A C A HOWE INTERNATIONAL LIMITED MICROMINE ...alhambraresources.com/news/releases/ACA Howe_MCS_Dombraly...MICROMINE CONSULTING SERVICES James Hogg, BSc, MSc, MAIG Marta Sostre, BSc,

A C A HOWE INTERNATIONAL LIMITED


MICROMINE CONSULTING SERVICES

TECHNICAL REPORT

INTRODUCTION AND

RESOURCE ESTIMATION FOR THE

DOMBRALY GOLD DEPOSIT IN

NORTH-CENTRAL KAZAKHSTAN

for

ALHAMBRA RESOURCES LTD

by

ACA HOWE INTERNATIONAL LIMITED John G Langlands, BSc, FGS, FIMMM, C Eng

James Emberton, BSc, FIMMM, C Eng

and MICROMINE CONSULTING SERVICES

James Hogg, BSc, MSc, MAIG Marta Sostre, BSc, MSc, MAUSIMM

22 March 2012 Berkhamsted

Herts, UK


TABLE OF CONTENTS

Page SUMMARY…………………………………………………………………………………………….…………… I

1 INTRODUCTION .......................................................................................................................................... 1

1.1 TERMS OF REFERENCE ............................................................................................................................... 2

1.2 ACA HOWE INTERNATIONAL LIMITED ................................................................................................... 2

1.3 MICROMINE CONSULTING SERVICES ..................................................................................................... 2

1.4 UNITS ............................................................................................................................................................... 3

2 RELIANCE ON OTHER EXPERTS ............................................................................................................. 3

3 PROPERTY DESCRIPTION AND LOCATION .......................................................................................... 3

3.1 LOCATION ...................................................................................................................................................... 3

3.2 LICENCE AND TENURE ................................................................................................................................ 5

4 REGIONAL GEOLOGY ................................................................................................................................ 7

5 ADJACENT PROPERTIES ........................................................................................................................... 8

6 DOMBRALY PROJECT ................................................................................................................................ 9

6.1 DOMBRALY - PROPERTY DESCRIPTION AND LOCATION ................................................................. 15

6.2 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ... 15

6.3 DOMBRALY - HISTORY .............................................................................................................................. 16

6.3.1 CHRONOLOGICAL OPERATIONS ................................................................................................... 16

6.3.2 HISTORICAL MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ......................... 18

6.3.3 HISTORICAL ESTIMATES OF MINERAL POTENTIAL ................................................................. 19

6.4 DOMBRALY - GEOLOGICAL SETTING .................................................................................................... 20

6.4.1 DOMBRALY - REGIONAL GEOLOGY.............................................................................................. 20

6.4.2 DOMBRALY - LOCAL AND PROPERTY GEOLOGY ...................................................................... 20

6.5 DOMBRALY - DEPOSIT TYPES .................................................................................................................. 20

6.6 DOMBRALY - MINERALISATION ............................................................................................................. 21

6.7 DOMBRALY - EXPLORATION.................................................................................................................... 22

6.8 DRILLING ...................................................................................................................................................... 26

6.8.1 TRENCHING......................................................................................................................................... 28

6.8.2 BULK DENSITY ................................................................................................................................... 28

6.9 DOMBRALY - SAMPLE PREPARATION, ANALYSES AND SECURITY ................................................ 29

6.10 DOMBRALY - DATA VERIFICATION ........................................................................................................ 32

6.10.1 QA/QC ANALYSIS ............................................................................................................................... 32

6.10.2 INTRODUCTION .................................................................................................................................. 32

6.10.3 QUALITY CONTROL SUBMISSION .................................................................................................. 33

6.10.4 QUALITY CONTROL SAMPLE MATERIALS .................................................................................. 33

6.10.5 QUALITY CONTROL ASSESSMENT................................................................................................. 35

6.10.6 QA/QC ASSESSMENT .......................................................................................................................... 36

6.10.7 CONCLUSIONS OF QAQC STUDY .................................................................................................... 48

6.10.8 RECOMMENDATIONS ....................................................................................................................... 48

6.11 DOMBRALY - MINERAL PROCESSING AND METALLURGICAL TESTING ...................................... 49

6.12 MCS OCTOBER-DECEMBER 2011 MINERAL RESOURCE ESTIMATES .............................................. 52

6.12.1 SOFTWARE USED ............................................................................................................................... 52

6.12.2 INPUT DATA SUMMARY ................................................................................................................... 52

6.12.3 INPUT DATA ........................................................................................................................................ 52

6.12.4 DATA VALIDATION ............................................................................................................................ 53

6.12.5 DESCRIPTIVE AND CLASSICAL STATISTICS................................................................................ 53

6.12.6 AU DISTRIBUTION.............................................................................................................................. 53

6.12.7 NATURAL CUT OFF ............................................................................................................................ 55

6.12.8 DOMAIN INTERPRETATION AND MODELLING ........................................................................... 56

6.12.9 DOMAIN STATISTICS......................................................................................................................... 65

6.12.10 TOP CUTS ............................................................................................................................................. 65

6.12.11 COMPOSITES ....................................................................................................................................... 66

6.12.12 GEOSTATISTICS ................................................................................................................................. 67

6.12.13 MCS OCTOBER-DECEMBER 2011 IDW BLOCK MODEL ESTIMATION..................................... 67

6.12.14 BLOCK MODEL ATTRIBUTES .......................................................................................................... 70

6.12.15 RESOURCE CLASSIFICATION.......................................................................................................... 72

6.12.16 MODEL VALIDATION ........................................................................................................................ 72

6.12.17 NOVEMBER 2011 IDW RESOURCE ESTIMATE REPORTING ...................................................... 74

6.13 DOMBRALY - ADJACENT PROPERTIES .................................................................................................. 87

6.14 DOMBRALY - OTHER RELEVANT DATA AND INFORMATION ........................................................... 87

6.15 DOMBRALY - INTERPRETATION AND CONCLUSIONS ........................................................................ 87

6.16 DOMBRALY - RECOMMENDATIONS ....................................................................................................... 89

7 REFERENCES AND OTHER SOURCES OF INFORMATION ................................................................ 92

8 DATE AND SIGNATURE PAGES .............................................................................................................. 94


LIST OF TABLES

TABLE 1. SARY-ARKA LICENCE COORDINATES .......................................................................................... 4

TABLE 2. SUMMARY OF DOMBRALY EXPLORATION 1969 TO 1972.......................................................... 16

TABLE 3. SUMMARY OF DOMBRALY SAMPLING 1969 TO 1972 ................................................................. 17

TABLE 4. DOMBRALY RESOURCE AND ‘RESERVE’ ESTIMATE 2007 ...................................................... 18

TABLE 5. DOMBRALY OFFICIAL RESOURCE AND ‘RESERVE’ ESTIMATE 2007 ................................... 19

TABLE 6. SUMMARY OF DOMBRALY WORK 2002-2005 ............................................................................... 22

TABLE 7. 2010-2011 DDH SIGNIFICANT INTERCEPTS ................................................................................. 24

TABLE 8. 2010-2011 RC SIGNIFICANT INTERCEPTS .................................................................................... 25

TABLE 9. SUMMARY OF DOMBRALY SAMPLING QUANTITIES ............................................................... 27

TABLE 10. NAME AND GRADE OF CRM SAMPLES ........................................................................................ 34

TABLE 11. NUMBER OF QA/QC SAMPLES IN THE DIFFERENT 2010-2011DRILLING CAMPAIGNS ...... 36

TABLE 12. NUMBER OF QA/QC SAMPLES USED IN PRE-2010 EXPLORATION CAMPAIGNS (EW*

TRENCHES). ....................................................................................................................................... 36

TABLE 13. DOMBRALY MARCH 2011 RESOURCE ESTIMATE SAMPLE SUMMARY ............................... 52

TABLE 14. SUMMARY OF BASIC PARAMETERS FOR DOMBRALY AU INPUT DATASET ...................... 53

TABLE 15. DOMBRALY MINERALISED DOMAIN RAW DATA DESCRIPTIVE STATISTICS ................... 65

TABLE 16. DOMBRALY MINERALISED DOMAIN COMPOSITE DATA DESCRIPTIVE STATISTICS...... 67

TABLE 17. DOMBRALY IDW SEARCH ELLIPSOID PARAMETERS ............................................................. 69

TABLE 18. DOMBRALY INTERPOLATION PARAMETERS ........................................................................... 70

TABLE 19. LOW GRADE STOCKPILE BLOCK MODEL ATTRIBUTES......................................................... 70

TABLE 20. PIT INFILL BLOCK MODEL ATTRIBUTES ................................................................................... 71

TABLE 21. IN SITU BLOCK MODEL ATTRIBUTES ......................................................................................... 71

TABLE 22. BASIC SUMMARY STATISTICS FOR DOMBRALY BULK DENSITY SAMPLES ...................... 71

TABLE 23. COMPARISON OF DOMAIN RAW, COMP AND BLOCK MODEL MEANS ................................ 73

TABLE 24. COMPARISON OF DOMAIN WIREFRAME AND BLOCK MODEL VOLUMES ......................... 73

TABLE 25. DOMBRALY NOVEMBER 2011 TOTAL RESOURCES .................................................................. 75

LIST OF FIGURES

FIGURE 1: LOCATION MAP SHOWING ALHAMBRA LICENCE, PROSPECTS AND SURROUNDING

GOLD PROJECTS ................................................................................................................................ 6 FIGURE 2: GEOLOGICAL MAP OF THE DOMBRALY AREA WITH SOIL AND ROCK CHIP SAMPLES . 10 FIGURE 3: GEOLOGICAL MAP OF THE DOMBRALY AREA WITH CROSS SECTION LINES (L16, L40)

AND LOND SECTION LINE (A-B) .................................................................................................... 11 FIGURE 4: CROSS SECTION OF THE DOMBRALY DEPOSIT ALONG LINE L16 ........................................ 12 FIGURE 5: CROSS SECTION OF THE DOMBRALY AREA ALONG LINE L40.............................................. 13 FIGURE 6: LONG SECTION OF THE DOMBRALY AREA, LINE A-B............................................................. 14 FIGURE 7: SAMPLING PROCESSING SCHEME ............................................................................................... 31 FIGURE 8: LEGEND FOR STANDARD GRAPHS ............................................................................................... 36 FIGURE 9: DIAMOND DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE STD=21.600 PPM

AU ........................................................................................................................................................ 37 FIGURE 10: DIAMOND DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE STD=13.600 PPM

AU ........................................................................................................................................................ 37 FIGURE 11: DIAMOND DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE STD= 5.530 PPM

AU ........................................................................................................................................................ 38 FIGURE 12: DIAMOND DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE STD= 1.460 PPM

AU ........................................................................................................................................................ 38 FIGURE 13: DIAMOND DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE STD=0.520 PPM

AU ........................................................................................................................................................ 39 FIGURE 14: REVERSE CIRCULATION DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE

STD=21.600 PPM AU .......................................................................................................................... 39 FIGURE 15: REVERSE CIRCULATION DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE

STD=13.600 PPM AU .......................................................................................................................... 40 FIGURE 16: REVERSE CIRCULATION DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE

STD=5.530 PPM AU ............................................................................................................................ 40 FIGURE17: REVERSE CIRCULATION DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE

STD=1.460 PPM AU ............................................................................................................................ 41 FIGURE 18: REVERSE CIRCULATION DRILLHOLE SAMPLE GRAPH PLOT OF STANDARD SAMPLE

STD=0.520 PPM AU ............................................................................................................................ 41 FIGURE 19: SCATTERPLOT COMPARISON BETWEEN ORIGINAL DIAMOND DRILLHOLE FIRE ASSAY

SAMPLES AND PULP DUPLICATES. .............................................................................................. 42


FIGURE20: SCATTERPLOT COMPARISON BETWEEN ORIGINAL REVERSE CIRCULATION

DRILLHOLE FIRE ASSAY SAMPLES AND PULP DUPLICATES. ............................................... 43 FIGURE21: SCATTERPLOT COMPARISON BETWEEN ORIGINAL GRAB SAMPLES FIRE ASSAY

SAMPLES AND PULP DUPLICATES ANALYSED IN THE ORIGINAL LAB. .............................. 44 FIGURE22: SCATTERPLOT COMPARISON BETWEEN ORIGINAL GRAB SAMPLES FIRE ASSAY

SAMPLES AND PULP DUPLICATES ANALYSED IN AN EXTERNAL CONTROL LAB. ........... 44 FIGURE23: SCATTERPLOT COMPARISON BETWEEN ORIGINAL DIAMOND DRILLING DRILLHOLE

FIRE ASSAY SAMPLES AND LAB DUPLICATES. ......................................................................... 45 FIGURE24: SCATTERPLOT COMPARISON BETWEEN ORIGINAL REVERSE CIRCULATION DRILLING

DRILLHOLE FIRE ASSAY SAMPLES AND LAB DUPLICATES. ................................................. 46 FIGURE25: GRAPH PLOT FOR DDH 2010 BLANK SAMPLES ......................................................................... 47 FIGURE26: GRAPH PLOT FOR RC 2010 BLANK SAMPLES ............................................................................ 47 FIGURE 27: LOG NORMAL HISTOGRAM DISTRIBUTION FOR LOW GRADE STOCKPILE AU DATASET .

