, it was decided to treat the boundaries between the different units as hard boundaries.  That is, the blocks of a given unit were estimated using only the composite data that fell within the same unit.  This is geologically reasonable since many of the contacts are associated with thrust faulting.  But it is not known if there has been any post-mineralization movement of these faults.  The Main Volcanics are unquestionably better mineralized than the surrounding units.  The reason for this is not fully understood.  With this, it is not geologically unreasonable to see grade discontinuities at the contacts for this reason either.


The use of hard boundaries will have an impact on the local estimates because the data has been partitioned.  Overall, whether hard boundaries or soft boundaries are used or not would have a minimal effect on the global estimate.  The issue as to whether hard or soft boundaries are more appropriate should be resolved as more drilling is done and additional information is gathered.



17.5

Spatial Statistics


17.5.1

Gold Indicator Variograms


Indicator variograms were calculated for each of the indicator thresholds within each of the lithologic domains.  Variogram models were fitted for each.  Because the data was so heavily partitioned the results from the individual domains were generally unsatisfactory.  Many of the areas are relatively thin, especially in the Main Volcanics, making it very difficult to infer a model of vertical continuity.  For this reason, the use of the partitioned data for variogram calculations was abandoned and all of the data was used to calculate a set of average indicator variograms that were used over all domains.  The average indicator variograms that were used for estimation of the gold indicators in all domains are shown in Table 17.8.  



17.5.2

Oxide Indicator Variograms


The oxidation model was estimated using two oxide indicators, one for oxidized and one for trace (Table 17.9).  Both the oxidized indicator variogram and the trace indicator variogram are essentially horizontal.



17.5.3

KINT Dike Variograms


A continuous dike indicator was defined using the percentage of Kint dike within each logged interval.  The presence and behavior of the dikes north and south of the Lillian Fault are significantly different.  Different variograms were fitted for each of these dike domains (Table 17.10).  The variogram in the north dips steeply to the south.  The variogram in the south was rotated with the horizontal plane dipping to the south-west.



17.5.4

Amy Sequence, Lower Sands and Shale Variograms


Continuous indicators were defined using the percentage of Amy Sequence, Lower Sands and Shale within each logged interval (Table 17.11).  The Amy Sequence material occurs only in the Cambrian, south of the Lillian Fault.  The Lower Sands material occurs only in the Lower Sediments, and the Shale occurs throughout the model, largely paralleling the stratigraphy.  The units dip steeply north on the north side of the Lillian fault because of overturning, so separate shale variograms were calculated for the areas north and south of the fault.


TABLE 17.8

AVERAGE GOLD INDICATOR VARIOGRAMS


Indicator

Sill

Range X

Range Y

 Range Z

1

0.57

     
 

0.30

96

33

59

 

0.13

176

360

251

2

0.57

     
 

0.27

179

38

82

 

0.16

217

391

404

3

0.59

     
 

0.24

134

48

88

 

0.17

217

460

364

4

0.56

     
 

0.28

105

33

82

 

0.16

197

398

382

5

0.65

     
 

0.21

104

26

92

 

0.14

380

216

393

6

0.66

     
 

0.24

159

31

56

 

0.10

211

499

411

7

0.74

     
 

0.19

136

30

39

 

0.07

198

358

600

8 & 9

0.83

     
 

0.12

130

79

22

 

0.05

168

539

227



TABLE 17.9

OXIDE INDICATOR VARIOGRAMS


Indicator

Sill

Range X

Range Y

 Range Z

Oxidized

0.19

     
 

0.40

134

73

115

 

0.41

2317

2553

273

Trace

0.03

     
 

0.52

155

47

144

 

0.45

2867

1117

320


TABLE 17.10

KINT DIKE VARIOGRAMS


Domain

Sill

Range X

Range Y

 Range Z

North

0.30

     
 

0.51

64

54

616

 

0.19

119

552

696

South

0.23

     
 

0.65

259

19

33

 

0.12

368

254

431



TABLE 17.11

LOWER SANDS, SHALE & AMY SEQ. VARIOGRAMS


Domain

Sill

Range X

Range Y

 Range Z

L. Sand

0.22

     
 

0.46

63

189

233

 

0.32

633

2570

2

Amy Seq.

0.15

     
 

0.25

579

115

114

 

0.60

774

614

211

N. Shale

0.21

     
 

0.67

91

48

110

 

0.12

95

812

399

S. Shale

0.11

     
 

0.63

46

40

177

 

0.26

1000

1205

167



The Amy Sequence variogram dips shallowly to the East, while the Lower Sand variogram is essentially horizontal.  The North Shale variogram dips shallowly to the North-West, and the South Shale variogram is horizontal.



17.6

Resource Model


17.6.1

Model Extents


The resource model was constructed to encompass the drilling data and the defined geological model.  The entire project is done using UTM NAD27 Alaska coordinate system.  The model extents are shown in Table 17.12.

TABLE 17.12

MODEL EXTENTS


 

Minimum (m)

Maximum (m)

Extent (m)

Block Size (m)

No. of Blocks

East

427,500

430,800

3,300

15

220

North

7,264,300

7,266,700

2,400

15

160

Elevation

50

560

510

10

51



The selected block size was chosen because it is envisioned that the deposit will be mined with bulk mining methods that would not warrant smaller blocks but also because the drill hole spacing would not support a smaller block size.



17.6.2

Gold Estimation


The gold contained within each block was estimated using MIK with nine indicator thresholds.  The block model was tagged with the geological model using a block majority coding method.  The contact analysis indicated that there are significant grade discontinuities at the lithologic boundaries.  Hard boundaries were used between each of the units.  That is, each unit was estimated using only data that also fell within the same unit.  There was no potentially economic mineralization outside of the geological model and it was not estimated.  The gold kriging plan is shown in Table 17.13 for all units.


TABLE 17.13

GOLD KRIGING PLAN


Minimum No. of Composites

8

Maximum No. of Composites

48

Maximum Composites per Octant

6

Maximum No. of Composites per Hole

4

Block Discretization

4 x 4 x 1

Search Distances (m)

300 (Maj.), 250 (Semi-Maj.), 200 (Min.)

Search Rotation

Maj. -30º à170º, Semi-Maj. 100º



An octant search was used.  The kriging plan forces data to be available from a minimum of two octants and from two separate drill holes for an estimate to be made.  Each of the gold indicators was estimated independently.



17.6.3

Oxidation Estimation


Two levels of oxidation were estimated: oxidized and trace oxidation.  These levels correspond to the metallurgical testing and were therefore necessary to estimate to allow the application of the metallurgical recoveries to the model.  The oxidation level has been visually logged for each sample interval by ITH geologists.  Two oxidation indicators were used to estimate the oxidation.  Historically, oxidation has been logged using ten different descriptors ranging from “complete” to “ none”.  Any interval described as “moderate” or greater was classified as oxidized.  Any interval described as anything except “none” was classified as trace or better.  The two indicators were tagged on each of the samples as 1 (meeting the criteria) or 0 (not meeting the criteria).  Each indicator represents the probability of th e sample being oxidized.  These indicators were composited into 10m composites with the rest of the data.  The two indicators were estimated independently.  The kriging plans are shown in Table 17.14 and Table 17.15.


TABLE 17.14

OXIDIZED KRIGING PLAN


Minimum No. of Composites

8

Maximum No. of Composites

48

Maximum Composites per Octant

6

Maximum No. of Composites per Hole

4

Block Discretization

4 x 4 x 1

Search Distances (m)

300 (Maj.), 150 (Semi-Maj.), 100 (Min.)

Search Rotation

None



TABLE 17.15

TRACE OXIDIZATION KRIGING PLAN


Minimum No. of Composites

8

Maximum No. of Composites

48

Maximum Composites per Octant

6

Maximum No. of Composites per Hole

4

Block Discretization

4 x 4 x 1

Search Distances (m)

300 (Maj.), 150 (Semi-Maj.), 100 (Min.)

Search Rotation

None



The blocks were then coded as fully oxidized (coded as 1) if the probability of being oxidized was greater than 50%.  The blocks were coded as trace (coded as 2) oxidized if the probability of trace oxidization was greater than 50% and not already tagged as oxidized.  The remaining un-oxidized blocks were coded as 3.  As would be expected, the fully oxidized material is nearer the surface and consequently mostly in the Cambrian rocks.  The trace oxidization is pervasive.  Significant un-oxidized material is not encountered except in the lower sediments.

 

17.6.4

KINT Dike Estimation


The Kint dikes are significant metallurgically.  It was therefore necessary to estimate them.  The dikes are small enough that the drilling information is insufficient to build a deterministic model of the dike locations.  Consequently, the dikes were estimated using a probabilistic model.  In each block in the model, the probability of encountering dike was treated as the dike proportion within the block.


A single continuous dike indicator was used to estimate the presence of dikes.  The presence of dikes was logged for each logged interval.  The percentage of dike within the interval was logged, as in many cases the dike represented less than 100% of the interval.  The dike indicator was set to be the proportion of dike within the interval.  This indicator was then composited into 10m composites along with the rest of the data.


The presence and distribution of dikes is significantly different north and south of the Lillian Fault.  The two domains were estimated separately.  The kriging plan to estimate the proportion of dike within each block is shown in Table 17.16 and Table 17.17.


The Kint dikes are important for metallurgical but make up a very small portion of the total resource.  The Kint dikes average between 3 and 4% of the tonnage.



TABLE 17.16

KINT DIKE INDICATOR KRIGING PLAN – SOUTHERN DOMAIN


Minimum No. of Composites

8

Maximum No. of Composites

48

Maximum Composites per Octant

6

Maximum No. of Composites per Hole

4

Block Discretization

4 x 4 x 1

Search Distances (m)

300 (Maj.), 250 (Semi-Maj.), 150 (Min.)

Search Rotation

Maj. -55º à248º, Semi-Maj. 80º



TABLE 17.17

KINT DIKE INDICATOR KRIGING PLAN – NORTHERN DOMAIN


Minimum No. of Composites

8

Maximum No. of Composites

48

Maximum Composites per Octant

6

Maximum No. of Composites per Hole

4

Block Discretization

4 x 4 x 1

Search Distances (m)

300 (Maj.), 250 (Semi-Maj.), 50 (Min.)

Search Rotation

Maj. -80º à191º, Semi-Maj. 352º

 

17.6.5

Amy Sequence, Lower Sands and Shale Estimation


The Amy Sequence, Lower Sands and Shale units are significant metallurgically.  It was therefore necessary to estimate them.  The occurrences are small enough that the drilling information is insufficient to build a deterministic model of their locations.  Consequently, these were estimated using a probabilistic model.  In each block in the model, the probability of encountering these units was treated as the material proportion within the block.


A single continuous indicator was used to estimate the presence of the units.  The presence of Amy Sequence, Lower Sands and Shale was logged for each logged interval.  The percentage of these units within the interval was logged, as in many cases the lithology represented less than 100% of the interval.  The unit indicator was set to be the proportion of lithology within the interval.  This indicator was then composited into 10m composites along with the rest of the data.  The kriging plan to estimate the proportion of these units within each block is shown in Table 17.19.  Note that the Amy Sequence occurs only in the Cambrian, and that the Lower sands occur only in the Lower Sediments.



TABLE 17.18

LOWER SANDS, SHALE & AMY SEQ.  INDICATOR KRIGING PLAN


Minimum No. of Composites

8

Maximum No. of Composites

48

Maximum Composites per Octant

6

Maximum No. of Composites per Hole

4

Block Discretization

4 x 4 x 1

Search Distances (m) – Lower Sand

Major 300, Int. 150, Minor 100

Search Rotation – Lower Sand

Major  0º à Azimuth 290º

Search Distance (m) – Amy Sequence

Major 300, Int. 150, Minor 100

Search Rotation – Amy Sequence

Major  0º à Azimuth 104º

Search Distance (m) – South Shale

Major 300, Int. 300, Minor 75

Search Rotation – South Shale

Major  0º à Azimuth 342º

Search Distance (m) – North Shale

Major 300, Int. 230, Minor 50

Search Rotation – North Shale

Major  -7º à Azimuth 293º



17.7

Model Validation


Various forms of model validation were undertaken and are shown below.  In all cases, the model appears to be unbiased and fairly represent the drilling data.  The composite data was declustered by estimating a nearest-neighbour value into each block.


17.7.1

Global Bias Check


The global average of the declustered composite values is 0.370 g/t Au and the corresponding average block value (E-Type estimate, or block average calculated from MIK bins) is 0.373 g/t.  Block estimation using Ordinary Kriging yielded a global average of 0.375 g/t.  The estimated block values are within 1 – 1.5% of the composite values.  This is reasonable and within the expectations of the model.