.............................................................................................................................................................. 54 FIGURE 28: LOG NORMAL HISTOGRAM DISTRIBUTION FOR PIT INFILL AU DATASET ....................... 54 FIGURE 29: LOG NORMAL HISTOGRAM DISTRIBUTION FOR IN-SITU AU DATASET ............................. 55 FIGURE 30: PLAN VIEW OF THE LOW GRADE STOCKPILE AND PIT INFILL AU DOMAIN MODELS ... 58 FIGURE 31: 3D VIEW OF THE LOW GRADE STOCKPILE AND PIT INFILL AU DOMAIN MODELS ......... 59 FIGURE 32: PLAN VIEW OF THE IN-SITU AU DOMAIN MODELS (COLOURS) ........................................... 60 FIGURE 33: 3D VIEW OF THE IN-SITU AU DOMAIN MODELS (COLOURS) ................................................. 61 FIGURE 34: CROSS SECTIONAL VIEW OF LOW GRADE STOCKPILE AU DOMAIN MODELS ................. 62 FIGURE 35: CROSS SECTIONAL VIEW OF PIT INFILL AU DOMAIN MODELS ........................................... 63 FIGURE 36: CROSS SECTIONAL VIEW OF IN-SITU AU DOMAIN MODELS ................................................. 64 FIGURE 37: HISTOGRAM OF SAMPLE INTERVALS – LOW GRADE STOCKPILE ...................................... 66 FIGURE38: HISTOGRAM OF SAMPLE INTERVALS – IN-SITU ...................................................................... 66 FIGURE 39: PLAN VIEW OF DOMBRALY LOW GRADE STOCKPILE BLOCK MODEL - AU GRADE

DISPLAY ............................................................................................................................................. 77 FIGURE 40: PLAN VIEW OF DOMBRALY PIT INFILL BLOCK MODEL - AU GRADE DISPLAY ................ 78 FIGURE 41: PLAN VIEW OF DOMBRALY IN-SITU BLOCK MODELS - AU GRADE DISPLAY.................... 79 FIGURE 42: PLAN VIEW OF DOMBRALY BLOCK MODELS - AU GRADE DISPLAY ................................... 80 FIGURE 43: 3D VIEW LOOKING NE OF DOMBRALY LOW GRADE STOCKPILE AND PIT INFILL

BLOCK MODELS - AU GRADE DISPLAY....................................................................................... 81 FIGURE 44: 3D VIEW LOOKING NE OF DOMBRALY IN SITU BLOCK MODELS - AU GRADE DISPLAY 82 FIGURE 45: 3D VIEW LOOKING NE OF DOMBRALY BLOCK MODELS - AU GRADE DISPLAY ............... 83 FIGURE 46: 3D VIEW LOOKING NW OF DOMBRALY LOW GRADE STOCKPILE AND PIT INFILL

BLOCK MODELS - AU GRADE DISPLAY....................................................................................... 84 FIGURE 47: 3D VIEW LOOKING NW OF DOMBRALY IN SITU BLOCK MODELS - AU GRADE DISPLAY ...

.............................................................................................................................................................. 85 FIGURE 48: 3 D VIEW LOOKING NW OF DOMBRALY BLOCK MODELS - AU GRADE DISPLAY ............... 86

LIST OF APPENDICIES

APPENDIX 1. MCS DOMBRALY SITE VISIT REPORT ................................................................................ 100 APPENDIX 2. DOMBRALY PROJECT DATABASE LISTING AND VALIDATION REPORTS ................. 133 APPENDIX 3. DOMBRALY DOMAIN TOP CUT STATS AND GRAPHS ..................................................... 139 APPENDIX 4. DOMBRALY VALIDATION CROSS SECTIONS.................................................................... 145 APPENDIX 5. DOMBRALY IDW RESOURCE NOV 2011 .............................................................................. 151


i

SUMMARY

This report presents the findings of recent NI 43-101 compliant computerised 3 dimensional resource

estimations for the Dombraly gold project, Akmola Oblast, Kazakhstan which were undertaken between

October 2011 and November 2011. The study was undertaken by ACA Howe International Limited (ACA Howe) and Micromine Consulting Services UK (MCS).

It is the opinion of ACA Howe and MCS that resources estimated as part of this study meet with CIM/JORC Inferred and Indicated category classifications based upon quality of input data, modelling

and estimation methodology, interpolation criteria based on sample density, search and interpolation

parameters, understanding and robustness of the geological model, drilling and sample density.

The resource estimation has an effective date of November 27th 2011 and represents a maiden NI 43-101

compliant resource estimation for the project.

Alhambra has six exploration projects within the polygons of one large exploration licence area

extending in aggregate to 9262.5 square kilometres, located in northern Kazakhstan. The area is served

by a network of paved and dirt roads and a railway. There are airports at Aksu near Stepnogorsk and a medium sized airport at Kokshetau.

Saga Creek Gold Company LLP (Saga Creek) is the local operating company in Kazakhstan, which is owned 100% by Alhambra.

The license area is in the Charsk Gold Belt between the Vasilkovskoe gold deposit on the northwest and

the Asku gold deposits on the southeast. All operations are conducted through the 100% Alhambra owned subsidiary Saga Creek.

Dombraly forms one of six exploration projects:

Dombraly,

Shirotnaia,

Zhanatobe,

Kerbay,

North Balusty,

Vasilkovskoe East.

The area forms part of the Caledonian-age Kokchetav-North Tienshan basin and fold system.

Increasingly felsic intrusive magmatism is related to 3 orogenic cycles ranging in age from the Pre-Cambrian to the Hercynian. Mineralisation is mainly hosted by Middle to Upper Ordovician volcano-

sedimentary rocks of mainly mafic and intermediate composition with interbedded lavas, tuffs and

terrigenous clastic rocks.

The mineralisation is described as volcano-sedimentary hosted, orogenic type. Zhanatobe, Kerbay and

Vasilkovskoe East are also influenced by intrusive magmatism.

Alhambra commissioned ACA Howe and MCS to compile relevant data and complete resource

estimations and prepare a Technical Report on the Dombraly gold project, Akmola Oblast, Kazakhstan.

The resource estimations and this Technical Report are prepared in accordance with CIM best practice methodology and NI 43-101 for release to the TSX and other markets, and issue on SEDAR. Resource

data and resource estimation sections of the report were prepared by Mr. J N Hogg MSc, MAIG, Senior

Geologist of MCS and Ms. M Sostre MSc, AUSIMM, Resource Geologist of MCS. Other sections were

prepared by Mr. J G Langlands, BSc, FGS, FIMMM, C Eng and Mr. J Emberton, BSc, FIMMM, C Eng or by J G Langlands and J N Hogg, together.


ii

The scope of work, restricted to the Dombraly resource estimation study, included:

Data preparation, compilation, review and validation,

Site visits,

Geological and grade domain interpretation and modelling,

Block model resource estimation for in-situ, low grade stockpile and pit infill material,

Preparation of a Technical Report in English, using Canadian National Instrument (NI

43-101) reporting standards.

The data on which this report is based are those which were available to ACA Howe and MCS up to November 2011. Exploration projects are ongoing and, in due course, subsequently available data may

enable further updated CIM compliant mineral resource estimates to be made for Dombraly.

Mineralisation at Dombraly comprises low grade stockpile, pit infill and In situ deposits. Pit infill material which forms backfill to the open pit is derived from the low grade stockpile.

The mineralisation at Dombraly is typically quartz veins, quartz veinlets and disseminated zones, containing native gold and pyrite with sporadic galena, arsenopyrite and specularite (haematite). All

mineralised zones and quartz veins strike north-northwest and dip to the northeast at 30 to 65 degrees.

Quartz veins are up to 350 metres (m) long and 1.5 m wide while veinlet and disseminated

mineralisation zones reach 600 m in length and 90 m in width, remaining open at both ends. These zones are explored by core drilling to depths of 360 m and are open down dip. The oxidation boundary

is variable in depth and in some tectonic zones extends to 380 m from surface.

Between 2002 and 2006, Saga Creek explored the Dombraly area including the mineralisation below the

open pit and sampling of the backfill and low grade stockpile. During this period, 8 trenches totalling

1741 m, 613 rotary air-blast (RAB) drill holes amounting to 4152 m and 10 core holes for a total of 2394 m were completed. Sampling included 354 samples from the trenches, 2076 from RAB drill holes

and 1958 from core holes. The results of the 2010 exploration programme at Dombraly, including 13

DDH and 37 reverse circulation (RC) resource holes in the low grade stockpile and around the open pit,

are included in the present report.

Ongoing exploration programmes are designed to test and further quantify the remaining oxide

resources below and adjacent to the abandoned open pit and to test the various mine waste materials for suitability for gold extraction by cyanide leaching. Additionally, the work is to test the lateral, oxide and

deeper, sulphidic extensions of the deposit and to identify and explore other gold bearing occurrences in

the area.

For the 2011 resource modelling, raw data used in interpretation and modelling consists of data from

recent and historical diamond drilling and RC drilling, low grade stockpile trench and RAB sampling

exploration work undertaken by Alhambra and previous explorers.

Raw data used as input to grade estimation consists of recent 2010 diamond, RC drill data and 2005 low

grade stockpile trench data.

The boundaries to mineralised zone domains digitised in cross section were interpreted using best practice industry standard techniques, snapping to drillhole assay intervals, and utilising lithology where

applicable to improve accuracy of location of mineralised zone in 3 dimensions, and to reduce the

inclusion of waste within the mineralised wireframes.

Grade domain modelling was completed for Au, with minimum width of approximately 1m. Where appropriate, geology was used in combination with grade values to assist zone interpretation.

The following grade domain models were generated across the 3 recognised deposit zones:


iii

Low Grade Stockpile: 2 domains interpreted using an Au cut off value of 0.13g/t Au.

Pit Infill: 1 domain interpreted using an Au cut off value of 0.1g/t Au.