17.7.2

Visual Validation


The model was visually compared to the composite gold data in both N-S and E-W sections.  The estimates were checked to see that they appeared to be consistent with the data and that they were geologically reasonable.  In all cases everything appeared reasonable.



17.7.3

Swath Plots


Swaths were taken through the model and the averaged block values (e-type MIK estimates) and the averaged declustered composite values (nearest-neighbour estimates) were compared on E-W, N-S and vertical swaths (Figure 17.3).  The kriged values have a small amount of spatial smoothing, but generally compare quite favourable to the composite values.



[exhibit1054.jpg]

Figure 17.3  Swath plots of E-type estimate vs. nearest neighbour.



17.8

Post-processing of MIK Model


The post-processing of the indicator kriging was done with the GSLIB post processing routine (postik).  It is necessary to provide a maximum grade of the distribution. This grade can be calculated as:


Zmax = Zcn + 3(Zn - Zcn)

where Zcn is the uppermost indicator threshold, and Zn is the mean of values > Zcn.


From the data in Table 17.9, the maximum grade used in the post-processing was calculated to be 5.82 ppm.



17.8.1

Change of Support


The multiple indicator kriging produces an estimate of the distribution of grade within a block rather than just a single average grade of a block.  The distribution produced is the distribution of composite sized units within the block not minable units.  It is therefore necessary to correct the distribution so that the distribution represents selective mining units (SMU’s) not composite sized units.  This correction is called a change of support correction.  Since the average grade of the block is the same whether mined in one scoop or mined by a core drill, the correction does not change the average grade of the block only reduces the variance of the distribution.


The variance reduction factor is the ratio of the variance of an SMU within a block to the variance of a composite within a block.  This is calculated using average variogram values.  The variance of the SMU within the block is the variance of a composite within a block minus the variance of a composite within an SMU.  Since the estimated blocks are small relative to the data spacing the effective block size was taken to be 40m by 40m (approximately ½ the drill spacing).


The method used for the change of support was an indirect lognormal correction.  This correction uses the ratio of standard deviations rather than the ratio of variances.  This is just the square root of the ratio of variances.


The mining SMU was assumed to be 5m by 5m selectivity.  This is reasonable for the envisioned size of the operation.  If the envisioned size of the operation were to grow significantly, the SMU size should be increased.


The following factors were derived using the variogram model.


[exhibit1056.jpg]= 0.716


[exhibit1058.jpg] = 0.590


[exhibit1060.jpg]


= 0.42


This correction is applied on a block-by-block basis with a global reduction target of 0.42.  This is done on a trial and error basis to find the block reduction factor that will achieve the target global variance reduction of 0.42.  A reduction factor of 0.28 was used by block.  The variance reduction factor is higher than that used in the October 2009 report.  This reflects the fact that the data has been restricted to the constraining probability grade shell and exhibits more variability given the removal of low grade, generally more continuous data.



17.9

Resource Classification


The resource was broken down into two categories: Indicated and Inferred.  The estimation variance from the estimation of the second indicator (median indicator) was used to determine the classification.  Along with the estimation of variance, the number of composites used, number of drill holes used and the distance to the nearest composite was saved for each block estimated.  The estimation variance provides a good measure of the confidence in the estimate.  The estimation variance will remain relatively low when data is near and evenly spaced around the block being estimated.  When the estimate starts extrapolating away from data, the estimation variance will rise rapidly.  An examination of plots of distance to the nearest sample versus variance (Figure 17.4), along with visual inspection of the model relative to the composite data were used to determine the acceptable estimati on variance thresholds.


Blocks estimated with an estimation variance less than 0.33 and with a minimum of 4 octants informed, should be considered Indicated.  Blocks with an estimation variance less than 0.43, and a minimum of 4 octants informed, should be considered Inferred.  Blocks with an estimation variance greater than 0.43 were considered to be too unreliable for further consideration.  Note that in the previous resource estimate (Klipfel, et al., 2009b) a minimum of only three octants was required for inclusion in the resource.  The decision to change to 4 octants for this estimate was based on a review of the previous estimate in conjunction with the revised geological model.  The selection thresholds have increased as compared to the October 2009 report – this is a reflection of the additional data available from infill drilling in several locations within the deposit.


On average, Indicated blocks are within 35m of the nearest composite, and are informed by 26 composites from at least 7 drill holes.  On average, Inferred blocks are within 87m of the nearest composite, and are informed by 20 composites from at least 5 drill holes.



[exhibit1062.jpg]

Figure 17.4.  Distance to the nearest composite vs. kriging variance.



18.0

Other Relevant Data and Information


No additional information or explanation is known by the authors to be necessary to make the technical report understandable and not misleading.



19.0

Interpretation and Conclusions


The Livengood property is centered on Money Knob and adjacent ridges and is an area considered by many for a long time to be the lode source for gold in the Livengood placer deposits which have produced in excess of 500,000 ounces of gold.  Anomalous gold in soil samples occurs in a northeast trend over an area of approximately 6 x 2 km with a principal concentration of surface anomalies in a smaller area measuring approximately 2.3 x 1.1 km.  Drilling by past companies, AGA, and ITH identified wide intervals (>100 m @ ≥ 1.0 g/t Au) of gold mineralization with local higher grade narrow intervals beneath the soil anomaly and in rocks beneath thrust surfaces which are not expressed geochemically at the surface.  The presence of mineralization over broad areas beneath thrust faults and the ever expanding area of drill hole intercepts suggests that there is still further discovery potential at Livengood.


The style of mineralization shows some similarities with several types of gold deposits including orogenic, sediment-hosted disseminated (SHD or Carlin type), and Intrusion-Related-Gold Systems (IRGS) of the Tintina Gold Belt.  However, the geochemical and metallogenic associations of As, Sb, ± Bi, and lack of some features typical of SHD’s indicates that Livengood is most comparable to IRGS type deposits and is typical of other such deposits within the host Tintina Gold Belt.


Gold mineralization at Livengood is hosted in a thrust interleaved sequence of Late Proterozoic to Palaeozoic ophiolitic rocks thrust emplaced over a Devonian sequence of sedimentary and volcanic rocks.  Mineralization is related to a ~90 million year old set of monzonite to diorite dikes that intrude the thrust stack along thrust faults.  Mineralization is hosted primarily by Devonian volcanics and Cretaceous dikes, but occurs in all rock types and consists of gold associated with arsenopyrite and to a lesser extent pyrite.  Other associated minerals include stibnite, marcasite, pyrrhotite, and minor to trace amounts of chalcopyrite and sphalerite.


Four stages of alteration are currently recognized.  These include biotite, albite, sericite, and carbonate.  These stages are interpreted to reflect alteration of host rocks by a fluid with decreasing temperature and evolving chemistry over time.


Overall, mineralization and alteration appear to be controlled by the thrust fault architecture and possibly by later normal faults.


The original surface geochemical anomaly in soil that attracted initial exploration in this location probably reflects only a portion of the mineralization present.  Mineralization has been shown to continue down-dip along and/or beneath thrust surfaces and therefore be blind at the surface.  This point along with the fact that the area drilled currently represents only a portion of the original surface geochemical anomaly suggests that the identification of more mineralization over a broader area is likely.


An updated resource estimate has been calculated and is based on all drill data through February 28, 2010.  This new estimate includes gold assay results from 64 drill holes received after the October 2009 resource estimate.  The current resource estimate increases the total tonnes and ounces in the Indicated category and reduces the number of ounces and tonnes in the Inferred category for cutoff grades of 0.3, 0.5, and 0.7 g/t Au.  This change is due to the addition of new lower grade tonnes in the northeastern area of the deposit.  Also, some previously inferred higher grade blocks were downgraded or extinguished through a combination of drilling and use of more rigorous statistical criteria for allowing certain blocks to be included in the resource.


It is important to note that compared to the October 2009 resource estimate, the tonnage and total ounces estimated has increased in the Indicated category and has decreased in the Inferred category for cutoff grades of 0.30, 0.50, and 0.70 g/t Au (Table 19.1).  This change is due, in part, to addition of newly defined lower grade tonnes in the northeastern area of the deposit.  In addition, in this estimate, the composite data were declustered prior to indicator bin selection and data in four octants were required to estimate a block compared to 3 octants in the 2009 estimate.  The net result is that some previously inferred higher grade blocks were downgraded or extinguished through a combination of drilling and use of more rigorous statistical criteria for allowing certain blocks to be included in the resource.  Application of multiple indicator kriging methods have provided new in sights into the character of mineralization and offered an improved, more robust block model for the resource estimation.  Comparison of block model with geologic sections interpreted by ITH geologists (Figures 7.8-7.12) reveals good correspondence (Figures 19.1 to 19.6).  These sections also show the potential of mineralized material to continue to depth, particularly down-dip.



TABLE 19.1

COMPARISON OF RESOURCE ESTIMATES

OCTOBER 2009 AND MARCH 2010


Classification

Au Cutoff (g/t)

Tonnes (millions) 10_2009

Tonnes (millions) 03_2010

Au (g/t)

10_2009

Au (g/t)

03_2010

Million Ounces Au

10_2009

Million Ounces Au

03_2010

Indicated

0.30

525

702

.65

0.60

11

13.5

Inferred

0.30

336

278

.61

0.56

6.6

5.0

Indicated

0.50

297

369

.85

0.78

8.1

9.3

Inferred

0.50

164

112

.84

0.77

4.4

3.0

Indicated

0.70

158

184

1.07

0.98

5.4

5.8

Inferred

0.70

78

56

1.11

0.99

2.8

1.8



It is concluded that a substantial gold resource has been identified at Money Knob and the surrounding area.  Dedicated drilling has continuously enlarged the resource over the past several years.  Current metallurgical studies are underway and results indicate that gold is recoverable through heap leach, and combined mill, CIP, CIL, gravity, and flotation techniques.  Continuation of planned and in-progress metallurgical and ore processing studies will enable assessment of the best material processing and gold recovery techniques.  As results for this work are completed, new cost estimates that incorporate optimized gold recovery techniques will be used for a more comprehensive development plan and economic assessment.


[exhibit1063.jpg]

Figure 19.1.  Block model for section 428625 E.  Grid squares are 200m.


[exhibit1064.jpg]

Figure 19.2.  Block model for section 428850 E.  Grid squares are 200m.


[exhibit1065.jpg]

Figure 19.3.  Block model for section 428925 E.  Grid squares are 200m.

[exhibit1066.jpg]

Figure 19.4.  Block model for section 429075 E.  Grid squares are 200m.


[exhibit1067.jpg]

Figure 19.5.  Block model for section 429525 E.  Grid squares are 200m.


[exhibit1068.jpg]

Figure 19.6.  Block model for section 429675 E.  Grid squares are 200m.

20.0

Recommendations



20.1

Recommended Exploration Program


Exploration of the Livengood project should continue with the aim of advancing the project toward a prefeasibility status.  ITH plans to drill 40,000 m in 2010 to accomplish this goal.  The proposed program is an appropriate amount of drilling for the needs of the project and the time available in the field season.  Activities that will help advance the project in this direction include those listed below.


1.

Continue step out drilling to identify the extent of mineralization, particularly:

a.

to the northeast of the Sunshine Zone,

b.

immediately northwest of the known Sunshine Zone,

c.

down dip of currently identified mineralization in the Sunshine Zone,

d.

to the southwest along the trend of the surface geochemical anomaly, and

e.

to the south of Money Knob and southeast of the Core zone.


2.

Focus infill drilling on areas where Inferred resource blocks can be converted to Indicated resources.


3.

Drill close spaced holes to define a variographic cross in order to better determine the drill spacing required to convert indicated resources into measured resources.


4.

Drill core holes to gather sample material for advance metallurgical testing.


5.

Continue and advance metallurgical, ore characterization, and mineral processing studies.  This should include:

a.

Evaluation of flotation methods,

b.

Production of sufficient volumes of flotation and gravity concentrates to evaluate treatment options for the concentrates,

c.

Quantification of the distribution of preg-robbing carbon in the deposit, and

d.

Use SEM studies to better characterize gold mineralization, its exact mineral association, and relationship to gangue.


6.

Assess geotechnical characteristics of the mineralized zone.


7.

Begin the sterilization process for land that might be covered by facilities.  This should start with surface geochemical surveys to be followed up with drilling on potentially mineralized zones.


8.

Continue and expand environmental base line studies including:

a.

Expansion of surface water quality studies to include additional stations to cover expanded land holdings and measurement of stream flow,

b.

Expansion of aquatic macro fauna studies to cover expanded land holding,

c.

Initiation of subsurface hydrological investigations, and

d.

Installation of meteorological stations on the property.


9.

Complete the combined mill/heap leach preliminary economic analysis that is currently in progress.  This should evaluate the basic economic, logistic, and processing factors for a mining operation at Livengood.



20.2

Budget for 2010


ITH has proposed expenditure of approximately $13 million dollars in 2010 for further evaluation of the Livengood project (Table 20.1).  This budget will be allocated primarily to drilling and geological analysis of the deposit.  The budget is appropriate for the amount of drilling planned and feasible within the time allocated and the company has sufficient funds to accomplish this goal.  The authors recommend implementation of this program in order to accomplish ITH’s goal of advancing the Livengood project.


TABLE 20.1

2010 EXPLORATION BUDGET


Expenditure

2010 $ M

Comments

Land

0.6

Claim and lease fees

Geological and Contract Services

2.5

Contract/consulting fees

Drilling

4.7

Drilling, supplies, surveying, preparation, hole abandonment

Geochemistry

1.5

Rock, soil, drill core and cuttings, prep and assay

Environmental and Metallurgy Studies

2.5

 

Admin and Operations

0.8

Office, salaries, travel, reporting, permitting

TOTAL

12.6

 


 

21.0

References


Albanese, M.D., 1983, Geochemical reconnaissance of the Livengood B-3, B-4, and C-4 quadrangles, Alsak; summary of data on stream-sediment, pan concentrate, and rock samples: Report of Investigations – Alaska. Division of Geological and Geophysical Surveys, v. 83-1, 55 pp.


Arbogast, B., Lee, G., Light, T., 1991, Analytical results and sample locality map of stream-sediment and heavy –mineral-concentrate samples from the Livengood 1 degree X 3 degree Quadrangle, Alaska: U.S. Geological Survey, Open file Report OF 91-0023A and B, 121 pp., 1 sheet.


Athey, J.E., and Craw, P.A., 2004, Geologic maps of the Livengood SW C-3 and SE C-4 Quadrangles, Tolovana mining district, Alaska, Preliminary Interpretive report 2004-3, Division of Geological and Geophysical Surveys., 1 sheet, 24 p.


Athey, J.E., Szumigala, D.J., Newberry, R.J., Werdon, M.B., and Hicks, S.A., 2004, Bedrock geologic map of the Livengood SW C-3 and SE C-4 Quadrangles, Tolovana mining district, Alaska, Preliminary Interpretive report 2004-3, Division of Geological and Geophysical Surveys.


Brooks, A. H., 1916, Preliminary report on the Tolovana district:  U.S. Geological Survey Bulletin 642, p. 201-209.


Burns, L.E., and Liss, S. A., 1999, Portfolio of aeromagnetic and resistivity maps of part of the Livengood mining district, Alaska, central Livengood Quadrangle:  Public-Data File, Alaska Division of Geological and Geophysical Surveys, Report 99-16, 16 pp.


Carew, T., 2010, Livengood Deposit – March 2010 resource update final: independent consultant’s report; Reserva International, LLP.


Chapman, R.M., and Weber, F.R., 1972, Geochemical analyses of bedrock and stream sediment samples from the Livengood Quadrangle, Alaska: U.S. Geological Survey, Open File Report, OF 72-0067, 2 sheets.


Chapman, R.M., Weber, F.R., and Taber, B., 1971, Preliminary map of the Livengood Quadrangle, Alaska; map and explanation: U.S. Geological Survey, Open File Report, OF 71-0066, 2 sheets.


CIM, 2005, CIM Definition Standards – For Mineral Resources and Mineral Reserves; Prepared by the CIM Standing Committee on Reserve Definitions, Adopted by CIM Council on December 11, 2005, 10 pp.


Cobb, E.H., 1972, Metallic mineral resources map of the Livengood Quadrangle, Alaska: U.S. Geological Survey, Misc. field Studies, Report MF-0413, 2 sheets.


Ebert, S., Dodd, S., Miller, L., and Petsel, S., 2000, The Donlin Creek Au-As-Sb-Hg deposit, southwestern Alaska, in Geology and Ore Deposits 2000, The Great Basin and Beyond, Symposium Proceedings, Geological Society of Nevada, ed., Cluer, J.K., Price, J.G., Struhsacker, E.M., Hardyman, R.F., and Morris, C.L., p. 1069-1081.

Foster, R.L., 1968, Potential for lode deposits in the Livengood gold placer district, east-central Alaska: U.S. Geological Survey Circular, Report: C-590, 18 pp.


Giroux Consultants Ltd, 2007, Livengood resource evaluation, Consultants report to ITH, 22 pp.


Giroux Consultants Ltd, 2008, Livengood resource evaluation, Consultants report to ITH, 35 pp.


Giroux Consultants Ltd, 2009, Livengood resource evaluation, January, 2009, Consultants report to ITH, 36 pp.


Goldfarb, 1997, Metallogenic evolution of Alaska:  Economic Geology Monograph 9, Society of Economic Geologists, pp. 4-34.


Goldfarb, R., Hart, C., Miller, M., Miller, L., Farmer, G.L., Groves, D., 2000, The Tintina gold belt – a global perspective, in The Tintina Gold Belt: Concepts, Exploration, and Discoveries, British Columbia and Yukon Chamber of Mines, Cordilleran Roundup Special Volume 2, p. 5-31.


Hazen Research Inc., 2007, Characterization of Livengood Gold Ore, Hazen Project 10504: letter report dated February 7, 2007.


ITH, 2009, International Tower hill Expands Livengood Gold Resource by 64%, October 13, press release, 6 pp.


Karl, S.M., Ager, T.A., Hanneman, K., and Teller, S.D., 1988, Tertiary gold-bearing gravel at Livengood, Alaska, in Geologic Studies in Alaska by the USGS in 1987, U.S.G.S. Circular 1016, p. 61-63.


Klipfel, P., 2006, Summary report on the Livengood Project, Tolovana District, Alaska,  consultants report to ITH, 37 pp.


Klipfel, P. and Giroux, G., 2008a, Summary report on the Livengood Project, Tolovana District, Alaska; consultants report to ITH, 62 pp.


Klipfel, P. and Giroux, G., 2008b, October 2008 Summary report on the Livengood Project, Tolovana District, Alaska; consultants report to ITH, 90 pp.


Klipfel, P., Giroux, G. and Puchner, C, 2008, Summary report on the Livengood Project, Tolovana District, Alaska; technical report to ITH, 72 pp.


Klipfel, P. and Giroux, G., 2009, January 2009 Summary report on the Livengood Project, Tolovana District, Alaska: consultant’s technical report to ITH, 99 pp.


Klipfel, P., Barnes, T., and Pennstrom, 2009a, July 2009 summary report on the Livengood Project, Tolovana District, Alaska: consultant’s technical report to ITH, 120 pp.


Klipfel, Carew, and Pennstrom, 2009b, October 2009 summary report on the Livengood Project, Tolovana District, Alaska: consultant’s technical report to ITH, 144 pp.


Larson, L., 2002, Petrographic report – AngloGold, Samples – Fort Knox, Livengood, and West Pogo, consultants report to AngloGold, 79 pp.


Levinson, A.A., 1974, Introduction to Exploration Geochemistry: Applied Publishing Ltd., 614 pp.


McCoy, D., Newberry, R.J., Layer, P., DiMarchi, J.J., Bakke, A., Masterman, J.S., and Minehane, D.L., 1997, Plutonic related gold deposits of interior Alaska, Society of Economic Geologists, Economic Geology Monograph 9, pp. 91-241.


Newberry, R.J., McCoy, D.T., and Brew, D.A., 1995, Plutonic-hosted gold ores in Alaska: Igneous vs. metamorphic origins: Resource Geology special issue, N0 18, pp. 57-100.


Northern Land Use Research, Inc., 2009, Cultural resources survey of proposed 634-acre Talon Gold core drilling program parcel near Livengood, Interior Alaska; consultants report to ITH, 60 pp.


Northern Land Use Research, Inc., 2010, Cultural resources survey of Talon Gold core drilling program parcels near Livengood, Interior Alaska: 2009 survey results and recommendations:  consultants report to Talon Gold, 190 pp.


Plafker, G. and Berg, H.C., 1994, Overview of the geology and tectonic evolution of Alaska, in Plafker, G. and Berg, H.C. eds., The Geology of Alaska: Geological Society of America, Boulder Co.  The Geology of North America, v. G1, p. 989-1017.


Poulsen, K.H., 1996, Carlin-type gold deposits and their potential occurrence in the Canadian Cordillera: in Current Research, 1996a, Geological Survey of Canada, pp x-x.


Reifenstuhl, R.R., Dover, J.H., Pinney D.S., Newberry, R.J., Clautice, K.H., Liss, S.A., Blodgett, R.B., Bundtzen, T.K., and Weber, F.R., 1997, Geologic map of the Tanana B-1 quadrangle, central Alaska, RI 97-15A, State of Alaska, Department of Natural Resources, Division of Geological & Geophysical Surveys, 17 p., 1 sheet, scale 1:63,360.


Robinson, M.S., 1983, Bedrock geologic map of the C-4 quadrangle, east-central Alaska: Report of Investigations – Alaska. Division of Geological and Geophysical Surveys, vol. 83-4, 1 sheet.


Rose, A.W., Hawkes, H.E., and Webb, J.S., 1979, Geochemistry in Mineral Exploration: Academic Press, 657 pp.


Rudd, J., 1999, Project report of the 1998 geophysical survey data for part for the Livengood mining district,  Alaska, central Livengood Quadrangle:  Alaska Division of Geology and Geophysics, Report, 99-17, 132 pp.


Silberling, N.J., Jones, D.J., Monger, J.W.H., Coney, P.J., Berg, H.C., and Plafker, G., 1994, Lithotectonic terrane map of Alaska and adjacent parts of Canada: Geology of Alaska, Geological Society of America. Vol. G-1, Plate 3.


Sillitoe, R.H., 2009, Geologic model and potential of the Livengood gold prospect, Alaska; consultants report to ITH, 13 pp.

 

Smith, M., 2000, The Tintina gold belt: an emerging gold district in Alaska and Yukon, in The Tintina Gold Belt: Concepts, Exploration, and Discoveries, British Columbia and Yukon Chamber of Mines, Cordilleran Roundup Special Volume 2, p. 1-3.


Smith, T.E., 1983, Bedrock geology of the Livengood C-3 Quadrangle, east-central Alaska: Report of Investigations – Alaska. Division of Geological and Geophysical Surveys, v. 83-5, 1 sheet.


Three Parameters Plus, 2009,  Livengood Project baseline surface water sampling report – 2009, consultants report to Talon Gold US Inc., 389 pp.


Waythomas, C.F., Ten Brink, N.W., Ritter, D.F., 1984, Surficial geologic map of the Livengood B-3, B-4, C-3 and C-4 quadrangles, Yukon-Tanana Upland: Report of Investigations – Alaska, Division of Geological and Geophysical Surveys, v. 84-6, 1 sheet.


22.0

Date and Signature Page


The effective date of this technical report, entitled “March, 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” is March 16, 2010.



Dated: March 18, 2010



Signed:


(signed) Paul Klipfel

[Sealed: CPG#10821]

Dr. Paul Klipfel, Ph.D, CPG#10821





(signed) T. Carew

[Sealed]

Timothy J. Carew P.Geo.





(signed) William J. Pennstrom, Jr.


William Pennstrom, Jr. QP-MMSA




23.0

Certificates of Authors


CERTIFICATE OF PAUL D. KLIPFEL, PH.D.


I, Paul D. Klipfel, Ph.D., do hereby certify that:


1.

I am President of :

Mineral Resource Services, Inc.

4889 Sierra Pine Dr.

Reno, NV 89519


2.

I have graduated from the following Universities with degrees as follows:

a.

San Francisco State University,   

B.A. geology

1978

b.

University of Idaho,

M.S. economic geology

1981

c.

Colorado School of Mines

M.S. mineral economics

1988

d.

Colorado School of Mines

Ph.D. economic geology

1992


3.

I am a member in good standing of the following professional associations:

a.

Society of Mining Engineers

b.

Society of Economic Geologists

c.

Geological Society of America

d.

Society for Applied Geology

e.

American Institute of Professional Geologists

f.

Sigma Xi


4.

I have worked as a mineral exploration geologist for 30+ years since my graduation from San Francisco State 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 professional associations and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.


6.

I am responsible for the preparation of all sections of the technical report titled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” and dated March 16, 2010 (the “Technical Report”) relating to the Livengood property except sections 16 and 17.  I have visited the Livengood property on seven occasions, the most recent being February 20-24, 2010.