In-Situ: 11 domains were interpreted using an Au cut off value of 0.2g/t Au.

Inverse distance weighting (IDW³) method of interpolation was used, which is a non-geostatistical

(classical) method of grade interpolation. In this method, each input sample is weighted according to some power of the inverse distance from the block to be estimated. Interpolation weights are only

applied to samples found within the search neighbourhood. There are no strict rules for choosing a

power; for gold a value of two or three is often used, with three most common. For iron a power of two

may be appropriate. The lower the power, the more the grades are smoothed, to the point where using a low power will produce a result which deviates only slightly from the global mean of the data. On the

other hand, higher powers will produce a result that approaches a nearest-neighbour interpolation, with

the sample nearest the block contributing almost all of the weight.

A power of three was chosen for this interpolation given the deposit, grades and commodity type.

The Dombraly block model resource reporting has been based on criteria that were established

according to best practice geological modelling techniques, current understanding of the geological model, historical interpretations and discussion with Alhambra personnel as described in previous

sections.

Potentially economic mineral resources are being reported by use of an economic cut-off grade

dependent upon the cost of mining and processing the mineralization and the selling price of the final product.

The economic cut-off grade for Dombraly established using grade and block revenue factors.

Due to the early stage status of the development of the Dombraly deposits, a number of assumptions have been made with regard to inputs to the calculation of the economic cut-off grade for reporting.

Inputs for the calculation of block revenue for the Dombraly deposit are US$ value per ppm, and

assumed metal % values in concentrate (product).

Inputs for oxide material are based upon actual mining cost data from Alhambra’s nearby Uzboy open

pit operation, and estimated costs for transitional and primary material taken from recent PEA studies

undertaken on the nearby Uzboy deposit.

Key input data for cut off calculation include:

Gold price - US$1,394/oz

Mining Method – open pit

Oxide processing method – heap leach

Transitional and primary processing method – gravity CIL

Recovery – Oxide 70%; Transitional/Primary 85%

Oxide mining cost – US$1/t (low grade stockpile and pit infill)/US$1.7/t (in-situ)

Transitional and Primary mining costs – US$1.95/t

Processing costs – US$3.85/t (oxide), US$6.47/t (transitional and primary)

Using the Au block grade, the above Au metal price and recovery, MCS estimated the revenue per

mined block.

For a mineralised block to be considered economic it must generate higher revenue than it costs to mine.

For a block to be considered economic it must therefore generate greater than US$4.85/t of revenue for


iv

the low grade stockpile and pit infill material, and US$5.55/t and US$8.42/t, for in-situ oxide and

transitional/primary material respectively.

MCS used Micromine software to filter those blocks in the resource model with value greater than the

calculated cost to mine values for economic cut-off grade determination and resource reporting.

Cut off grades used for reporting are 0.1g/t Au for thelow grade stockpile, 0.2g.t Au for the pit infill zone and 0.1g/t for in-situ oxide material, and 0.2g/t Au for in-situ transitional and primary material

types.

It is MCS’ opinion that the assumptions made for input to economic cut-off grade determination and reporting of potentially economic resources are reasonable given the current understanding of the

geology, mineralisation, anticipated mining and processing methods and comparison with similar type

operations.

At Dombraly, a total of 9.3 million tonnes of Inferred resources, grading at 1.01 g/t Au for 301,000

ounces Au have been identified. An additional 0.6 million tonnes of Indicated resources grading

at 1.22 g/t Au have been identified for 22,000 ounces.

A summary of in situ classified inferred resources as of November 2011 for the Dombraly Deposits are presented in the tables below.

Dombraly Low Grade Stockpile Total Resource by Category and Material Type

CUTOFF¹ MATERIAL CLASS²

Density Volume Tonnes Au³ Au Au

t/m3 x 1000 m3 x 1000 t g/t g Oz

Oxide

Indicated 1.67 284 473 1.26 597,000 19,000

0.10g/t Inferred 1.67 578 963 1.07 1,033,000 33,000

Dombraly Pit Infill Total Resource by Category and Material Type



t/m3 x 1000 m3 x 1000 t g/t g Oz

Oxide

Indicated 1.73 50 86 0.97 83,000 3,000

0.20g/t Inferred 1.73 525 908 0.82 747,000 24,000

Dombraly In Situ Total Resource by Category and Material Type



t/m3 x 1000 m3 x 1000 t g/t g Oz

0.1g/t Oxide Inferred 2.63 1,043 2,700 0.99 2,700,000 87,000

0.2g/t Transitional Inferred 2.61 249 646 1.16 750,000 24,000

0.2g/t Primary Inferred 2.71 1,364 3,671 1.12 4,099,000 132,000

0.1g/t Total Inferred 2.64 2,807 7,446 1.02 7,601,000 244,000 ¹ Cut off value used here represents economic cut off determined from block revenue factor calculation methodology and input gold price of

US$1,394/Oz.

²Class represents resource category under CIM and JORC reporting guidelines.

³ Top cuts of 10g/t Au and 6g/t Au have been applied to in situ domains A and F gold assay data respectively. Top cuts of 10g/t Au and 5g/t Au

applied to low grade stockpile (lower), and pit infill domains respectively.

Classification of resources for the low grade stockpile and pit infill are restricted to Indicated and Inferred.

Classification of in situ resources is restricted to inferred category due to the following factors which introduce uncertainty:


v

o Limited number of valid drill holes and drill samples clustered in small areas. o The number of valid drillholes are widely spaced along domain extents

o A very low number of valid samples per mineralised domain

o A low number or no bulk density data for a number of domains and sub-domains

o Lack of QAQC data, and quality control issues

On working through the estimation process, it became clear that although the in situ deposit models are

coherent and robust based upon an interpretation of combined historical and recent (valid) drilling, the domains require significant additional drill testing to increase valid input sample data numbers and

sample density for both grade and bulk density determination, and improved resource block

classification.

Quality control sample data analysis and interpretation raised a number of issues with respect to assay

precision and repeatability. This could be due to nugget effect or sampling error, and will require

follow up investigation studies.

Due to these reasons the restriction and selection of resource classification currently applicable to the

deposit areas are deemed appropriate, particularly for in-situ domains.

Mineral processing and recovery studies were reported in 2005 and 2006 by Kazmechanobr.

Comprehensive leaching test work on composite sample DLT-1 of oxide, reported a head grade of 1.96

g/t Au. This work showed that the agglomerated, minus 50 mm oxide mineralisation can be processed

using heap leaching with a predicted 74.6% recovery of gold during commercial scale production. ACA

Howe noted that very high cement consumption was required. The material of the sample, while typical, may not be spatially and statistically representative of the whole of the oxide mineralisation.

Column leach test work was carried out on sample DLT-1 -add from the oxide dump terrace, which is reported to grade 1.52 g/t Au. The combined gold recovery on resin and in the stripping solutions during

the column leach test, was relatively low at 53.33%. This indicates to ACA Howe that commercial scale

recovery will be lower still. Again, high cement consumption was required to maximise gold recovery.

Enlarged bottle roll tests showed that gold recovery from ore crushed to 80% minus 0.074 mm was around 89%. Good gold recovery from sample DLT-1 -add depends for the most part on fine grain size

following crushing. It is clear that studies of crushing and agglomeration and leaching will be required

to optimise gold recovery from this type of material. ACA Howe suggests that this ore type may be more suitable for processing by conventional CIL technology (Carbon in Leach). This dump terrace

material, while described as characteristic, is not spatially and statistically representative of the whole of

the oxide mineralisation and it has different mineralogical and relatively poorer metallurgical characteristics compared with sample DLT-1 material which may be from undisturbed oxide material.

Nine small core samples from oxide, transitional and sulphide mineralisation zones were submitted for

petrographic and chemical analysis and bottle roll cyanidation tests. The sampling intervals range from 25 to 327 m downhole and the weights of samples range from 0.29 to 1.92 kilograms (kg). Gold grades

in the samples range from 1.54 to 23.2 g/t Au. Standard bottle roll tests were conducted, including direct

cyanidation and adsorption leaching with resin. Cyanide leaching showed high gold recovery rates from oxide and transitional samples of 94% to 98%. Gold recovery from sulphide samples by direct

cyanidation was considerably lower at 29% to 83%. These sulphide ores have a preg-robbing effect, i.e.

the ore contains active natural adsorbent such as organic carbon which adsorbs gold leached into solution. It is the opinion of ACA Howe that these nine core samples are not spatially and statistically

representative of the whole of the oxide, transitional and sulphide mineralisation and the results of this

study, alone, cannot be used to map the boundaries between these types of mineralisation.

Metallurgical test work on sulphide mineralisation sample DLT-2 includes gravity, froth flotation,

cyanidation and adsorption leaching on an 85.6 kg composite sample of drill core from downhole depths

of 127 to 348 m. The average gold content is reported as 7.19 g/t Au. The sample rock types are


vi

described as carbonaceous sedimentary and tuffaceous sedimentary rocks with dispersed disseminated

pyrite. Detrimental impurities include arsenic and carbon. Gravity separation tests recovered 53.45% of the head grade gold. Froth flotation tests recovered 86.13% of the gold. Direct cyanidation leach testing

recovered only 17.65% of the gold into cyanide solution due to the presence of carbonaceous shale in

the ore, which acts as a natural adsorbent of gold. Cyanidation leach tests with resin adsorbent recovered

79.23% of the gold to the resin and 64.81% of the gold was recovered from the loaded resin in gravity tails. It is the opinion of ACA Howe that the sulphidic gold ore represented by sample DLT-2 could be

successfully processed by one of two schemes using 1. Gravity separation and froth flotation or 2.

Gravity separation and conventional carbon- or resin-in-leach technology and that further work is required to determine which route is economically most beneficial. It is the opinion of ACA Howe that

the Dombraly sulphide ore material represented by sample DLT-2, while described as characteristic, is

probably not spatially and statistically representative of the whole of the sulphide mineralisation. However, a wide range of metallurgical recovery characteristics has been illustrated, which is a useful

guide for future studies.

The current resource estimations build upon those performed in 2007, where Saga Creek estimated resources of oxide + backfill + low grade stockpile in Category C2 as 4.25 million tonnes at 1.52 g/t Au;

sulphide to 300 m depth in Category C2 as 0.58 million tonnes at 5.42 g/t Au; sulphide from 300 m to

500 m depth in Category P1 as 0.51 million tonnes at 6.02 g/t Au.

The current models and estimations for Dombraly near surface zones are by no means exhaustive.

Based on available information, ACA Howe and MCS believe that the exploration and resource development of Dombraly is progressing well and that there is scope to develop a potentially

economically viable gold resource.

Strike and dip directions remain open for a large number of lodes and existing stope and mine level wireframes indicate the presence of additional structures some previously mined and developed others

yet to be adequately explored and exploited.

Results of the block model estimations for the deposit zones modelled using limited data collected thus

far for the areas are positive, and offer excellent potential for development of significant resources

within the immediate Dombraly mine area. The interpretation of higher grading zones will no doubt

further aid prioritisation of drill targeting and future resource/reserve development.

Open pit production took place in 1985 to 1988 when the price of gold was very low, at US$320 to

US$460 per troy ounce, using a gold cut-off grade of 2.5 g/t Au. With improved heap leaching techniques and current gold prices, the economics of potential gold production at Dombraly are again

attractive even at much lower grade, well below the historical cut-off grade.

A number issues and sensitivities have been highlighted as part of this study and are outlined below and

expanded upon in the report. These issues ultimately impact on the robustness and confidence of the

geological and resource model and should be considered for improved assessment, estimation of higher

classification of resources and mine planning.

Data Collection

Analysis

Domain modelling

Metallurgical testwork


1

1 INTRODUCTION

This report is prepared for Alhambra Resources Limited, of Calgary, Alberta, Canada (Alhambra, the issuer, the Company) by ACA Howe International Limited (ACA Howe) and Micromine Consulting

Services (MCS).

Alhambra is currently undertaking exploration and development of gold deposits and prospects on their exploration and mining licences in north-central Kazakhstan, known us the Uzboy Project. Dombraly is

one of a group of exploration projects distributed around the active Uzboy gold mine, comprising:

Dombraly,

Shirotnaia,

Zhanatobe,

Kerbay,

North Balusty,

Vasilkovskoe East.

Alhambra commissioned ACA Howe and MCS to compile relevant data and complete resource

estimations and prepare a Technical Report on the Dombraly gold project, Akmola Oblast, Kazakhstan.

The resource estimations and this Technical Report are prepared in accordance with CIM best practice methodology and NI 43-101 for release to the TSX and other markets, and issue on SEDAR. Resource

data and resource estimation sections of the report were prepared by Mr. J N Hogg MSc, MAIG, Senior

Geologist of MCS and Ms. M Sostre MSc, AUSIMM, Resource Geologist of MCS. Other sections were

prepared by Mr. J G Langlands, BSc, FGS, FIMMM, C Eng and Mr. J Emberton, BSc, FIMMM, C Eng or by J G Langlands and J N Hogg, together.

Saga Creek Gold Company LLP (Saga Creek) is the local operating company in Kazakhstan, which is owned 100% by Alhambra.

Alhambra’s Qualified Person (QP) is Elmer B. Stewart, M. Sc. P. Geol., who is a technical consultant

for Alhambra. He is the QP responsible for monitoring the supervision and quality control of the exploration programmes.

Site visits and meetings were undertaken by Mr. Evgenij Zhuravlyov, Senior Geologist, Micromine

Consulting Services (Kazakhstan), between the dates 12th and 14

th August 2011, in the company of Mr.

Evgeny Plyushchev, Alhambra (Saga Creek) Consulting Geologist and other Alhambra site personnel.

The purpose of the site visit and meetings was to review deposit geology, receive and review the

available data base, review data collection methodologies and determine modelling criteria for the purpose of completing an NI 43-101 compliant resource estimation study. A copy of the site visit report

is presented in Appendix 1.

A personal inspection of the Dombraly property was made during the visit to the Uzboy Project by ACA

Howe Senior Associate Mining Engineer, Bruce Brady, and Senior Associate Metallurgist, Gary Patrick, on 25 November 2010. It was concluded that, for present purposes, Dombraly is an advanced

exploration project. A substantial amount of historical and more recent exploratory work has been

carried out by previous and current owners and exploration activity is ongoing.

The items for resource estimation comprise: 1. in-situ mineralisation, 2. mineralised low grade stockpile

and 3. mineralised open pit back-fill material zones.

This report is based on the findings of the MCS and ACA Howe site visits, desk study, input data

review, data validation, deposit modelling, block model grade interpolation and resource estimation.

The report follows the format for National Instrument 43-101 reports, and provides a template for future NI 43-101 compliant reports.


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The mineral resources calculated as part of this study are considered to be reportable resource estimates

compliant with the NI 43-101 reporting requirements and CIM/JORC codes suitable to be filed on SEDAR.

The sources of information used in the preparation of this report are listed in the section: References and

Sources of Information and are cited in the text where appropriate.

1.1 TERMS OF REFERENCE

The scope of work, restricted to the Dombraly resource estimation study, includes:

Data preparation, compilation, review and validation,

Site visits,

Geological and grade domain interpretation and modeling,

Block model resource estimation for in-situ, low grade stockpile and pit infill material,

Preparation of a Technical Report in English, using Canadian National Instrument (NI

43-101) reporting standards.

The data on which this report is based are those which were available to ACA Howe and MCS up to

November 2011. Exploration projects are ongoing and, in due course, subsequently available data may enable further updated CIM compliant mineral resource estimates to be made for Dombraly.

Mineralisation at Dombraly comprises low grade stockpile, pit infill and In situ deposits. Pit infill material which forms backfill to the open pit is derived from the low grade stockpile.

1.2 ACA HOWE INTERNATIONAL LIMITED

ACA Howe International Limited is an internationally recognised, independent geology and mining

consultancy with offices in Canada where it was established in 1961 and in the United Kingdom where

it has operated since 1978.

ACA Howe, its directors, employees and associates do not hold:

any rights to subscribe for shares in Alhambra Resources Ltd, either now or in the future;

any vested interests in any concessions held by Alhambra Resources Ltd;

any rights to subscribe to any interests in any of the concessions held by Alhambra Resources Ltd,

either now or in the future;

any vested interests in either any concessions held by Alhambra Resources Ltd or any adjacent

concessions;

any right to subscribe to any interests or concessions adjacent to those held by Alhambra

Resources Ltd, either now or in the future.

ACA Howe's only financial interest is the right to charge professional fees at normal commercial rates,

plus normal overhead costs, for work carried out in connection with the investigations reported here.

Payment of professional fees is not dependent either on project success or project financing.

1.3 MICROMINE CONSULTING SERVICES

Micromine Consulting Services (MCS) is an internationally recognised, independent geology and

mining consultancy with offices in the United Kingdom, China, Australia, Mongolia, and Kazakhstan.

Micromine Consulting Services have been providing services to the exploration and mining industry since 1986.


3

MCS, its directors, employees and associates do not hold:

any rights to subscribe for shares in Alhambra Resources Ltd either now or in the future

any vested interests in any concessions held by Alhambra Resources Ltd

any rights to subscribe to any interests in any of the concessions held by Alhambra

Resources Ltd, either now or in the future

any vested interests in either any concessions held by Alhambra Resources Ltd or any

adjacent concessions

any right to subscribe to any interests or concessions adjacent to those held by

Alhambra Resources Ltd, either now or in the future

Micromine’s only financial interest is the right to charge professional fees at normal commercial rates,

plus normal overhead costs, for work carried out in connection with the investigations reported here. Payment of professional fees is not dependent either on project success or project financing.

1.4 UNITS

All units of measurement used in this report are metric unless otherwise stated. Tonnages are reported

as metric tonnes (t), precious metal values (gold and silver) in grams per tonne (g/t) or parts per million (ppm) and base metal values (tin, copper, lead and zinc) are reported in weight percent (%) or parts per

million (ppm). Other references to geochemical analysis are in parts per million (ppm) or parts per

billion (ppb) as reported by the originating laboratories.

Data is captured and located using Gauss-Kruger grid coordinates (GK) based on the Pulkovo 1942

datum. The Property is located in GK Zone 12.

2 RELIANCE ON OTHER EXPERTS

ACA Howe and MCS have relied upon the accuracy of all information provided by Alhambra and other

sources cited in this report and have no reason to believe that the information is not accurate.

ACA Howe and MCS have relied upon the translations into English from Russian, provided by Alhambra and Saga Creek personnel, of all the documentary data cited in the report and listed in the

References and Other Sources of Information where such sources are labeled with the bracketed words

(Contains information originally reported in Russian).

In addition to their own observations during discussions and from the literature, the writers have relied

on information provided by Mr. Evgeny Plyushchev, Alhambra Consulting Geologist, Mr. Richard Gorton, Alhambra Consulting Geologist, and other site personnel. Part of the information may be

presented to the writers only in verbal communication without any written evidence. The writers

highlight “verbal” information in this report. The writers consider that the information provided by these

individuals is reliable and relevant.

3 PROPERTY DESCRIPTION AND LOCATION

3.1 LOCATION

The Dombraly exploration project is located in Figure 1.


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Subsequent to a licence area reduction which was implemented in 2001 (see below), the Dombraly,

Shirotnaia, Kerbay and North Balusty Projects are located within the boundaries of the East Area of the Sary-Arka Licence with serial number MG 1029-D, held by Saga Creek.. Zhanatobe is located within

the separate Mamayskoe Area of the same licence. The Vasilkovskoe East project is located in the West

Area of the same licence.

The East Area extends to approximately 3024.5 square kilometers (km²). The Mamayskoe Area extends

to 350 km². The Taskuduk Area extends to 157 km². The West Area extends to 5731 km². The total area

of the Sary-Arka Licence is 9262.5 km².

The corners of these licence areas are shown in Table 1 below and the license polygons are shown in

Figure 1. All Government property descriptions utilise latitudes and longitudes and both the Sary-Arka and Uzboy Licences are described in this way. No property boundaries are surveyed on the ground.

TABLE 1. SARY-ARKA LICENCE COORDINATES

The Dombraly exploration project of this report lies within East Area.

Entire Sary-Arka Licence: MG 1029-D is 9262.50 km²

EAST AREA (3024.50 km2)

1 53o26'00'' 71o17'00''

2 53o26'00'' 71o40'30''

3 53o00'00'' 72o20'00''

4 52o29'00'' 72o18'00''

5 52o28'00'' 71o48'00''

6 52o40'00'' 72o05'00''

7 53o10'00'' 71o46'00''

8 53o10'00'' 71o17'00''

WEST AREA (5731.00 km2)

1 53o18'00'' 69 o 30'00''

2 53 o 37'00'' 69 o 49'00''

3 53 o 37'00'' 70 o 20'00''

4 53o26'00'' 70o20'00''

5 53o26'00'' 70o41'00''

6 53o10'00'' 70o41'00''

7 53o10'00'' 70o54'00''

8 52o47'00'' 70o54'00''

9 52o47'00'' 69o58'00''

10 53o06'00'' 69o55'00''

MAMAYSKOE AREA (350.00 km2)

1 52o40'00'' 71o15'00''

2 52o40'00'' 71o33'00''

3 52o30'00'' 71o33'00''

4 52o30'00'' 71o20'00''

5 52o35'00'' 71

o15'00''

TASKUDUK AREA (157.00 km2)

1 53o01'05'' 71o10'00''

2 53o04'00'' 71o17'00''

3 52o57'00'' 71o26'00''

4 52o54'00'' 71o19'00''


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There are two exclusion areas in East Area: ‘Baylusty Deposit’ of 1.32 km² and ‘North Baylusty’ of 2

km². There are seven exclusion areas in West Area: ‘Borovoye Health Resort’ of 479 km², ‘Alexandrovskoe Deposit’ of 2.3 km², ‘Area No. 1’ of 0.04 km², ‘Area No. 2’ of 0.13 km², ‘Area No. 3’

of 0.66 km², Madeniet Nos. 1 and 2 of 68.8 km² for placer gold.

The Taskuduk Area of Licence MG 1029-D does not contain any of the exploration prospects identified here and, therefore, is not mentioned further in this report.

The property boundaries were located by plotting the latitudes and longitudes of the corner points of

Licence No. MG 1029-D, East Area and Mamayskoe Area on Google Earth satellite imagery mosaic maps. This plot produced boundaries which enclose, or bear the expected spatial relationship with,

recognisable ground features visible on the satellite imagery maps, such as the following. Dombraly is

identified in bold text in the following list.

Alhambra’s Uzboy open pit, heap leach gold mine (Lat. 53o21’33” N, Long. 70

o59’30” E);

The Vasilkovskoe gold mine which is located about 20 kilometres (kms) west of Alhambra’s

West Area licence polygon (Lat. 53°25'44.57"N, Long. 69°14'3.66"E);

The Aksu uranium open pit and waste heap which are located just inside the southern boundary

of Licence No. MG 1029-D, East Area (Lat. 52o28’44” N, Long. 71

o59’41” E);

Dombraly, open pit, dumps, exploration trenches and access roads (Lat. 52o54’55” N,

Long. 72o04’33” E);

Shirotnaia, trenches and area of drill holes of 2007 and 2010 (Lat. 52°30'40.59"N, Long.