7.

Prior to being retained by ITH in 2006, I have not had prior involvement with the property that is the subject of the Technical Report.


8.

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


9.

I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101.


10.

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


Dated this 18th day of March, 2010


(signed)  Paul Klipfel

[Sealed:  AIPG#10821]

Signature of Qualified Person


Paul D. Klipfel, Ph.D, CPG[AIPG]


Print name of Qualified Person




CERTIFICATE OF TIMOTHY J. CAREW


I, Tim Carew, P. Geo. do hereby certify that:


1.

I am the Principal of :

Reserva International LLC

P.O. Box 19848

Reno, NV 89511 USA


2.

I have graduated from the following Universities with degrees as follows:

a.

University of Rhodesia,

B.Sc. Geology

1973

b.

University of Rhodesia,

B.Sc. (Hons) Geology

1976

c.

University of London (RSM)

M.Sc. Mineral Prod. Management

1982


1.

I am a member in good standing of the following professional associations:

a.

Association of Professional Engineers and Geoscientists of British Columbia

b.

Institute of Mining, Metallurgy and Materials

c.

Canadian Institute of Mining and Metallurgy

d.

Society of Mining Engineers


2.

I have worked in mining geology and engineering for over 35 years since my graduation from the University of Rhodesia.


3.

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 professional associations and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.


4.

I am responsible for the preparation of section 17 of the technical report titled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” and dated March 16, 2010 (the “Technical Report”) relating to the Livengood property.  I have visited the Livengood property on three occasions for a total of twenty six days, the most recent being from February 14 -24, 2010.


5.

Prior to being retained by ITH in September, 2009, I have not had prior involvement with the property that is the subject of the Technical Report.


6.

I am not aware of any material fact or material change with respect to the subject matter of Section 17 of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.


7.

I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101.


8.

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


Dated this 18th day of March, 2010


(signed)  T. Carew

[Sealed]

Signature of Qualified Person


Timothy J. Carew P.Geo.


Print name of Qualified Person




CERTIFICATE OF WILLIAM J. PENNSTROM JR.

I, William J. Pennstrom Jr., do hereby certify that:

1.

I am self employed as a Consulting Process Engineer and President of:

Pennstrom Consulting Inc.
2728 Southshire Rd.
Highlands Ranch, CO 80126


2.

I graduated in 1983 with a Bachelors of Science degree in Metallurgical Engineering from the University of Missouri - Rolla, Rolla, Missouri and in 2001 with a Master of Arts degree in Management from Webster University, St. Louis, Missouri.


3.

I am a Founding Registered Member of the Society for Mining, Metallurgy, and Exploration (SME) and am a recognized Qualified Professional (QP) Member with expertise in Metallurgy of the Mining and Metallurgical Society of America (MMSA).


4.

I have worked in the Mineral Processing Industry for a total of 29 years since before, during, and after my attending the University of Missouri.


5.

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


6.

I am responsible for the preparation of section 16 of the technical report titled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” and dated March 16, 2010 (the “Technical Report”) relating to the Livengood property.  I have visited the Livengood Project site for two days during May of 2009.


7.

Prior to being retained by ITH in May, 2009, I have not had prior involvement with the property that is the subject of the Technical Report.


8.

I am not aware of any material fact or material change with respect to the subject matter of Section 16 of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.


9.

I am independent of the issuer applying all of the tests per Section 1.5 of NI 43-101.


10.

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


Dated the 18th day of March, 2010.


(signed)  William J. Pennstrom Jr.