71°59'56.27"E);

Zhanatobe, old mine located in the northeast corner of the Mamayskoe Area of Licence MG

1029-D (Lat. 52°38'57.76"N, Long. 71°31'2.50"E) and access tracks and old earthworks, just

east of centre of prospect at RAB hole JT-0713 (Lat. 52°37'26.46"N, Long. 71°28'42.79"E);

Kerbay, few old tracks and lines on 2002 satellite imagery, centre of prospect is trench KT-78/1

(Lat. 52°37'21.45"N, Long. 72° 7'32.91"E);

North Balusty, several trenches of various orientations, centre of prospect is at centre of the

KGK drill line (Lat. 52°58'30.25"N, Long. 72° 9'6.72"E).

Vasilkovskoe East, several areas of apparently backfilled trench sites and apparently

rehabilitated drill sites, e.g. Akshasor (Lat. 53°16'39.89"N. Long. 69°46'35.80"E).

3.2 LICENCE AND TENURE

The following licence information which is believed to be up to date, is extracted from a report by ACA Howe (2009). The legal status of the licence has not been verified by ACA Howe, who are not qualified

to make any judgement on legal issues.

Alhambra’s 100% ownership of Saga Creek came about as a result of various transactions reported by ACA Howe (2009), which are summarised below.

In 1996 Cameco Gold Inc. (Cameco) formed a Joint Venture with Goodwin Golems LLP (Goodwin), a local Kazakhstan company based in Almaty. The Joint Venture company was named the Saga Creek

Gold Company LLP (Saga Creek) and was registered in the form of a joint venture with limited liability

to carry-out mineral exploration and exploitation in Kazakhstan. The Joint Venture was managed and operated by Cameco with Cameco’s interest being 80% while Goodwin’s interest was 20%. Cameco

funded 100% of all expenditures.

Two exploration and exploitation licences in northern Kazakhstan were initially granted to Cameco; the Sary-Arka Licence (MG 1029-D issued April 8, 1996), a large area of land extending from a point

immediately east of the city of Kokshetau at the northwest end of the licence area, extending in a

southeasterly direction to the mining town of Bestube in the southeast. A second, smaller, licence, the Uzboy Licence (MG 719D – issued May 23, 1996), was located internally within the Sary-Arka licence

and thus was an area excluded from it.

FIGURE 1: LOCATION MAP SHOWING DOMBRALY AREA, PROSPECTS AND

SURROUNDING GOLD PROJETS

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The licences were consolidated and re-issued to Saga Creek on 10 February 1997 and are valid for 25 years, with the initial 5 years specifically devoted towards exploration and the remaining 20 years for

mining purposes.

Saga Creek, under Cameco management, carried out exploration efforts on both licences during 1996 and 1997 but concentrated their efforts on the Uzboy prospect itself. In April 2000, Cameco pulled out

of the Joint Venture and transferred its share and obligations to Marsa Aktiengesellschaft (Marsa) of

Mauren, Liechtenstein. After the transfer, Saga Creek was owned by Marsa (80%) and Goodwin (20%).

The ownership of Goodwin was sold in March 2002 to Teragol Investments Limited (Teragol)

incorporated in Cyprus. Saga Creek carried out exploration activity on the Uzboy licence until the end of 2001. Marsa funded all expenditures since the transfer of ownership through to 2001. Cameco no

longer retains any interest in either the joint venture or the licences.

Pursuant to a Partnership Unit Purchase and Exchange Agreement dated March 21, 2002 (Agreement), Alhambra agreed to purchase all of the partnership units of Saga Creek and Goodwin Golems, and as a

result, obtained the licences to the Uzboy Property in exchange for 4,000,000 common shares of

Alhambra. Alhambra also agreed to grant to Marsa and Teragol, a net smelter return on production from the Uzboy Property, based on up to 3% of gross revenue in the event that the weighted average price of

gold is equal to or greater than US$350 per ounce and less revenue for lower gold prices.

The TSX Venture Exchange approved the Transaction in October 2003.

The Republic of Kazakhstan Contract on Foreign Investments had been put in force in 1997 and dictated

that 50% of the initial licence was to be relinquished over time (5 years from the date of granting the licence). By September 2001, an area in excess of 50% of the total initial licence area had been

relinquished in accordance with the licence conditions. Two mining allotments were granted in

December 2001 to Saga Creek on areas of recognised gold mineralisation at the Uzboy prospect itself.

4 REGIONAL GEOLOGY

The Dombraly, Shirotnaia, Zhanatobe, Kerbay, North Balusty and Vasilkovskoe East project areas are

all located within the northeast part of the Caledonian-age Kokchetav-North Tienshan basin and fold

system. Four major elongate structural zones within the system form the framework for the region.

Archaean to middle or late Palaeozoic sedimentary, magmatic and metamorphic rocks form the

basement of the region. These are intruded by Cambrian to Devonian age ultrabasics, gabbros, diorites, granodiorites and granites.

Major deep-seated faults cross the region. Most were generated during the Proterozoic and tend to separate the different crustal blocks.

The Pre-Cambrian formations are associated with the cores of the oldest anticlinoria. They may be

divided into two large complexes: Early and Late separated by a regional, folded unconformity. The Early Pre-Cambrian includes complicated, dislocated metamorphic layers ranging from Archaean to

Middle Proterozoic age. This complex of rocks has two clear sub-divisions that are dramatically

different in degree of metamorphism. The lower part includes rocks metamorphosed to amphibolite facies. Metamorphism of the upper part of the Early Pre-Cambrian is greenschist facies.

The Late Pre-Cambrian is represented by Upper Proterozoic, slightly metamorphosed, volcano-

sedimentary formations. These underlie the faunally characterised Lower Palaeozoic.

Cambrian rocks consist of conglomerates, conglomerate-breccias, coarse sandstones and argillites with

isolated pods of limestones and calcareous sandstones.


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The Ordovician period is represented by thick volcanogenic formations, mainly of mafic and intermediate composition. The formations are interbedded with layers of terrigenous clastic material and

tuffs and lavas.

The Silurian - Devonian sequence is composed of a basal conglomerate, up to 35 – 90 m in thickness, interbedded with sandstones, tuffaceous sandstones and argillites with lenses of limestones. Overlying

rocks are mainly rhyolites with subordinate andesites and tuffs.

Carboniferous rocks are mainly limestones often lying unconformably on the older rocks. They are

overlain by Namurian sandstones and brown coals.

Large areas are covered with continental sandy-clayey lateritic deposits of the Upper Cretaceous,

Palaeogene, Neogene and Quaternary age. Lacustrine and alluvial deposits are widely developed.

Intrusive magmatism is related to three orogenic cycles. The oldest is represented by gneissose rocks within the Pre-Cambrian series. The early Caledonian intrusives are mainly basic in composition:

gabbros, norites, also pyroxenites, peridotites and dunites. Later intrusions are of granite and

granodiorite. Hercynian intrusives are mainly granites and syenites which form rounded massifs with areas of more than 100 square kilometres.

Four sets of faults are distinguished, with north-northwesterly, northwesterly, northeasterly and easterly trends. Some faults were formed during the Pre-Cambrian, others are connected with formation of the

Caledonian synclinoria, or have been developed during the Hercynian orogenic cycle. All major faults

feature a long history of development during which they have changed direction of displacement many

times, both horizontally and vertically.

Dombraly, Shirotnaia, Zhanatobe, Kerbay, North Balusty and Vasilkovskoe East lie within the Charsk

Gold Belt (ACA Howe, 2009, Figure 1) which extends through the northeastern part of Kazakhstan from China to Russia, trending east-southeastwards over a width of 300 kms and a distance of 1800 kms

within Kazakhstan (Figure 1). The Charsk Gold Belt contains a number of substantial gold deposits and

mines including those near the Uzboy Project area, which are mentioned in the Adjacent Properties

section below.

5 ADJACENT PROPERTIES

The qualified persons who prepared this report have been unable to verify the available information on

the adjacent properties and this information is not necessarily indicative of the mineralisation on the Alhambra property.

There are two gold mines located adjacent to or close to the greater Uzboy Project area of licences, also in the Charsk Gold Belt: Aksu gold mine and Vasilkovskoe gold mine. Aksu is adjacent to Shirotnaia,

located 2.3 km south of the southern boundary of the East Area of Alhambra’s Licence MG 1029-D.

Vasilkovskoe is located 21.5 kms to the northwest of the westernmost corner of the West Area of

Alhambra’s Licence MG 1029-D and 116 kms west of the Uzboy mine (Figure 1).

Aksu

Three gold mines in northern Kazakhstan are owned by KazakhGold in which Polyus had a controlling interest, now sold back to the Assaubayev family, namely - Aksu, Zholymbet and Bestobe (Stiskin et

al., December 2010). Aksu is south of and adjacent to Alhambra’s Shirotnaia exploration project. Each

of these mines is reported to process about 350,000 tonnes of ore per year

(http://www.mbendi.com/indy/ming/gold/as/kz/p0005.htm#5). Correcting obvious errors reporting thousands instead of millions on the website http://www.kazakhgold.com/operations/resources, as of

June 2005, Kazakhstan Standard reserve and resource totals for the three mines of B+C1 reserves plus

http://www.kazakhgold.com/operations/resources


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C2+P1 resources of 46.6 million ounces of gold, are reported. The category B+C1 reserves, amounting

to 13.2 million ounces of gold, were split between the mines as follows: Aksu 44%

Zholymbet 33%

Bestobe 23%

A major exploration programme and mine expansion plan were to be implemented. On the website http://www.kazakhgold.com/documents/327/Swift%20-%20Investor%20presentation.pdf, a map

indicates that these mines were expected to have resources of 8 - 10 million ounces of gold by the end of

2011.

Vasilkovskoe

The Vasilkovskoe gold mine, just west of Alhambra’s West Area licence polygon, is owned by Kazzinc of which 51% is owned by Glencore International AG of Switzerland (Glencore) (Stiskin et al.,

December 2010). Vasilkovskoe has a reported resource of 7.7 or 12.6 million ounces of gold (dates

unknown) (Alhambra, January 2011a, Figure 1) and is said to be the largest gold mine in Kazakhstan.

The gold is reported to be associated with tin and tungsten and the mineralisation has a low total sulphide content. Production of 955 kg of gold is recorded in 2003. Bekzatov (2004) reported the field

has been developed since 1979, “proven reserves” of 370 tonnes (11.9 million troy ounces) of gold at an

average grade of 2.8 g/t Au and the possibility of open pit mining to a depth of 300 m. As of 2004, about 10 percent of the deposit had been mined and processed by heap-leaching 1 million tonnes of ore

per year, producing 600 to 900 kg of gold per year. Plans had been made to increase production to 7.4

million tonnes of ore by constructing a new gold extraction plant which was put on stream in 2010. More recent, detailed information on the mine is difficult to find. Glencore itself reports that Glencore

owns 51% of Kazzinc and Kazzinc owns 100% of the Vasilkovskoe gold mine and 48.3% of the

Novoshirokinskoe gold mine in Russia. Together, Kazzinc’s six base metal and precious metal mines

have an annual production capacity of 300,000 tonnes of zinc metal; 130,000 tonnes of lead metal; 280,000 tonnes of copper concentrate; 700,000 troy ounces of gold and 6,000,000 troy ounces of silver

(http://www.glencore.com/kazzinc.html ).

6 DOMBRALY PROJECT

The information sources on which this report is based are cited in the text in abbreviated form and

described in detail in the References and Sources of Other Information section below. A major report

by Begayev and Teleshev et al. (2007), approved by SevKaznedra, a department of the Ministry of

Energy and Mines, describes the Dombraly Project in great detail as on 1 January 2007. The most recent descriptive summary of the Dombraly exploration project is by Alhambra (January 2011a). The

following report sections are based on these sources. The following text is complemented by Figure 2 to

Figure 6 inclusive.

http://www.kazakhgold.com/documents/327/Swift%20-%20Investor%20presentation.pdf

http://www.glencore.com/kazzinc.html

FIGURE 2: GEOLOGICAL MAP OF THE DOMBRALY AREA

WITH SOIL AND ROCK CHIP SAMPLES

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FIGURE 3: GEOLOGICAL MAP OF THE DOMBRALY AREA WITH CROSS SECTION LINES

(L16, L40) AND LONG SECTION LINE (A-B)

(GK - ZONE 13 - PULKOVO 1942)

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FIGURE 4: CROSS SECTION OF THE DOMBRALY DEPOSIT ALONG LINE L16

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FIGURE 5: CROSS SECTION OF THE DOMBRALY AREA ALONG LINE L40

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FIGURE 6: LONG SECTION OF THE DOMBRALY AREA ALONG LINE A-B

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6.1 DOMBRALY - PROPERTY DESCRIPTION AND LOCATION

Available information on property description and location, which is common to all the exploration

projects, may be found above in the report section with that name. The following information comes largely from Alhambra (2011a) and Begayev et al. (2007).

The Dombraly (also known as Dombraly II) gold exploration project is located in Enbekshilder District

of the Akmola Province in northern Kazakhstan, 64 kms north-northeast of the town of Stepnogorsk and 88 kms southeast of Alhambra’s Uzboy open-pit gold mine. The project area lies well within the East

Area polygon of Alhambra’s Licence MG 1029-D.

The operational centre of the project is the partially backfilled open pit, independently located on

Google Earth at Latitude 52°54'55.27"N and Longitude 72°04'32.41"E. The known gold mineralisation

in and adjacent to the old open pit has an estimated north-south strike length of 600 m and a total area of

approximately 20 hectares or 0.2 km². The prospective area is enlarged by outlying soil arsenic anomalies defined by the 50 ppm As contour and highly gold-anomalous rock chip samples over an area

of 2.5 kms north-south by 2 kms east-west, with the old open pit centrally located on the western side.

The ongoing exploration programme is to test and quantify the remaining oxide resources below and

adjacent to the abandoned open pit and to test the various mine waste materials for suitability for gold

extraction by leaching. Additionally, the work is extended to test the lateral, oxide and deeper, sulphidic extensions of the deposit and to identify other gold bearing occurrences in the area. All enterprises in

Kazakhstan must conduct environmental monitoring, provisions for which are made in three stages:

prior to the start of operations;

during the progress of work;

after the completion of work.

The first stage of ecological monitoring has been carried out while conducting hydrogeological studies. The background characteristics of the environment were determined: soils, water and air and a

programme for monitoring for the next stages was developed.

Ecological monitoring will be carried out within the framework of an agreement between Saga Creek

and the Karaganda Branch of Azimuth Energy Services on the basis of a scheme approved by the

Territorial Environmental Control Administration.

6.2 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE

AND PHYSIOGRAPHY

Dirt roads connect the area to the nearest villages of Zolotaia Niva (15 kms to northwest), Valikhanovo

(30 kms to southwest) and Aksu railway station (50 kms to south). From the latter there is a 16 km

asphalt road to Stepnogorsk (pop. 60,000), the site of Alhambra’s gold extraction plant and the Saga Creek subsidiary operating headquarters. The dirt roads are fit for transport only in summer time.

Farming is the main industry of the region.

The project area lies on part of an extensive plain with low relief ridges and wide internal-drainage

valleys and lake depressions. Elevations range from 182 m to 321 m above sea level, with local relief of

1 to 5 m, less frequently 8 to 10 m. The average altitude in the vicinity of Dombraly is 240 m.

The region is characterised by widespread shallow swampy depressions, fringed by small bushes and

aspen and birch forests. It belongs to the tipchak-feathergrass steppes of North Kazakhstan. The swamps

regularly dry up, appearing again after the spring melting of snows and in rainy periods.


16

The region has a sharp continental climate. The coldest months are January and February with average

monthly temperatures of -17C to -20C with the minimum reaching -35C. June and July have the

highest average monthly temperatures of 18C to 22C with the maximum reaching 35C. It starts to freeze in September, and warm weather comes in the middle of May. Hence, there is a restriction on the

operating season for exploration and this will probably also apply for mining operations.

Annual rainfall ranges from 250 to 300 mm. The long-term average yearly precipitation is 268 mm. The

major portion of precipitation falls in autumn and winter. July and August are the rainiest summer

months. The prevailing wind directions are westerly and southwesterly in summer; northwesterly and westerly in winter. High winds and dust storms bring problems for farming prior to the hot summer

season.

Electric power for Dombraly is supplied by a 35 kV line from the village of Zolotaya Niva 15 kms to

the northwest. Adequate quantities of potable and process water for mining needs could be supplied

from local boreholes. Factors such as plant sites, potential tailings and waste disposal sites will be considered as the project develops.

With regard to the sufficiency of surface rights for mining operations, the availability and sources of

power, water, mining personnel, potential tailings storage areas, potential waste disposal areas, heap leach pad areas and potential processing plant sites, no difficulties are foreseen by ACA Howe since

mining is clearly an acceptable and very active industry in this general area, Alhambra is already

producing gold from Uzboy, the land is generally flat and sparsely populated and this advanced exploration project is relatively close to the town of Stepnogorsk, Alhambra’s operational headquarters.

6.3 DOMBRALY - HISTORY

6.3.1 CHRONOLOGICAL OPERATIONS

The Dombraly deposit was discovered in 1966 by the northern party of “Zholymbet” Prospecting - Exploration Group who worked on it during 1966-1968. Further exploration was conducted between

1969 and 1972 by the Tselinograd Prospecting - Exploration Group of Central - Kazakhstan Territorial

Geological Administration.

In 1969 to 1972, the work included a topographic survey, construction of regional and local maps,

trenching, shaft sinking, underground level workings, core drilling and sampling. Surface exploration and sampling was by means of trenches and bell-pits, spaced at 12 to 25 m. Underground sampling and

mapping was carried out at level -30 m by drifting and cross-cutting every 20 to 25 m. Drilling was

carried out on grids 20 х 40 m, 40 х 60 m and 40 x 80 m to a depth of 100 to 180 m.

Summaries of exploration activities during this period are given in Table 2 and Table 3 below.

TABLE 2. SUMMARY OF DOMBRALY EXPLORATION 1969 TO 1972

Type of Work Units 1969-1972

Prospecting and exploratory drilling Linear metre 3785.6

Test drilling Linear metre 1840.0

Mechanical trenching Linear metre 5416.0

Bell-pits Linear metre 232.8

Shafts with section 9m2

Linear metre 32.2

Cross-cutting 4m2

Linear metre 720.0


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TABLE 3. SUMMARY OF DOMBRALY SAMPLING 1969 TO 1972

Type of Work Units 1969-1972

Core sampling Linear metre 3732.8

Chip (spot) sampling Pieces 425

Subsurface channel sampling Linear metre 1954.7

Trench channel sampling Linear metre 887.2

Trench spot sampling Pieces 63

Exploratory core drilling was carried out to study the mineralisation at a depth, below the oxide zone; to

trace controlling structures and to find new mineralised zones and quartz veins. Drilling was carried out

on lines 40 to 60 m apart, with 40 to 80 m between each hole, bearing 260 degrees magnetic at an angle

of 80 degrees and across the strike of mineral bearing structures and zones. Hole diameters were 127

mm to depths of 5 to 6 m, then 108 mm to depths of 50 to 60 m and then 90 mm to the end of each hole.

Drill core was split along its axis and sampled in lengths of 0.2 m to 1.2 m in the mineralisation, 1 m to

2 m for weaker mineralisation and 3 m to 5 m for unmineralised rocks.

In 1969 to 1972, geophysical well-logging of those drill holes deeper than 200 m was carried out by a

specialist team of the Central Geophysical Expedition used the following suite of tools: apparent

resistivity (AR), spontaneous potential (SP), electrode potential (EP); natural radioactivity; hole deviation; IP exploration - drilling holes; SP exploration - drilling holes; electrical correlation method

(ECM) - direct current. Geophysical studies were conducted using an automatic electronic well-logging

station (AEKS). Diagrams were drawn to 1:200 scale. Gamma-logging was carried out using a radioactive logger (NGGK – Neutron Gamma-Gamma Logging device and RSKM – Radioactive

Logging device).

During the mining operations, systematic sampling was carried out. Channel samples with a cross-

section 5 by 10 cm and from 0.1 to 1.2 m long were collected. In the shaft, the samples were collected

from all four walls. From the eastern and western walls, continuous vertical sampling was carried out

with sample lengths from 0.2 m to 1 m. From the northern and southern walls, horizontal channels were sampled at vertical interval of 1 to 1.5 m. In drifts, working faces were sampled at intervals of 1 to 2 m

by horizontal sections. From every section two or three channel samples were collected with lengths of

0.3 to 1.2 m. In the crosscuts, opposite walls were sampled by continuous sectional channels with the length of sections not exceeding 1.2 m.

Collection of channel samples from trenches and pits was carried out using the same procedure as for underground workings. However, samples taken from unaltered rock were over lengths of 3 to 5 m.

Sampling in the open pit was carried out in channels on every bench level. The channels were cut across

the strike of mineralisation at intervals of 10 m along the strike, the length of individual samples ranging from 0.1 to 1.0 m.

Quartz veins with thickness of more than 15 cm and their wall-rocks were sampled separately.

Two gold-bearing bodies were delineated in the oxidised zone. Oxide zone resources were estimated to

a depth of 170 m from surface using a system of geological blocks. The oxide resources were approved

and certified by the Territorial Reserves Committee, Production Geological Enterprise in Protocol # 3-411 of February 27, 1981.

At a cut-off grade of 3.0 g/t Au the following resources were approved:

Category С1 – 87793 tons of ore, 949 kg of gold at an average gold grade of 10.8 g/t

Category С2 – 77017 tons of ore, 631 kg of gold at an average gold grade of 8.2 g/t.

From 1985 to 1988 open pit mining to a depth of 60 m in the oxide material was carried out by a small

mining cooperative, Prospecting Crew “Enbek” of Mining and Processing Works “Kazzoloto”. The


18

cooperative is reported to have produced 140,000 tonnes of ore grading 6.96 g/t Au, using a 2.5 g/t Au

mining cut-off and produced 949.0 kg of gold.

In summary, since 1966, the Dombraly prospect has been intensely explored by various methods

including trenches, pits and shallow underground workings. Wider spaced drilling has probed the

transitional material and sulphides to a depth of 170 m. From 1985 to 1988 open pit mining produced 949.0 kg of gold. The pit has been partially backfilled and reclaimed.

Saga Creek, Alhambra’s wholly owned operating subsidiary in Kazakhstan, resumed exploration in 2002, the details and results of which are described in later sections of this report.