Signature of Qualified Person


William J. Pennstrom Jr. QP-MMSA


Print name of Qualified Person







24.0

Appendices


APPENDIX 1:  Claim/Property Information


Owner

File Number

Tenure Name

Date Acquired

MTRS Location

Alaska State Lease

Alaska Mental Health Land Trust

9400248

AMHLT - ML

1-Jul-2004

F008N005W

Federal Patented Claims

Griffin heirs

MS 1990, Patent 1041576

Mastodon

18-Jan-2007

F008N005W

Griffin heirs

MS 1990, Patent 1041576

Piedmont

18-Jan-2007

F008N005W

Griffin heirs

MS 1990,

Patent 1041576

Yukon

18-Jan-2007

F008N005W

Federal Unpatented Claims

Richard Hudson

55469

ANNE

21-Apr-2003

F008N005W24

Richard Hudson

55466

BLACK ROCK

21-Apr-2003

F008N005W24

Richard Hudson

55471

BRIDGET

21-Apr-2003

F008N005W24

Richard Hudson

55453

DOROTHEA

21-Apr-2003

F008N005W23

Richard Hudson

55470

EILEEN

21-Apr-2003

F008N005W24

Richard Hudson

55455

FOSTER

21-Apr-2003

F008N005W24

Richard Hudson

55454

LENORA

21-Apr-2003

F008N005W23

Richard Hudson

55459

NICKIE

21-Apr-2003

F008N005W24

Richard Hudson

55464

OLD SMOKY

21-Apr-2003

F008N005W23

Richard Hudson

55468

PATRICIA

21-Apr-2003

F008N005W13

Richard Hudson

55460

PATRICK

21-Apr-2003

F008N005W23

Richard Hudson

55458

SAUNDERS

21-Apr-2003

F008N005W23

Richard Hudson

55452

SHARON

21-Apr-2003

F008N005W23

Richard Geraghty

55462

SUNSHINE #1

21-Apr-2003

F008N005W23

Richard Geraghty

55463

SUNSHINE #2

21-Apr-2003

F008N005W23

Richard Hudson

55467

TRAPLINE

21-Apr-2003

F008N005W24

Richard Hudson

55457

TWERPIT

21-Apr-2003

F008N005W24

Richard Hudson

55456

VANCE

21-Apr-2003

F008N005W24

Richard Hudson

55461

WHITE ROCK

21-Apr-2003

F008N005W23

Richard Hudson

55465

WITTROCK

21-Apr-2003

F008N005W23

Ronald Tucker

37580

Lillian No. 1

30-Sep-1968

F008N005E22

Ronald Tucker

37581

Satellite

30-Sep-1968

F008N005E22

Ronald Tucker

37582

Nickel Bench R.L.*

30-Jun-1972

F008N005E22 & 15

Ronald Tucker

37583

The Nickel*

12-Aug-1965

F008N005E22

Ronald Tucker

37584

Overlooked*

6-Sep-1975

F008N005E22

Ronald Tucker

37585

The Lad*

12-Aug-1965

F008N005E22

State Claims

 Karl Hanneman and Bergelin Family Trust

330936

LUCKY 55

14-May-1981

F009N004W33

Karl Hanneman and Bergelin Family Trust

330937

LUCKY 56

14-May-1981

F009N004W33

Karl Hanneman and Bergelin Family Trust

330938

LUCKY 64

13-May-1981

F009N004W32 F009N004W33

Karl Hanneman and Bergelin Family Trust

330939

LUCKY 65

14-May-1981

F009N004W33

Karl Hanneman and Bergelin Family Trust

330940

LUCKY 66

14-May-1981

F009N004W33

Karl Hanneman and Bergelin Family Trust

330941

LUCKY 72

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330942

LUCKY 73

13-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330943

LUCKY 74

13-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330944

LUCKY 75

14-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330945

LUCKY 76

14-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330946

LUCKY 82

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330947

LUCKY 83

13-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330948

LUCKY 84

13-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330949

LUCKY 85

14-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330950

LUCKY 86

14-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330951

LUCKY 91

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330952

LUCKY 92

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330953

LUCKY 93

13-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330954

LUCKY 94

13-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330955

LUCKY 95

14-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330956

LUCKY 96

14-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330957

LUCKY 101

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330958

LUCKY 102

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330959

LUCKY 103

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330960

LUCKY 104

12-May-1981

F008N004W05

Karl Hanneman and Bergelin Family Trust

330961

LUCKY 105

12-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330962

LUCKY 106

12-May-1981

F008N004W04

Karl Hanneman and Bergelin Family Trust

330963

LUCKY 202

13-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330964

LUCKY 203

13-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330965

LUCKY 204

15-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330966

LUCKY 205

13-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330967

LUCKY 206

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330968

LUCKY 207

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330969

LUCKY 208

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330970

LUCKY 302

13-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330971

LUCKY 303

13-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330972

LUCKY 304

15-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330973

LUCKY 305

13-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330974

LUCKY 306

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330975

LUCKY 307

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330976

LUCKY 308

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330977

LUCKY 404

15-May-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

330978

LUCKY 405

13-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

330979

LUCKY 406

14-May-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

338477

LUCKY 198

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338478

LUCKY 199

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338479

LUCKY 295

17-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338480

LUCKY 296

17-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338481

LUCKY 297

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338482

LUCKY 298

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338483

LUCKY 299

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338484

LUCKY 392

21-Sep-1981

F008N005W11

Karl Hanneman and Bergelin Family Trust

338485

LUCKY 395

18-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338486

LUCKY 396

18-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338487

LUCKY 397

18-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338488

LUCKY 398

18-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338489

LUCKY 399

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338490

LUCKY 400

23-Sep-1981

F008N004W07 F008N004W08

Karl Hanneman and Bergelin Family Trust

338491

LUCKY 491

21-Sep-1981

F008N005W11

Karl Hanneman and Bergelin Family Trust

338492

LUCKY 492

21-Sep-1981

F008N005W11

Karl Hanneman and Bergelin Family Trust

338493

LUCKY 493

21-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338494

LUCKY 494

21-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338495

LUCKY 495

18-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338496

LUCKY 496

18-Sep-1981

F008N005W12

Karl Hanneman and Bergelin Family Trust

338497

LUCKY 497

18-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338498

LUCKY 498

18-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338499

LUCKY 499

17-Sep-1981

F008N004W07

Karl Hanneman and Bergelin Family Trust

338500

LUCKY 500

23-Sep-1981

F008N004W07 F008N004W08

Karl Hanneman and Bergelin Family Trust

338501

LUCKY 504

10-Sep-1981

F008N004W08

Karl Hanneman and Bergelin Family Trust

338502

LUCKY 505

10-Sep-1981

F008N004W09

Karl Hanneman and Bergelin Family Trust

338503

LUCKY 589

21-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338504

LUCKY 590

21-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338505

LUCKY 591

21-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338506

LUCKY 592

21-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338507

LUCKY 593

21-Sep-1981

F008N005W13

Karl Hanneman and Bergelin Family Trust

338508

LUCKY 594

21-Sep-1981

F008N005W13

Karl Hanneman and Bergelin Family Trust

338509

LUCKY 595

18-Sep-1981

F008N005W13

Karl Hanneman and Bergelin Family Trust

338510

LUCKY 596

18-Sep-1981

F008N005W13

Karl Hanneman and Bergelin Family Trust

338511

LUCKY 597

18-Sep-1981

F008N004W18

Karl Hanneman and Bergelin Family Trust

338512

LUCKY 598

18-Sep-1981

F008N004W18

Karl Hanneman and Bergelin Family Trust

338513

LUCKY 599

17-Sep-1981

F008N004W18

Karl Hanneman and Bergelin Family Trust

338514

LUCKY 689

22-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338515

LUCKY 690

22-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338516

LUCKY 691

22-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338517

LUCKY 692

22-Sep-1981

F008N005W14

Karl Hanneman and Bergelin Family Trust

338518

LUCKY 693

22-Sep-1981

F008N005W13

Karl Hanneman and Bergelin Family Trust

338519

LUCKY 694

22-Sep-1981

F008N005W13

Karl Hanneman and Bergelin Family Trust

338520

LUCKY 697

18-Sep-1981

F008N004W18

Karl Hanneman and Bergelin Family Trust

338521

LUCKY 698

18-Sep-1981

F008N004W18

Karl Hanneman and Bergelin Family Trust

338522

LUCKY 699

17-Sep-1981

F008N004W18

Karl Hanneman and Bergelin Family Trust

347943

LC 407

5-Jun-1982

F008N004W18

Karl Hanneman and Bergelin Family Trust

347945

LC 502

5-Jun-1982

F008N004W08

Karl Hanneman and Bergelin Family Trust

347946

LC 503

5-Jun-1982

F008N004W08

Karl Hanneman and Bergelin Family Trust

347947

LC 506

7-Jun-1982

F008N004W09

Karl Hanneman and Bergelin Family Trust

347948

LC 507

7-Jun-1982

F008N004W09

Karl Hanneman and Bergelin Family Trust

347949

LC 600

5-Jun-1982

F008N004W17 F008N004W18

Karl Hanneman and Bergelin Family Trust

347950

LC 601

5-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347951

LC 602

5-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347952

LC 603

5-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347953

LC 604

6-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347954

LC 605

6-Jun-1982

F008N004W16

Karl Hanneman and Bergelin Family Trust

347955

LC 695

10-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347956

LC 696

10-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347957

LC 700

6-Jun-1982

F008N004W17 F008N004W18

Karl Hanneman and Bergelin Family Trust

347958

LC 701

6-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347959

LC 702

6-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347960

LC 703

6-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347961

LC 704

6-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347962

LC 790

12-Jun-1982

F008N005W14

Karl Hanneman and Bergelin Family Trust

347963

LC 791

12-Jun-1982

F008N005W14

Karl Hanneman and Bergelin Family Trust

347964

LC 792

11-Jun-1982

F008N005W14

Karl Hanneman and Bergelin Family Trust

347965

LC 793

11-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347966

LC 794

11-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347967

LC 795

10-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347968

LC 796

10-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347969

LC 797

10-Jun-1982

F008N004W18

Karl Hanneman and Bergelin Family Trust

347970

LC 798

9-Jun-1982

F008N004W18

Karl Hanneman and Bergelin Family Trust

347971

LC 799

8-Jun-1982

F008N004W18

Karl Hanneman and Bergelin Family Trust

347972

LC 800

8-Jun-1982

F008N004W17 F008N004W18

Karl Hanneman and Bergelin Family Trust

347973

LC 801

8-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347974

LC 802

8-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347975

LC 803

8-Jun-1982

F008N004W17

Karl Hanneman and Bergelin Family Trust

347976

LC 891

12-Jun-1982

F008N005W14

Karl Hanneman and Bergelin Family Trust

347977

LC 892

11-Jun-1982

F008N005W14

Karl Hanneman and Bergelin Family Trust

347978

LC 893

11-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347979

LC 894

11-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

347980

LC 895

10-Jun-1982

F008N005W13

Karl Hanneman and Bergelin Family Trust

348802

LC 688

4-Jun-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348803

LC 787

4-Jun-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348804

LC 788

4-Jun-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348805

LC 884

31-May-1982

F008N005W16

Karl Hanneman and Bergelin Family Trust

348805

LC 884

31-May-1982

F008N005W16

Karl Hanneman and Bergelin Family Trust

348806

LC 885

31-May-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348807

LC 886

25-May-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348808

LC 887

2-Jun-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348809

LC 888

4-Jun-1982

F008N005W15

Karl Hanneman and Bergelin Family Trust

348810

LC 984

31-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348811

LC 985

31-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348812

LC 986

25-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348813

LC 987

4-Jun-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348814

LC 1083

30-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348815

LC 1084

30-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348816

LC 1085

30-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348817

LC 1086

25-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348818

LC 1183

29-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348819

LC 1184

29-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348820

LC 1185

29-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348821

LC 1186

25-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348822

LC 1282

28-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348823

LC 1283

28-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348824

LC 1284

28-May-1982

F008N005W21

Karl Hanneman and Bergelin Family Trust

348825

LC 1285

28-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348826

LC 1286

26-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348827

LC 1287

26-May-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348828

LC 1288

2-Jun-1982

F008N005W22

Karl Hanneman and Bergelin Family Trust

348829

LC 1382

27-May-1982

F008N005W28

Karl Hanneman and Bergelin Family Trust

348830

LC 1383

27-May-1982

F008N005W28

Karl Hanneman and Bergelin Family Trust

348831

LC 1384

27-May-1982

F008N005W28

Karl Hanneman and Bergelin Family Trust

348832

LC 1385

27-May-1982

F008N005W27

Karl Hanneman and Bergelin Family Trust

361326

LUCKY 90

24-Oct-1983

F008N004W06

Karl Hanneman and Bergelin Family Trust

361327

LUCKY 100

24-Oct-1983

F008N004W06

Karl Hanneman and Bergelin Family Trust

361328

LUCKY 200

24-Oct-1983

F008N004W07

Karl Hanneman and Bergelin Family Trust

361329

LUCKY 294

28-Oct-1983

F008N005W12

Karl Hanneman and Bergelin Family Trust

361330

LUCKY 300

24-Oct-1983

F008N004W07

Karl Hanneman and Bergelin Family Trust

361331

LUCKY 394

28-Oct-1983

F008N005W12

Karl Hanneman and Bergelin Family Trust

361332

LUCKY 401

24-Oct-1983

F008N004W08

Karl Hanneman and Bergelin Family Trust

361333

LUCKY 402

24-Oct-1983

F008N004W08

Karl Hanneman and Bergelin Family Trust

361334

LUCKY 403

24-Oct-1983

F008N004W08

Karl Hanneman and Bergelin Family Trust

361335

LUCKY 501

24-Oct-1983

F008N004W08

Talon Gold Alaska Inc

669377

LVG 1

02/20/10

F8N4W9SESE

Talon Gold Alaska Inc

669378

LVG 2

02/20/10

F8N4W16NWNE

Talon Gold Alaska Inc

669379

LVG 3

02/20/10

F8N4W16NWSW

Talon Gold Alaska Inc

669380

LVG 4

02/20/10

F8N4W16NWSE

Talon Gold Alaska Inc

669381

LVG 5

02/20/10

F9N4W20NW

Talon Gold Alaska Inc

669382

LVG 6

02/20/10

F9N4W20NE

Talon Gold Alaska Inc

669383

LVG 7

02/20/10

F9N4W21NW

Talon Gold Alaska Inc

669384

LVG 8

02/20/10

F9N4W21NE

Talon Gold Alaska Inc

669385

LVG 9

02/20/10

F9N4W22NW

Talon Gold Alaska Inc

669386

LVG 10

02/20/10

F9N4W22NE

Talon Gold Alaska Inc

669387

LVG 11

02/20/10

F9N4W20SW

Talon Gold Alaska Inc

669388

LVG 12

02/20/10

F9N4W20SE

Talon Gold Alaska Inc

669389

LVG 13

02/20/10

F9N4W21SW

Talon Gold Alaska Inc

669390

LVG 14

02/20/10

F9N4W21SE

Talon Gold Alaska Inc

669391

LVG 15

02/20/10

F9N4W22SW

Talon Gold Alaska Inc

669392

LVG 16

02/20/10

F9N4W22SE

Talon Gold Alaska Inc

669393

LVG 17

02/20/10

F9N5W25NW

Talon Gold Alaska Inc

669394

LVG 18

02/20/10

F9N5W25NE

Talon Gold Alaska Inc

669395

LVG 19

02/20/10

F9N4W30NW

Talon Gold Alaska Inc

669396

LVG 20

02/20/10

F9N4W30NE

Talon Gold Alaska Inc

669397

LVG 21

02/20/10

F9N4W29NW

Talon Gold Alaska Inc

669398

LVG 22

02/20/10

F9N4W29NE

Talon Gold Alaska Inc

669399

LVG 23

02/20/10

F9N5W25SW

Talon Gold Alaska Inc

669400

LVG 24

02/20/10

F9N5W25SE

Talon Gold Alaska Inc

669401

LVG 25

02/20/10

F9N4W30SW

Talon Gold Alaska Inc

669402

LVG 26

02/20/10

F9N4W30SE

Talon Gold Alaska Inc

669403

LVG 27

02/20/10

F9N4W29SW

Talon Gold Alaska Inc

669404

LVG 28

02/20/10

F9N4W29SE

Talon Gold Alaska Inc

669405

LVG 29

02/20/10

F9N5W35NW

Talon Gold Alaska Inc

669406

LVG 30

02/20/10

F9N5W35NE

Talon Gold Alaska Inc

669407

LVG 31

02/20/10

F9N5W36NW

Talon Gold Alaska Inc

669408

LVG 32

02/20/10

F9N5W36NE

Talon Gold Alaska Inc

669409

LVG 33

02/20/10

F9N5W35SW

Talon Gold Alaska Inc

669410

LVG 34

02/20/10

F9N5W35SE

Talon Gold Alaska Inc

669411

LVG 35

02/20/10

F9N5W36SW

Talon Gold Alaska Inc

669412

LVG 36

02/20/10

F9N5W36SE

Talon Gold Alaska Inc

669413

LVG 37

02/20/10

F8N5W3NW

Talon Gold Alaska Inc

669414

LVG 38

02/20/10

F8N5W3NE

Talon Gold Alaska Inc

669415

LVG 39

02/20/10

F8N5W3SW

Talon Gold Alaska Inc

669416

LVG 40

02/20/10

F8N5W3SE

Talon Gold Alaska Inc

669417

LVG 41

02/20/10

F9N4W27NW

Talon Gold Alaska Inc

669418

LVG 42

02/20/10

F9N4W27NE

Talon Gold Alaska Inc

669419

LVG 43

02/20/10

F9N4W27SW

Talon Gold Alaska Inc

669420

LVG 44

02/20/10

F9N4W27SE

Talon Gold Alaska Inc

669421

LVG 45

02/20/10

F9N4W34NW

Talon Gold Alaska Inc

669422

LVG 46

02/20/10

F9N4W34NE

Talon Gold Alaska Inc

669423

LVG 47

02/20/10

F9N4W34SW

Talon Gold Alaska Inc

669424

LVG 48

02/20/10

F9N4W34SE

Talon Gold Alaska Inc

669425

LVG 49

02/20/10

F8N4W4NE

Talon Gold Alaska Inc

669426

LVG 50

02/20/10

F8N4W3NW

Talon Gold Alaska Inc

669427

LVG 51

02/20/10

F8N4W3NE

Talon Gold Alaska Inc

669428

LVG 52

02/20/10

F8N4W2NW

Talon Gold Alaska Inc

669429

LVG 53

02/20/10

F8N4W2NE

Talon Gold Alaska Inc

669430

LVG 54

02/20/10

F8N4W4SE

Talon Gold Alaska Inc

669431

LVG 55

02/20/10

F8N4W3SW

Talon Gold Alaska Inc

669432

LVG 56

02/20/10

F8N4W3SE

Talon Gold Alaska Inc

669433

LVG 57

02/20/10

F8N4W2SW

Talon Gold Alaska Inc

669434

LVG 58

02/20/10

F8N4W2SE

Talon Gold Alaska Inc

669435

LVG 59

02/20/10

F8N4W10NW

Talon Gold Alaska Inc

669436

LVG 60

02/20/10

F8N4W10NE

Talon Gold Alaska Inc

669437

LVG 61

02/20/10

F8N4W11NW

Talon Gold Alaska Inc

669438

LVG 62

02/20/10

F8N4W11NE

Talon Gold Alaska Inc

669439

LVG 63

02/20/10

F8N4W10SW

Talon Gold Alaska Inc

669440

LVG 64

02/20/10

F8N4W10SE

Talon Gold Alaska Inc

669441

LVG 65

02/20/10

F8N4W11SW

Talon Gold Alaska Inc

669442

LVG 66

02/20/10

F8N4W11SE

Talon Gold Alaska Inc

669443

LVG 67

02/20/10

F8N4W16NE

Talon Gold Alaska Inc

669444

LVG 68

02/20/10

F8N4W15NW

Talon Gold Alaska Inc

669445

LVG 69

02/20/10

F8N4W15NE

Talon Gold Alaska Inc

669446

LVG 70

02/20/10

F8N4W14NW

Talon Gold Alaska Inc

669447

LVG 71

02/20/10

F8N4W14NE

Talon Gold Alaska Inc

669448

LVG 72

02/20/10

F8N4W16SW

Talon Gold Alaska Inc

669449

LVG 73

02/20/10

F8N4W16SE

Talon Gold Alaska Inc

669450

LVG 74

02/20/10

F8N4W15SW

Talon Gold Alaska Inc

669451

LVG 75

02/20/10

F8N4W15SE

Talon Gold Alaska Inc

669452

LVG 76

02/20/10

F8N4W14SW

Talon Gold Alaska Inc

669453

LVG 77

02/20/10

F8N4W14SE

Talon Gold Alaska Inc

669454

LVG 78

02/20/10

F8N4W21NW

Talon Gold Alaska Inc

669455

LVG 79

02/20/10

F8N4W21NE

Talon Gold Alaska Inc

669456

LVG 80

02/20/10

F8N4W22NW

Talon Gold Alaska Inc

669457

LVG 81

02/20/10

F8N4W22NE

Talon Gold Alaska Inc

669458

LVG 82

02/20/10

F8N4W23NW

Talon Gold Alaska Inc

669459

LVG 83

02/20/10

F8N4W23NE

Talon Gold Alaska Inc

669460

LVG 84

02/20/10

F8N4W21SW

Talon Gold Alaska Inc

669461

LVG 85

02/20/10

F8N4W21SE

Talon Gold Alaska Inc

669462

LVG 86

02/20/10

F8N4W22SW

Talon Gold Alaska Inc

669463

LVG 87

02/20/10

F8N4W22SE

Talon Gold Alaska Inc

669464

LVG 88

02/20/10

F8N4W23SW

Talon Gold Alaska Inc

669465

LVG 89

02/20/10

F8N4W23SE


* - Placer claim


Note: Meridian Township Range and Section (MTRS) Location is the Federal land location system.  Example F006S013E12 is a section of land located in the Fairbanks Meridian, Township 6 South, Range 13 East, Section 12.