6.3.2 HISTORICAL MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

An English translation prepared by Alhambra’s Calgary office, of a report by Begayev et al. (2007)

includes a summary of Kazakhstan standard ‘reserves’. In March 2006, the State Reserve Committee of

the Republic of Kazakhstan provided Saga Creek with detailed tabulated recommendations concerning estimation of resources and reserves. According to Begayev et al. (2007), these recommendations have

been completely executed. A report was required containing a resource estimate in categories С1 and С2,

and the estimate to be approved by the State Geology and Subsurface Management Committee of the Republic of Kazakhstan.

The given category C2 ‘reserves’ of the oxidised ores down to approximately 60 m, plus the backfilled rock mass within the pit outline plus the waste pile are here rounded down to 4.25 million tonnes at an

average grade of 1.52 g/t Au containing 6.48 tonnes of gold. The given category C2 ‘reserves’ of

sulphide ‘ores’ to a depth of 300 m in category C2 are here rounded down to 0.58 million tonnes at an

average gold grade of 5.42 g/t Au containing 3.12 tonnes of gold. Inferred resources in category P1 to a depth of 500 m are here rounded down to 0.51 million tonnes at an average grade of 6.02 g/t Au. These

results are tabulated below in Table 4.

Begayev (2007) describes inferred P1 Category Kazakhstan standard resources which are those in the

Table 4 below. It is acknowledged that an opportunity to increase the Kazakhstan standard ‘reserves’ of

the Dombraly oxides is limited to the southeastern flank of the known mineralisation. At present, this area is covered by waste rock and exploration work is thereby precluded except in the long term.

However, there is a possibility that oxides located 10 to 15 kms to the northeast at North Balusty may

eventually be integrated with those of Dombraly.

TABLE 4. DOMBRALY RESOURCE AND ‘RESERVE’ ESTIMATE 2007

Type Category of

certainty

Tonnes

million

Grade

g/t Au

Gold

t

Oxide + backfill + low

grade stockpile

C2 4.25 1.52 6.48

Sulphide to 300 m depth C2 0.58 5.42 3.12

Sulphide from 300 m to

500 m depth

P1 0.51 6.02 3.07

E. Plyushchev (2011, pers. comm., 26 January.) wrote that the Dombraly ‘reserves’ and resources were

reviewed by the Kazakhstan State Reserve Committee as follows in Table 5 below, in which numbers

are rounded to the second decimal place.


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TABLE 5. DOMBRALY OFFICIAL RESOURCE AND ‘RESERVE’ ESTIMATE

2007

Type Category of certainty

Tonnes million

Grade g/t Au

Gold t

Oxide C2 3.68 1.62 5.97

Oxide P1 0.57 0.89 0.51

Sulphide C2 0.58 5.42 3.12

Sulphide P1 0.51 6.02 3.07

It is clear from a comparison of the above Tables 4 and 5 that 4.25 million tonnes of C2 oxide of Table 4 have been officially reclassified as 3.68 million tonnes of C2 and 0.57 million tonnes of P1 category

oxides in Table 5 above.

These Kazakh standard ‘reserves’ and resources estimates are probably reliable in the context in which they were estimated and based on the available information at the time. However, these estimates are

not described in accordance with the categories set out in NI 43-101.

Formal, independent, CIM compliant resource estimation work on Dombraly was curtailed in August

2008 after the related site visit by an ACA Howe geologist (ACA Howe, 2 June 2008). The status of the

Dombraly resource estimation project was described by another ACA Howe geologist (ACA Howe, 6 September 2008).

CIM compliant mineral resources and mineral reserves had not been previously estimated for this

relatively advanced exploration project due to the fact that any CIM compliant resource estimates would be fragmentary, for two main reasons:

1. Only the top metre of Saga Creek’s RAB drill samples of the Dombraly waste dump were useable under CIM due to probable contamination of deeper samples in holes up to 14 m deep (White, 2008).

2. The drill sample and assay data of nine angle core holes drilled at depth into the sulphide zone below

the open pit were available. However, the drill and sample data above these intercepts up to the floor of the presently inaccessible flooded pit are from non-compliant historical drilling. The drill and assay data

from ten RC holes surrounding the pit were not then available. Therefore, available drill data could be

used only to estimate the deeper, presumably sulphidic, resource and not the resource immediately adjacent to the pit slopes and floor.

6.3.3 HISTORICAL ESTIMATES OF MINERAL POTENTIAL

The quantity and grade, expressed as ranges, of potential mineral deposits at Dombraly which is a target

of ongoing exploration are described below. The potential quantities and grades are conceptual in

nature, there has been insufficient exploration to define CIM compliant mineral resources and it is uncertain if further exploration will result in the target being delineated as a CIM compliant mineral

resource.

Alhambra (January, 2011a) reporting a News Release dated May 15, 2008, have estimated the oxide

potential as 5 to 7 million tonnes with gold grades ranging from 1.40 to 1.80 g/t Au based on:-

results from the historical and current exploration and historical mining information;

core drilling completed in 2005 which showed that the oxide portion of the gold mineralisation

extends to a depth of 100 m with a minimum strike length of 400 m, which is open along strike to the north and south;

the average grade of all channel samples collected from the waste pile produced from the old

open pit.


20

That being said, Saga Creek has prepared a detailed report, including Kazakhstan standard resource

estimates in ‘reserve’ category C2 and resource category P1, which may form the basis for a prefeasibility study in due course (Begayev et al., 2007).

Prior to the CIM compliant resource estimates reported below in 2012, the estimates of the potential

quantities and gold grades for NI 43-101 compliant disclosure were as follows, combining the C2 ‘reserve’ and P1 resource estimates reported by Begayev et al. (2007) and revised by the Kazakhstan

State Reserve Committee:

Dombraly oxide zone potential mineralisation plus backfill plus low grade stockpile:

4 to 5 million tonnes with between 0.9 and 1.6 g/t Au.

Dombraly sulphide zone potential mineralisation to a depth of 500 metres

0.9 to 1.3 million tonnes with between 5.4 and 6.0 g/t Au.

6.4 DOMBRALY - GEOLOGICAL SETTING

6.4.1 DOMBRALY - REGIONAL GEOLOGY

The regional geology is common to all the exploration projects and is described above in the report

section with that name.

6.4.2 DOMBRALY - LOCAL AND PROPERTY GEOLOGY

Middle to Upper Ordovician volcano-sedimentary rocks underlie the project area, including sandstone, tuffaceous sandstone, siltstone and siliceous limestone. The upper part of the sequence is composed of

porphyritic pyroxene andesite-basalt lava and tuff, interbedded with felsic volcanics. All rocks are

intruded by Late Ordovician sub-volcanic dolerite, andesite and trachyandesite. The sequence has been

isoclinally folded with fold limbs dipping northeast and with northward plunging fold axes.

The limb of the synclinal fold that contains the deposit, is complicated by two systems of faults:

1) The main fault or dislocation is a northeasterly trending thrust-breccia zone, dipping to the east and

southeast at angles of 60 to 70 degrees. The thickness of the zone on the southern limb of the fold

ranges from 3 to 6 m. On the northern limb of the fold and beyond, the zone could not be followed by

workings due to the presence of a thick clay-like material reaching a depth of 150-170 m. A small gold occurrence has been found within the breccia zone.

2) The mineralised host structures are a series of sub-parallel, northwest trending shatter zones that are associated with the main fault in the form of fringing left-lateral branches. The dip of the fault plane

ranges from 45 to 65 degrees eastwards. The shatter zones have been identified in mine workings, in

drill holes and by using geophysical methods. They are intensively chloritised and show an increased content of arsenic. The shatter zones are filled with angular fragments of tuffaceous sandstones,

siltstones of various colours (from greenish-grey to lilac) and limestones with a chloritic matrix.

Both faults are accompanied by numerous ruptures and shear fractures filled with quartz. Often, where faults come together, a network of fine, en-echelon cracks has formed, giving rise to stockwork

mineralisation.

6.5 DOMBRALY - DEPOSIT TYPES

The mineralisation is described as belonging to a volcano-sedimentary hosted, orogenic type and is represented by quartz veins, quartz veinlets and disseminated mineralisation zones containing native


21

gold and pyrite with sporadic galena, arsenopyrite and specularite (haematite) in the primary

mineralisation. The tuffaceous sedimentary rocks hosting the gold mineralisation are propylitised. The highest degree of hydrothermal-metasomatic alteration is found in the central part of the altered zone,

where the rocks are sericite-chlorite-quartz metasomatites. They contain broken areas penetrated by fine

veins and streaks of quartz within a siliceous matrix impregnated by sulphides.

The primary mineralisation is deeply weathered and oxidised to variable depths in excess of 60 m below

surface reaching maxima in tectonic zones. For example, in drill hole DDD 48-04 oxidation extends to

380 m from surface. The gold bearing oxides are the main target for short term open pit exploitation.

6.6 DOMBRALY - MINERALISATION

The mineralisation occurs as quartz veins, quartz veinlets and disseminated mineralisation zones

containing native gold and pyrite with sporadic galena, specularite (haematite) and arsenopyrite.

Metallurgical samples taken from the primary zone of gold mineralisation indicated sulphide content

less than 2% although some maximum values of 15 to 18% were established. All zones and quartz veins strike north-northwest and dip to the northeast at 30 to 65 degrees. Quartz veins are up to 350 m long

and 1.5 m wide while veinlet and disseminated mineralisation zones reach 600 m in length and 90 m in

width, remaining open at both ends. They are explored to depths of 360 m and are open down dip. The oxidation boundary is variable in depth reaching maxima in tectonic zones. In drill hole DDD 48-04

oxidation extends to 380 m from surface.

The mineralised bodies do not have obvious geological boundaries and have been separated only by

gold grade. Using a cut-off grade of 0.3 g/t Au, a main mineralised body and seven smaller lenses have

been defined within the limits of the zone of oxidation.

The Main Body contains 73.6% of the total gold. It has been studied by core drilling and underground

workings and has been traced along strike at surface. It has been tested to a depth of 300 m but has not

been explored down the plunge. Oxide ores were mined by an open pit to a depth of 60 m between 1985-1988. In plan, the body has an irregular lenticular shape, 300 to 500 m long and 50 to 60 m wide.

To the north, the body narrows sharply. It dips to the east at 45 to 60 degrees with a northerly pitch of

20 degrees in the zone of oxidation and 45 degrees at depth. In long section the body has an asymmetric lenticular shape and, in its lower part, it is split into a number of branches. A series of quartz veins and

zones of stockwork quartz occur along its axial plane.

Lenses 1, 2 and 3 lie immediately adjacent on the hangingwall side of the main lens.

Lens 1 is a branch of the Main Body. The gold content of Lens 1 comprises 1.3% of the total for the

whole deposit. The lens has a northwesterly strike with a northeast dip of 35 to 40 degrees. Its pitch matches that of the main ore body. In plan and in profile, the lens has a simple lenticular form. It is well

delineated by drill holes in the oxidation zone. In the primary zone, hole 401 hit only weak

mineralisation when testing its extension.

Lens 2 is also a branch on the hangingwall side of the Main Body. The gold content comprises 7.0% of

the total known gold. It is lenticular in plan and strikes northwest with a 40 degree dip to the northeast.

It pitches northwest at 50 degrees. The lens is well defined by holes in the oxidation zone and it has been tested by six drill holes to a depth of 250 m. The shape is a simple lens.

Lens 3 is another branch on the hangingwall side of the Main Body. It contains 9.5% of the total known gold. The lens also strikes northwest with a northeast dip at 40 degrees. It pitches northwest at 20

degrees. It appears to be well delineated by drilling.