APPENDIX 2: List Of Drill Holes



HOLE

EASTING

NORTHING

ELEVATION

HOLE LENGTH (m)

BAF-1       

430060.00

7266021.00

518.20

213.40

BAF-2       

430073.00

7266149.00

525.50

152.40

BAF-3       

429760.00

7266096.00

506.00

150.90

BAF-4       

430073.00

7265881.00

476.70

216.40

BAF-5       

430078.00

7265765.00

460.20

189.90

BAF-6       

429745.00

7265979.00

515.10

134.10

BAF-7       

430056.00

7266034.00

518.20

304.80

BAF-8       

430342.00

7266042.00

524.90

152.40

L-1         

429726.00

7265450.00

503.00

31.00

L-2         

429350.00

7265457.00

506.00

73.00

L-3         

429050.00

7265715.00

468.00

46.00

L-4         

429045.00

7265688.00

470.00

20.00

L-5         

428910.00

7265675.00

454.00

70.00

L-6         

428805.00

7265640.00

441.00

70.00

LC-TR-01    

428883.00

7266132.00

358.10

91.40

LC-TR-02    

428859.00

7266041.00

358.10

68.60

MK-04-01    

428734.38

7265596.00

421.50

109.70

MK-04-02    

428492.13

7265738.00

361.60

305.70

MK-04-03    

428674.66

7265520.50

412.20

208.80

MK-04-04    

428547.66

7265813.50

354.40

137.80

MK-04-TP1   

429594.00

7265670.00

510.00

2.00

MK-04-TP2   

429583.00

7265653.00

512.00

2.00

MK-04-TR1   

429541.09

7265537.00

524.70

34.00

MK-04-TR2E  

429598.03

7265763.00

514.80

85.00

MK-04-TR2S  

429598.03

7265763.00

514.80

20.00

MK-04-TR2W  

429597.06

7265763.50

514.80

85.00

MK-04-TR3   

429602.97

7265704.00

516.40

33.40

MK-04-TR5   

429570.00

7265621.00

512.00

15.00

MK-06-05    

429099.00

7266101.00

403.00

305.10

MK-06-06    

429299.00

7266298.00

405.00

205.40

MK-06-07    

428772.31

7265845.00

412.80

276.50

MK-06-08    

428915.28

7265897.00

408.70

288.30

MK-06-09    

427614.00

7264251.00

223.70

124.70

MK-06-10    

427533.00

7264335.00

228.20

10.40

MK-06-11    

427691.00

7264430.00

242.30

17.10

MK-07-12    

428915.28

7265897.00

408.70

282.90

MK-07-13    

428773.31

7265847.50

412.80

351.10

MK-07-14    

428774.81

7265846.00

412.80

44.80

MK-07-15    

428774.81

7265849.00

412.80

281.60

MK-07-16    

430220.00

7265985.00

531.30

332.80

MK-07-17    

428773.41

7265621.50

427.70

421.80

MK-07-18    

428853.63

7265780.00

431.80

301.10

MK-07-19    

429002.63

7265704.00

458.40

436.20

MK-07-20    

428851.72

7265720.00

435.30

244.30

MK-07-21    

428925.81

7265760.50

440.20

310.00

MK-07-22    

428703.31

7265764.00

408.50

382.80

MK-07-23    

429075.75

7265779.50

458.80

290.20

MK-07-24    

429529.81

7265631.00

508.90

372.20

MK-07-25    

428399.63

7265253.00

368.20

330.40

MK-07-26    

429900.00

7265470.00

438.00

28.40

MK-08-27    

429592.59

7265927.30

499.90

201.80

MK-08-28    

429518.31

7266005.70

485.90

229.20

MK-08-29    

429896.00

7265778.70

470.10

266.70

MK-08-30    

428891.91

7265737.88

438.70

345.20

MK-08-31    

429142.44

7265606.61

479.10

376.40

MK-08-32    

429186.50

7265431.15

474.10

400.00

MK-08-33    

429066.25

7265091.11

427.50

300.00

MK-09-34

428771.9

7265545

427.53

296.27

MK-09-35

428851.1

7265491

437.15

276.45

MK-09-36

428782.5

7265215

409.49

697.69

MK-09-37

429109.1

7265406

463.73

527.3

MK-09-38

429251.3

7265388

477.33

215.8

MK-09-39

429524.8

7265999

487.82

309.37

MK-08-TR01  

428869.84

7266061.44

342.40

21.30

MK-08-TR02  

428834.63

7266031.09

338.80

28.00

MK-08-TR03  

428834.63

7266031.09

338.80

4.10

MK-08-TR04  

428869.84

7266061.44

342.40

26.10

MK-1        

428945.00

7265820.00

442.00

76.00

MK-2        

428825.00

7265850.00

427.00

77.00

MK-3        

429500.00

7266190.00

465.00

28.00

MK-4        

429493.00

7266117.00

478.00

15.20

MK-4B       

429493.00

7266117.00

478.00

106.70

MK-5        

428660.00

7265925.00

368.00

0.00

MK-6        

428680.00

7265940.00

367.00

0.00

MK-RC-0001  

428996.00

7265778.00

449.00

321.60

MK-RC-0002  

429001.81

7265854.50

426.10

335.30

MK-RC-0003  

428703.19

7265998.50

335.90

222.50

MK-RC-0004  

428612.00

7265921.00

343.50

274.00

MK-RC-0005  

428561.81

7265841.50

350.00

269.80

MK-RC-0006  

429045.69

7265695.50

460.70

353.60

MK-RC-0007  

428846.00

7265843.00

423.60

286.50

MK-RC-0008  

428925.00

7265691.60

445.90

213.40

MK-RC-0009  

428997.91

7265632.10

456.50

246.90

MK-RC-0010  

428547.69

7265470.90

393.20

240.80

MK-RC-0011  

428925.69

7265626.30

448.00

225.60

MK-RC-0012  

428997.00

7265544.70

459.50

307.90

MK-RC-0013  

428624.19

7265480.10

403.20

225.60

MK-RC-0014  

428176.91

7265590.70

357.30

217.90

MK-RC-0015  

428323.09

7265696.50

349.20

195.10

MK-RC-0016  

428319.50

7265542.50

367.70

134.10

MK-RC-0017  

428779.09

7265774.00

423.20

297.20

MK-RC-0018  

428710.91

7265834.00

396.90

252.40

MK-RC-0019  

428550.00

7265925.00

330.00

54.90

MK-RC-0020  

428549.69

7265909.80

331.50

213.40

MK-RC-0021  

428470.00

7265852.10

330.50

213.40

MK-RC-0022  

428847.91

7265920.70

399.80

280.40

MK-RC-0023  

428849.31

7265622.60

437.70

288.00

MK-RC-0024  

428697.81

7265630.00

413.90

207.30

MK-RC-0025  

428920.91

7265909.10

404.50

213.40

MK-RC-0026  

428622.91

7265760.00

385.80

167.60

MK-RC-0027  

428559.09

7265703.80

381.60

129.50

MK-RC-0028  

428844.53

7266105.70

350.00

93.00

MK-RC-0029  

429057.91

7265856.70

432.50

256.00

MK-RC-0030  

428777.19

7265548.20

425.80

243.80

MK-RC-0031  

428926.47

7265548.00

447.20

303.30

MK-RC-0032  

428554.91

7265783.10

363.50

91.40

MK-RC-0033  

428849.41

7265566.50

437.10

335.30

MK-RC-0034  

429073.81

7265553.40

467.90

365.80

MK-RC-0035  

429071.91

7265468.10

467.90

330.70

MK-RC-0036  

429001.59

7265463.40

453.20

257.90

MK-RC-0037  

429149.41

7265558.70

483.50

295.70

MK-RC-0038  

428784.09

7265918.70

392.50

234.70

MK-RC-0039  

428999.09

7265410.20

450.70

277.40

MK-RC-0040  

428927.38

7265860.42

418.90

335.30

MK-RC-0041  

428850.69

7265504.08

437.50

262.10

MK-RC-0042  

428778.56

7265473.11

425.90

274.30

MK-RC-0043  

428940.28

7265472.30

446.40

265.20

MK-RC-0044  

428698.09

7265487.46

417.60

237.70

MK-RC-0045  

428922.00

7265395.50

441.10

317.00

MK-RC-0046  

429084.03

7265622.27

470.50

323.10

MK-RC-0047  

429152.56

7265477.69

475.40

326.80

MK-RC-0048  

429144.00

7265399.25

466.90

350.50

MK-RC-0049  

428697.66

7265404.66

416.90

274.30

MK-RC-0050  

429225.06

7265481.30

488.50

350.80

MK-RC-0051  

428699.75

7265549.36

416.60

239.30

MK-RC-0052  

428625.53

7265847.83

366.60

249.90

MK-RC-0053  

428543.97

7265549.99

393.20

204.20

MK-RC-0054  

429297.22

7265483.50

493.40

341.40

MK-RC-0055  

428706.44

7265926.89

368.90

262.10

MK-RC-0056  

428477.38

7265559.88

384.50

195.10

MK-RC-0057  

429374.31

7265486.84

504.80

304.80

MK-RC-0058  

428700.06

7266242.25

334.30

213.40

MK-RC-0059  

429450.22

7265478.31

511.60

262.10

MK-RC-0060  

429077.13

7265328.34

453.50

336.80

MK-RC-0061  

429225.78

7265326.36

468.30

302.10

MK-RC-0062  

429150.22

7265323.46

460.50

312.40

MK-RC-0063  

429299.63

7265329.00

474.40

359.70

MK-RC-0064  

429072.38

7265252.31

445.30

363.30

MK-RC-0065  

429302.81

7265425.01

484.80

346.00

MK-RC-0066  

429156.28

7265243.08

452.10

304.80

MK-RC-0067  

429155.28

7265174.77

448.20

349.00

MK-RC-0068  

429227.25

7265403.32

476.20

396.20

MK-RC-0069  

429147.53

7265098.42

434.70

256.00

MK-RC-0070  

429452.13

7265548.90

509.90

378.00

MK-RC-0071  

428928.31

7265326.22

435.50

301.80

MK-RC-0072  

428997.91

7265323.84

444.90

262.10

MK-RC-0073  

429521.63

7265549.72

513.20

335.30

MK-RC-0074  

428474.03

7265632.47

377.30

158.50

MK-RC-0075  

428477.16

7265481.85

386.50

243.80

MK-RC-0076  

429151.06

7265033.41

425.50

285.00

MK-RC-0077  

428475.91

7265930.18

312.10

152.40

MK-RC-0078  

429225.91

7265026.63

428.20

298.70

MK-RC-0079  

428399.41

7265859.17

320.00

161.50

MK-RC-0080  

428626.69

7265396.63

402.60

262.10

MK-RC-0081  

428841.59

7265250.01

419.90

243.80

MK-RC-0082  

429073.56

7265037.48

421.60

317.00

MK-RC-0083  

428911.13

7265169.42

420.60

300.20

MK-RC-0084  

429224.53

7265250.71

458.20

374.90

MK-RC-0085  

429599.09

7265554.41

510.80

326.10

MK-RC-0086  

429377.88

7265391.25

491.40

36.60

MK-RC-0087  

429148.47

7264949.83

417.20

254.50

MK-RC-0088  

429003.38

7265008.70

413.50

115.80

MK-RC-0089  

429003.38

7265008.70

413.50

374.90

MK-RC-0090  

429070.13

7264946.92

413.30

201.20

MK-RC-0091  

429007.06

7264947.97

407.40

283.50

MK-RC-0092  

429377.88

7265391.25

491.40

344.42

MK-RC-0093  

429226.13

7265103.86

439.00

323.09

MK-RC-0094  

429750.00

7265475.00

504.00

327.66

MK-RC-0095  

429600.00

7266000.00

513.00

268.22

MK-RC-0096

428780.91

7265217.91

410.00

262.13

MK-RC-0097

429897.41

7265464.74

447.73

237.74

MK-RC-0098

428925.00

7265112.11

415.29

219.46

MK-RC-0099

429296.66

7264946.83

419.03

268.22

MK-RC-0100

429214.03

7264951.65

418.33

274.32

MK-RC-0101

429294.00

7265027.91

429.73

295.66

MK-RC-0102

429296.25

7265176.16

453.02

274.32

MK-RC-0103

429229.09

7265170.67

449.21

306.63

MK-RC-0103a

429225.00

7265175.00

449.78

6.10

MK-RC-0104

429159.75

7264696.23

386.59

128.02

MK-RC-0105

429138.44

7264694.52

387.76

190.50

MK-RC-0106

429071.19

7265245.22

445.85

335.28

MK-RC-0107

429296.03

7264725.13

378.26

224.03

MK-RC-0108

429296.72

7265103.06

442.38

271.27

MK-RC-0109

428934.3

7265034.7

409.7

284.99

MK-RC-0110

428996.0

7265174.3

430.5

353.57

MK-RC-0111

429446.9

7265637.8

504.2

303.58

MK-RC-0112

429376.1

7265625.5

500.4

356.62

MK-RC-0113

429296.7

7265617.7

493.5

334.37

MK-RC-0114

429229.3

7265624.3

486.7

307.85

MK-RC-0115

428694.1

7264869.6

369.1

263.96

MK-RC-0116

428636.1

7264959.9

369.9

295.66

MK-RC-0117

428775.0

7265085.7

397.6

182.88

MK-RC-0118

428761.0

7264784.0

370.4

289.56

MK-RC-0119

428774.3

7265081.3

397.7

225.55

MK-RC-0120

428610.5

7264794.5

353.3

313.94

MK-RC-0121

428693.6

7265241.3

401.2

231.65

MK-RC-0122

428773.4

7264966.5

385.0

295.66

MK-RC-0123

428694.8

7265247.4

401.6

332.84

MK-RC-0124

428627.5

7265097.7

380.2

301.75

MK-RC-0125

428764.9

7265308.5

414.6

306.93

MK-RC-0126

428851.3

7265319.4

425.8

263.65

MK-RC-0127

428617.2

7265252.4

391.9

307.85

MK-RC-0128

429302.2

7265768.1

476.9

320.04

MK-RC-0129

428846.6

7265012.9

398.6

262.13

MK-RC-0130

429150.7

7265775.7

462.4

286.51

MK-RC-0131

428848.7

7264870.7

386.8

260.6

MK-RC-0132

428928.8

7264939.7

401.0

220.98

MK-RC-0133

428845.8

7265095.3

407.4

326.14

MK-RC-0134

428627.3

7265628.6

404.3

182.88

MK-RC-0135

429376.7

7265704.6

492.0

301.75

MK-RC-0136

428854.1

7265401.7

432.0

297.18

MK-RC-0137

429466.4

7265926.7

482.7

280.42

MK-RC-0138

428992.3

7265089.2

421.8

269.75

MK-RC-0139

429368.1

7265988.9

456.9

289.56

MK-RC-0140

428700.3

7265164.6

396.1

318.52

MK-RC-0141

429304.2

7265999.5

443.2

198.12

MK-RC-0142

428686.4

7265103.8

388.4

280.42

MK-RC-0143

430273.2

7266146

542.38

301.75

MK-RC-0144

429677.2

7265407

513.99

310.9

MK-RC-0145

430421

7266012

477.81

311.51

MK-RC-0146

429818.9

7265396

473.5

256.03

MK-RC-0147

429245.4

7264877

408.21

350.52

MK-RC-0148

430417.4

7266143

504.45

307.85

MK-RC-0149

429826

7265555

464.11

170.69

MK-RC-0150

429380

7264892

412.05

193.55

MK-RC-0151

429673.3

7265549

504.29

266.7

MK-RC-0152

430124.4

7265924

486.84

306.93

MK-RC-0153

429372.8

7265019

429.49

262.13

MK-RC-0154

429373.2

7265177

454.74

344.42

MK-RC-0155

429984.4

7265930

483.41

300.23

MK-RC-0156

429670.2

7265842

503.93

316.99

MK-RC-0157

429374.2

7265251

466.4

301.75

MK-RC-0158

429672

7265916

507.81

324.61

MK-RC-0159

429825.1

7265848

491.69

272.8

MK-RC-0160

429673.9

7266070

503.64

316.99

MK-RC-0161

429458.4

7264796

389.32

242.93

MK-RC-0162

429524.4

7266078

480.71

263.65

MK-RC-0163

429376.4

7264800

389.75

325.22

MK-RC-0164

429302

7264795

391.9

334.67

MK-RC-0165

429746.2

7265846

500.61

249.94

MK-RC-0166

429740.3

7265918

509.18

240.79

MK-RC-0167

429676.3

7265703

497.76

286.51

MK-RC-0168

429356.5

7264949

419.58

312.42

MK-RC-0169

430124.6

7266079

531.45

339.85

MK-RC-0170

429526

7265862

494.19

301.75

MK-RC-0171

429454.4

7264940

413.51

276.76

MK-RC-0172

429602.6

7264877

391.76

298.7

MK-RC-0173

429520.3

7264951

412.08

242.32

MK-RC-0174

429602.4

7265860

502.14

321.56

MK-RC-0175

428413.1

7265552

377.14

198.12

MK-RC-0176

429447.4

7265018

430.48

248.41

MK-RC-0177

429969.4

7266055

502.98

278.89

MK-RC-0178

429302.7

7264870

407.19

316.99

MK-RC-0179

428545.1

7265409

393.24

73.15

MK-RC-0180

429670.9

7265997

507.25

347.47

MK-RC-0181

429372.5

7265122

446.07

262.13

MK-RC-0182

428817.4

7265678

432.97

274.32

MK-RC-0183

429301.9

7265247

463.9

332.23

MK-RC-0184

428545.1

7265409

393.24

268.22

MK-RC-0185

429599

7266016

499.65

289.56

MK-RC-0186

429176.4

7265350

465.01

365.76

MK-RC-0187

429971.6

7265855

470.45

317.3

MK-RC-0188

429602.9

7266076

496.19

268.22

MK-RC-0189

429451.7

7265098

440.82

233.17

MK-RC-0190

429889.9

7265852

483.59

286.51

MK-RC-0191

430205.1

7265555

391.17

170.69

MK-RC-0192

429522.7

7265105

435.33

300.53

MK-RC-0193

430351.3

7265706

413.67

368.81

MK-RC-0194

429524.1

7265026

425.6

251.46

MK-RC-0195

430349.7

7266235

529.9

319.74

MK-RC-0196

430493.3

7265844

427.6

359.66

MK-RC-0197

428480.5

7265398

385.59

313.94

MK-RC-0198

430637.3

7265919

451.05

335.28

MK-RC-0199

430343.5

7266154

523.27

341.38

MK-RC-0200

428400.6

7265467

377.58

277.37

MK-RC-0201

429533.4

7265188

450.49

298.7

MK-RC-0202

430278.3

7266088

539.29

365.76

MK-RC-0203

430501.2

7265922

443.48

365.76

MK-RC-0204

429597.9

7265251

455.69

213.36

MK-RC-0206

429829

7265995

508.01

402.34

MK-RC-0207

429974.1

7265999

500.76

399.29

MK-RC-0208

429525.2

7265255

465.99

262.13

MK-RC-0209

429900.3

7265926

492.04

408.43

MK-RC-0210

429448.8

7265248

467.86

278.89

MK-RC-0211

429754.4

7266003

510.63

402.34

MK-RC-0212

429598.3

7265193

441.6

214.88

MK-RC-0213

429901.2

7266006

504.06

411.48

MK-RC-0214

429599.7

7265095

424.97

201.17

MK-RC-0215

429756.2

7266074

505.81

396.24

MK-RC-0216

429680.1

7265175

427.24

216.41

MK-RC-0217

429818.7

7265922

501.72

396.24

MK-RC-0218

429600.5

7265024

413.88

224.03

MK-RC-0219

429598.1

7264949

403.46

356.62

MK-RC-0220

429602.7

7265774

501.13

396.24

MK-RC-0221

430029.1

7265465

401.76

353.57

MK-RC-0222

429530.8

7265926

492.43

399.29

MK-RC-0223

429678.5

7265773

499.54

341.38

MK-RC-0224

429467.2

7265933

481.72

376.43

MK-RC-0225

429968.1

7265308

397.59

347.47

MK-RC-0226

429747.1

7265700

485.81

341.38

MK-RC-0227

429898.7

7265164

386.41

256.03

MK-RC-0228

429527.7

7266161

463.33

254.51

MK-RC-0229

429605.8

7265705

503.25

237.74

MK-RC-0231

429457.4

7266073

466.34

335.28

MK-RC-0233

429454.4

7266001

473.3

350.52

MK-RC-0234

429606.5

7265701

503.74

423.67

MK-RC-0236

429600.7

7265622

505.76

396.24

MK-RC-0237

429519.4

7266000

487.49

396.24

MK-RC-0239

429673.2

7265628

497.41

426.72

MK-RC-0240

429598

7266145

480.22

320.02

MK-RC-0242

429750.3

7265775

492.54

426.72

MK-RC-0244

429673.5

7266147

483.75

396.22

MK-RC-0250

429824.2

7265702

471.17

353.57

MK-RC-0256

429525.6

7265703

503.85

426.72

MK-RC-0257

429527

7265768

499.12

274.32

MK-RC-0258

429824

7266074

503

375

MK-RC-0259

429743

7265549

493

315

MK-RC-0260

430126

7266147

526

396

MK-RC-0261

430283

7266008

520

390

MK-RC-0262

429751

7265403

490

378

MK-RC-0263

430200

7266142

545

378

MK-RC-0264

429518

7265490

520

424

MK-RC-0265

430419

7266071

491

396

MK-RC-0266

430340

7266001

500

366

MK-RC-0267

429527

7265760

499

366

MK-RC-0268

430511

7266079

479

366

MK-RC-0269

429450

7265873

484

372

MK-RC-0270

429597

7265932

502

152

MK-RC-0271

430492

7266002

462

280

MK-RC-0272

430345

7265919

472

387

MK-RC-0273

430055

7266004

505

338

MK-RC-0274

430422

7265923

456

379

MK-RC-0275

430353

7266078

512

320

MK-RC-0276

428623

7265321

397

311

MK-RC-0277

430957

7265913

496

305

MK-RC-0278

428776

7265175

407

244

MK-RC-0279

428095

7265434

335

332

MK-RC-0280

429671

7265349

493

360

MK-RC-0281

427898

7265539

291

287

MK-RC-0282

431107

7265779

480

347

MK-RC-0283

428030

7265567

318

215

MK-RC-0284

429747

7265331

471

396

MK-RC-0285

430050

7266159

507

354

MK-RC-0286

429827

7265465

470

326

MK-RC-0287

428403

7265396

376

274

MK-RC-0288

430273

7266233

533

396

MK-RC-0289

428847

7266471

358

355

MK-RC-0290

430052

7265866

468

351

MK-RC-0291

429756

7264873

370

235

MK-RC-0292

429380

7266155

432

187

MK-RC-0293

429600

7265475

534

369

MK-RC-0294

429375

7266000

462

319

MK-RC-0295

429825

7265325

451

99

MK-RC-0296

429823

7265323

451

274

MK-RC-0297

429375

7265925

466

315

MK-RC-0298

429675

7265475

524

369

MK-RC-0299

429750

7265250

438

265

MK-RC-0300

429450

7266150

449

152

MK-RC-0301

429675

7265250

444

223

MK-RC-0302

429375

7265325

479

308

MK-RC-0303

429375

7266075

445

256

MN-1        

428864.00

7266045.00

358.10

106.70

MN-2        

428864.00

7266045.00

358.10

106.70

MN-3        

428745.00

7266065.00

335.30

106.70

TL-10       

428183.00

7265586.00

358.00

79.00

TL-11       

429528.00

7266520.00

370.00

105.00

TL-12       

429223.00

7266654.00

318.00

200.00

TL-13       

429054.00

7266654.00

307.00

150.00

TL-14       

427780.00

7265504.00

266.50

124.00

TL-6        

433265.00

7269380.00

277.00

43.90

TL-7        

428443.00

7266477.00

317.00

101.00

TL-8        

428443.00

7266477.00

317.00

192.00

TL-9        

428443.00

7266477.00

317.00

105.00

EX-2 4 exhibit2.htm CONSENT OF TIMOTHY J. CAREW Consent of Tim Carew

 

TIM CAREW, P.Geo.


Consent and Certificate

of Professional (Qualified Person)


British Columbia Securities Commission

Pacific Centre

12th Floor, 701 West Georgia Street

Vancouver, British Columbia

CANADA  V7Y 1L2

The Toronto Stock Exchange

Suite 2700 – 650 West Georgia Street

Vancouver, British Columbia

CANADA  V6B 4N9

Alberta Securities Commission

4 th Floor, 300 – 5th Avenue S.W.

Calgary, Alberta

CANADA  T2P 3C4

Ontario Securities Commission

Suite 1903 - 20 Queen Street West

Toronto, Ontario

CANADA  M5H 3S8

International Tower Hill Mines Ltd.

Suite 1920 – 1188 West Georgia Street

Vancouver, British Columbia

CANADA  V6E 4A2

 


Re:  International Tower Hill Mines Ltd. (the “Issuer”)


I, Timothy J. Carew, P.Geo., of P.O. Box 19848, Reno, Nevada, U.S.A. 89511, have prepared, and am the author of, section 17 of the report entitled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” dated March 16, 2010 (the “Report”).


1.                                                                                                                        

I hereby consent to:


(a)

the public filing of the Report on SEDAR and EDGAR, in the public files of the Securities Commissions of British Columbia, Alberta and Ontario, with the Toronto Stock Exchange and with the United States Securities and Exchange Commission;


(b)

the use of and reliance upon the Report in connection with the disclosure in the Issuer’s Press Release dated March 10, 2010 and Material Change Report dated April 21, 2010 (collectively, the “Disclosures”) and for any and all required regulatory filings, acceptances or approvals in connection with the use of, and reference to, the Mineral Resource Estimate for the “Livengood” Property, Alaska as set out in the Report (including any subsequent press releases and material change reports); and


(c)

the inclusion of extracts from, or a summary of, the Report in the Disclosures.


2.                                                                                                                        

I hereby consent to the use of my name “Timothy J. Carew”, and to the use of the name of “Reserva International LLC”, a private geological and mining consulting services firm, of which I am the Principal, in the Disclosures and in any subsequent press releases or material change reports in reference to the Report.


3.                                                                                                                        

I hereby certify that I have read the Disclosures and that the Disclosures fairly and accurately represent the information contained in the Report.


Dated this 23rd day of April, 2010



(signed) Tim Carew


Timothy J. Carew, P.Geo.


EX-3 5 exhibit3.htm CERTIFICATE OF TIMOTHY J. CAREW Certificate of Timothy J. Carew



CERTIFICATE OF TIMOTHY J. CAREW


I, Tim Carew, P. Geo. do hereby certify that:


1.

I am the Principal of :

Reserva International LLC

P.O. Box 19848

Reno, NV 89511 USA


2.

I have graduated from the following Universities with degrees as follows:

a.

University of Rhodesia,

B.Sc. Geology

1973

b.

University of Rhodesia,

B.Sc. (Hons) Geology

1976

c.

University of London (RSM)

M.Sc. Mineral Prod. Management

1982


3.

I am a member in good standing of the following professional associations:

a.

Association of Professional Engineers and Geoscientists of British Columbia

b.

Institute of Mining, Metallurgy and Materials

c.

Canadian Institute of Mining and Metallurgy

d.

Society of Mining Engineers


4.

I have worked in mining geology and engineering for over 35 years since my graduation from the University of Rhodesia.


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 professional associations and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.


6.

I am responsible for the preparation of section 17 of the technical report titled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” and dated March 16, 2010 (the “Technical Report”) relating to the Livengood property.  I have visited the Livengood property on three occasions for a total of twenty six days, the most recent being from February 14 -24, 2010.


7.

Prior to being retained by ITH in September, 2009, I have not had prior involvement with the property that is the subject of the Technical Report.


8.

I am not aware of any material fact or material change with respect to the subject matter of Section 17 of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.


9.

I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101.


10.

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


Dated this 18th day of March, 2010


(signed)  T. Carew

[Sealed]

Signature of Qualified Person


Timothy J. Carew P.Geo.


Print name of Qualified Person





EX-4 6 exhibit4.htm CONSENT OF PAUL D. KLIPFEL Consent of Paul D. Klipfel

 

PAUL D. KLIPFEL Ph.D, CPG


Consent and Certificate

of Professional (Qualified Person)


British Columbia Securities Commission

Pacific Centre

12th Floor, 701 West Georgia Street

Vancouver, British Columbia

CANADA  V7Y 1L2

The Toronto Stock Exchange

Suite 2700 – 650 West Georgia Street

Vancouver, British Columbia

CANADA  V6B 4N9


Alberta Securities Commission

4 th Floor, 300 – 5th Avenue S.W.

Calgary, Alberta

CANADA  T2P 3C4

Ontario Securities Commission

Suite 1903 - 20 Queen Street West

Toronto, Ontario

CANADA  M5H 3S8

International Tower Hill Mines Ltd.

Suite 1920 – 1188 West Georgia Street

Vancouver, British Columbia

CANADA  V6E 4A2

 


Re:  International Tower Hill Mines Ltd. (the “Issuer”)


I, Paul D. Klipfel, Ph.D, CPG (AIPG) #10821, of 4889 Sierra Pine Drive, Reno, Nevada, U.S.A. 89519, have prepared, and am the author of, all sections and/or parts of the report entitled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” dated March 16, 2010 (the “Report”), other than sections 16 and 17 thereof.


1.                                                                                                                        

I hereby consent to:


(a)

the public filing of the Report on SEDAR and EDGAR, in the public files of the Securities Commissions of British Columbia, Alberta and Ontario, with the Toronto Stock Exchange and with the United States Securities and Exchange Commission;


(b)

the use of and reliance upon the Report in connection with the disclosure in the Issuer’s Press Release dated March 10, 2010 and Material Change Report dated April 21, 2010 (collectively, the “Disclosures”) and for any and all required regulatory filings, acceptances or approvals in connection with the use of, and reference to, the Report (including any subsequent press releases and material change reports); and


(c)

the inclusion of extracts from, or a summary of, the Report in the Disclosures.


2.                                                                                                                        

I hereby consent to the use of my name “Paul D. Klipfel”, and to the use of the name of “Mineral Resource Services Inc.”, a private geological consulting firm, of which I am the President, in the Disclosures and in any subsequent press releases or material change reports in reference to the Report.


3.                                                                                                                        

I hereby certify that I have read the Disclosures and that the Disclosures fairly and accurately represent the information contained in the Report.


Dated this 23rd day of April, 2010



(signed) Paul Klipfel


Paul D. Klipfel, Ph.D, CPG (AIPG) #10821.


EX-5 7 exhibit5.htm CERTIFICATE OF PAUL D. KLIPFEL Certificate of Paul D. Klipfel


CERTIFICATE OF PAUL D. KLIPFEL, PH.D.


I, Paul D. Klipfel, Ph.D., do hereby certify that:


1.

I am President of :

Mineral Resource Services, Inc.

4889 Sierra Pine Dr.

Reno, NV 89519


2.

I have graduated from the following Universities with degrees as follows:

a.

San Francisco State University,   

B.A. geology

1978

b.

University of Idaho,

M.S. economic geology

1981

c.

Colorado School of Mines

M.S. mineral economics

1988

d.

Colorado School of Mines

Ph.D. economic geology

1992


3.

I am a member in good standing of the following professional associations:

a.

Society of Mining Engineers

b.

Society of Economic Geologists

c.

Geological Society of America

d.

Society for Applied Geology

e.

American Institute of Professional Geologists

f.

Sigma Xi


4.

I have worked as a mineral exploration geologist for 30+ years since my graduation from San Francisco State 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 professional associations and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.


6.

I am responsible for the preparation of all sections of the technical report titled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” and dated March 16, 2010 (the “Technical Report”) relating to the Livengood property except sections 16 and 17.  I have visited the Livengood property on seven occasions, the most recent being February 20-24, 2010.


7.

Prior to being retained by ITH in 2006, I have not had prior involvement with the property that is the subject of the Technical Report.


8.

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


9.

I am independent of the issuer applying all of the tests in section 1.4 of NI 43-101.


10.

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


Dated this 18th day of March, 2010


(signed)  Paul Klipfel

[Sealed:  AIPG#10821]

Signature of Qualified Person


Paul D. Klipfel, Ph.D, CPG[AIPG]


Print name of Qualified Person





EX-7 8 exhibit6.htm CONSENT OF WILLIAM J. PENNSTROM, JR. Consent of William J. Pennstrom, JR.

 

WILLIAM J. PENNSTROM, JR.


Consent and Certificate

of Professional (Qualified Person)


British Columbia Securities Commission

Pacific Centre

12th Floor, 701 West Georgia Street

Vancouver, British Columbia

CANADA  V7Y 1L2

The Toronto Stock Exchange

Suite 2700 – 650 West Georgia Street

Vancouver, British Columbia

CANADA  V6B 4N9

Alberta Securities Commission

4 th Floor, 300 – 5th Avenue S.W.

Calgary, Alberta

CANADA  T2P 3C4

Ontario Securities Commission

Suite 1903 - 20 Queen Street West

Toronto, Ontario

CANADA  M5H 3S8

International Tower Hill Mines Ltd.

Suite 1920 – 1188 West Georgia Street

Vancouver, British Columbia

CANADA  V6E 4A2

 


Re:  International Tower Hill Mines Ltd. (the “Issuer”)


I, William J. Pennstrom, Jr., of 2728 Southshire Road, Highlands Ranch, Colorado, USA 80126, have prepared, and am the author of, section 16 of the report entitled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” dated March 16, 2010 (the “Report”).


1.                                                                                                                        

I hereby consent to:


(a)

the public filing of the Report on SEDAR and EDGAR, in the public files of the Securities Commissions of British Columbia, Alberta and Ontario, with the Toronto Stock Exchange and with the United States Securities and Exchange Commission;


(b)

the use of and reliance upon the Report in connection with the disclosure in the Issuer’s Press Release dated March 10, 2010 and Material Change Report dated April 21, 2010 (collectively, the “Disclosures”) and for any and all required regulatory filings, acceptances or approvals in connection with the use of, and reference to, the Mineral Resource Estimate for the “Livengood” Property, Alaska as set out in the Report (including any subsequent press releases and material change reports); and


(c)

the inclusion of extracts from, or a summary of, the Report in the Disclosures.


2.                                                                                                                        

I hereby consent to the use of my name “William J. Pennstrom, Jr.”, and to the use of the name of “Pennstrom Consulting Inc.”, a private process engineering consulting firm, of which I am the President, in the Disclosures and in any subsequent press releases or material change reports in reference to the Report.


3.                                                                                                                        

I hereby certify that I have read the Disclosures and that the Disclosures fairly and accurately represent the information contained in the Report.


Dated this 23rd day of April, 2010



(signed) William J. Pennstrom Jr.


William J. Pennstrom, Jr.


EX-8 9 exhibit7.htm CERTIFICATE OF WILLIAM J. PENNSTROM, JR. Certificate of William J. Pennstrom JR.



CERTIFICATE OF WILLIAM J. PENNSTROM JR.

I, William J. Pennstrom Jr., do hereby certify that:

1.

I am self employed as a Consulting Process Engineer and President of:

Pennstrom Consulting Inc.
2728 Southshire Rd.
Highlands Ranch, CO 80126


2.

I graduated in 1983 with a Bachelors of Science degree in Metallurgical Engineering from the University of Missouri - Rolla, Rolla, Missouri and in 2001 with a Master of Arts degree in Management from Webster University, St. Louis, Missouri.


3.

I am a Founding Registered Member of the Society for Mining, Metallurgy, and Exploration (SME) and am a recognized Qualified Professional (QP) Member with expertise in Metallurgy of the Mining and Metallurgical Society of America (MMSA).


4.

I have worked in the Mineral Processing Industry for a total of 29 years since before, during, and after my attending the University of Missouri.


5.

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


6.

I am responsible for the preparation of section 16 of the technical report titled “March 2010 Summary Report on the Livengood Project, Tolovana District, Alaska” and dated March 16, 2010 (the “Technical Report”) relating to the Livengood property.  I have visited the Livengood Project site for two days during May of 2009.


7.

Prior to being retained by ITH in May, 2009, I have not had prior involvement with the property that is the subject of the Technical Report.


8.

I am not aware of any material fact or material change with respect to the subject matter of Section 16 of the Technical Report that is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.


9.

I am independent of the issuer applying all of the tests per Section 1.5 of NI 43-101.



10.

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


Dated the 18th day of March, 2010.


(signed)  William J. Pennstrom Jr.


Signature of Qualified Person


William J. Pennstrom Jr. QP-MMSA


Print name of Qualified Person





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