Lens 4 is adjacent to the footwall side of the main body. It contains 2.3% of the total known gold. It strikes northwest and dips at 40 to 50 degrees to the northeast. It pitches at 70 degrees to the northwest.

It appears well delineated by drilling.


22

Lens 5 is located on the footwall side of the Main Body and is en echelon with Lens 4. It contains 3.4% of the total known gold. It strikes northwest with a northeast dip of 70 to 80 degrees. This lens has been

only partially delineated by drilling.

Lens 6 has been mapped on the hangingwall side of Lens 3. It strikes northwest with a northeast dip of 40 to 45 degrees. The lens has been only partially delineated by drilling.

Lens 7 lies on the footwall side of Lens 5. It also strikes northwest with a 70 degrees northeast dip. It is estimated the gold content of the lens may be 2.5% of the total for the whole deposit but further drilling

is required in depth and on its southeastern side.

There is another gold showing located 2.5 kms north-northwest along the main mineralisation trend at

the open pit called Liparitovoe.

6.7 DOMBRALY - EXPLORATION

Between 2002 and 2006 Saga Creek, the wholly owned subsidiary of Alhambra, explored the Dombraly

area including the down dip and strike extensions of the gold mineralisation. It included exploration of the mineralisation below the profile of the open pit and sampling of the backfill andlow grade stockpile.

During this period, 8 trenches totalling 1741 m, 613 RAB drill holes amounting to 4152 m and 10 core

holes for a total of 2394 m were completed. Sampling included 354 samples from the trenches, 2076 from RAB drill holes and 1958 from core holes. A summary of work during 2002-2005 is given in Table

6 below.

TABLE 6. SUMMARY OF DOMBRALY WORK 2002-2005

Description Area Units Number Total

Core drilling Open pit site m 10 2394

RAB drilling Around of open pit m 47 1004

RAB drilling Open pit m 49 424

RAB drilling Low grade stockpile m 110 1284

Trenches Low grade stockpile m3 3656.7

Trench channel sampling sample 1742

Drill cutting sampling sample 1361

Core sampling sample 1958

Technological (process) sampling sample 3

Bulk sampling sample 61

Technological mapping sampling sample 11

Blocks for determination of bulk

density sample 5

In 2005, a topographic field survey was carried out at 1:1000 scale during which all the old and new

exploratory works were integrated. The topographic survey was carried out by the appropriate

department of the Karaganda branch of Azimuth Energy Services Ltd.

Between 2004 and 2005, core drilling and RAB drilling programmes were carried out by

Stepnogorskaya GRP Ltd (Stepnogorskaya). For some of the core drilling, Stepnogorskaya used Boart Longyear SSK type wire-line equipment and double-tube core barrels complete with a hydraulic

submersible impact tool and an ejector. The exploratory core holes were drilled at an angle of 60

degrees at a diameter 108 mm for the first 5 m. Thereafter, the hole diameter was 76 mm, producing 42

or 40 mm diameter cores. Depending on the drilling conditions and equipment used, drill run lengths ranged from 0.3 to 1.5 m, with an average length of 1.0 m in normal conditions and reaching 2.0 m with

SSK equipment. The depth of inclined holes ranged from 97.0 m to 202.8 m; vertical holes from 44.0 m


23

to 54.0 m. The angle of 60о for exploratory drilling allowed for mineral intersections at an angle close to

90о. The average core recovery was 90 to 92% and 95% for Boart Longyear drill strings. The core

samples were marked and placed into standard boxes.

The air-flush RAB holes were intended to test the extension of the oxidised gold mineralisation at depth,

as well as above the level of ground water. These holes were drilled vertically with a diameter of 76 mm. Slurry recovery was in the range 95-100%. However, although drilling below the level of ground

water is technically feasible the slurry material is lost and enriched with heavy residues. For verification

of RAB drill sample quality, eight trenches 2.5 m deep were dug along the lines of holes. The trenches were tested with 2.0 m long vertical trench channel samples. The results of correlation of trench wall

channel sampling with upper intervals of the RAB drilling proved satisfactory, but not so at depth.

The grid for the RAB drill holes within the open pit envelope was based on lines 40 m apart with holes

spaced at 10 m intervals on the lines. The holes were drilled down to the level of the pit water. Vertical

air-drilled holes on the surface of the dumps were drilled on grids of 40 x 50 m and 60 x 50 m. Drilling

was continued to the base of the dump.

The east-west lines for exploration of the oxidation zone were laid out at 40 m intervals. Continuity of

the mineralisation along the extension of the various bodies was verified by reference to the position of the mine workings.

The achieved density of the exploration grid fully complies with the recommendations for prospecting and exploration of mineral reserves issued by the State Committee on Reserves of Kazakhstan with

regard to exploration of С2 reserves in category III of structural complexity.

The main mineralised body has been well-explored and completely delineated all over the area of the site and to its full depth of oxidation, except at its southernmost end. Here it is narrow and of low gold

grade.

The mine “waste” in the open pit was explored by air flush drilling down to the level of the pit water.

The level of exploration of the dumps is sufficiently uniform, and those reserves may be classified as

category С2.

In 2010-2011 Saga Creek completed further testing of the mineralised low grade stockpile, pit infill and

in-situ gold mineralised zones. Exploration of the oxide, transitional and primary zones was done by

RC drilling and diamond drillcore holes using Longyear equipment.

Exploration core drilling was done by skid-mounted drilling rigs ZIF-650 and SKB-5. Thirteen holes for

a total of 3441 m were drilled at an angle of minus 60 degrees westwards. Drilling was done with 108 mm diameter equipment to a maximum depth of 5 m. After installation of a collar casing, the drilling

continued using Longyear double-tube core barrels. The drilling diameter was 76 mm and the diameter

of the core was 40-42 mm. Average core recovery was greater than 90%. Core was marked and stored in

standard boxes.

Thirty-seven RC holes were drilled for a total of 880 m to test oxide gold mineralisation in the

mineralised low grade stockpile and pit infill zones. The diameter of drill rods was 76 mm. Recovery of slurry material was 59-100. Within the low grade stockpile, drill holes (RCW prefix) were vertical,

within the pit infill zone, RC drill holes (RCD prefix) were angled minus 60 degrees to the west.

To confirm the quality of the earlier RAB drilling in the mineralized low grade stockpile, a total of 8

trenches with a depth of 2.5 m were excavated along the drill lines. Trenches were sampled vertically

with sample length of 2 m. Comparison of trench and RAB drill results show satisfactory correlation.

Significant intercepts from Alhambra drilling conducted on the project are presented in Tables 7 and 8

below.

¹


24

TABLE 7. 2010-2011 DDH SIGNIFICANT INTERCEPTS

Hole ID From To Length DataType Au Fa Area East North RL

DDD1502 0.0 1.0 1.0 core sample 2.03 In Situ 13303515 5868843 230





















DDD4803 296.5 297.5 1.0 core sample 0.88 In Situ 13303430 5869482 -28

DDD0701/1 78.3 79.9 1.6 core sample 0.59 In Situ 13303396 5868930 160
















DDD4804 415.4 417.4 2.0 core sample 3.06 In Situ 13303474 5869483 -

132

¹ Intercepts calculated using a 0.5g/t Au trigger value, minimum width 1m, no more than 2m consecutive waste, overall grade

greater than 0.5g/t Au.¹


25

TABLE 8. 2010-2011 RC SIGNIFICANT INTERCEPTS

Hole ID From To Length DataType Au Fa Area East North RL

RCD0801 0 24 24 RC chip sample 1.14 Pit Infill 13303408 5869093 207


RCW0101 5 7 2 RC chip sample 0.90 Low grade stockpile 13303264 5868493 241
























26









RCD1402 23 26 3 RC chip sample 1.00 In Situ 13303350 5869146 195











¹ Intercepts calculated using a 0.5g/t Au trigger value, minimum width 1m, no more than 2m consecutive waste, overall grade

greater than 0.5g/t Au.

6.8 DRILLING

The following information was compiled from Saga Creek’s report of 2007 (Begayev et al., 2007) and modified in consultation with Saga Creek, for the description of sampling methodologies prior to the

recent 2010-2011 exploration programmes.

It is reported that systematic sampling of mine workings was carried out between 1969 and 1972.

Channel samples were collected from the walls of the shaft and drives. Additionally, samples were taken

from trenches and pits within the open pit as described above under History.

In 1969 to 1972, two 600 kg bulk samples were collected from the underground mine workings at the

30-m horizon for a study of gold recovery. During 1969 to 1972, thirty-five waxed samples were

collected from mine workings and openings and sent to the Aksu Mine laboratory for the determination of bulk density.

Slurry sampling was conducted during RAB drilling. A 1.5 to 2.0 kg sample of the cuttings was split from the discharge and then further split into a sample for analysis and its duplicate. This control of the

slurry sampling for RAB holes was developed during the exploration of the oxidation zone of the Uzboy

deposit in 2003. A total of 68 duplicate samples were analysed and compared to their originals and it is

reported that variation in results was calculated as plus or minus 2.4%.

Diamond drill samples ranging from 0.8 m to 1.5 m in length depending on the nature of the rock are

split longitudinally by diamond saw. The resulting sample weighed approximately 1.3 to 2.8 kg with a core diameter of 40 – 42 mm. In the oxidised zone, where the core presented a loose mass, the samples


27

were divided into two manually under a geologist's supervision. The samples were packed into bags

marked with the hole number and depth interval, and sent to the laboratory of KhAL “Quartz” in Semipalatinsk for sample preparation. The remaining core was packed into wooden crates for storage in

the core-shed of Saga Creek in Stepnogorsk.

In 2004, a bulk sample weighing 525 kg was collected from the open pit for heap leaching tests on the oxide material. Processing studies were conducted by Kazmekhanobr of Almaty.

Table 9 below is a summary of sampling derived from the report by Begayev et al. (2007).

TABLE 9. SUMMARY OF DOMBRALY SAMPLING QUANTITIES

Description Units 1969-1972 2002-2005

Collection of core samples Linear metres

3732.8 1958

Collection of chip samples from holes Pc. 425

Collection of slurry samples from RAB holes Pc. 1361

Collection of channel samples from underground Linear

metres 1954.7

Collection of channel Samples from Trenches Linear metres

887.2 1742

Collection of chip samples from trenches Pc. 63

Collection of pillars or blocks for bulk density Pc. 3 5

Collection of bulk samples Pc. 5 61

Collection of samples for technological mapping Pc. 11

For the recent (2010-2011) exploration programmes, geological exploration works on the deposits were

carried out according to conventional methods. Documentation of the mine workings and holes was carried out directly at the mine sites. Sampling was conducted at the Saga Creek gold processing plant.

Core documentation was carried out during the core drilling process on a daily basis. Geological information was entered into the logs recording drilling intervals (runs), the core length, % core

recovery, core sketches of separate fragments, lithologies and mineralisation, sampling intervals and

sample numbers. During the performance of geological work the hole is closed by the district geologist

and check measurements of the hole are taken.

Once the hole is completed, core handling and processing is undertaken, which includes geological logs,

drilling logs (conducted by the drilling foreman with filling in the drilling parameters and possible geology-technological information and issues), start and completion dates of hole drilling, survey

measurement. Drill core was routinely photographed.

Drill holes documentation for logging and processing is hardcopy only, later transferred to electronic format for utilisation in 3d modeling software.

Core boxes were laid out in drill sequence for processing and sampling. The geologist marks sample intervals, putting sample labels along sample interval boundaries strictly identified by the logging and

sample documentation. MCS have reviewed geological logs and are satisfied they capture the pertinent

and relevant information

Following the geological documentation and the identification of ore zones (mineralisation) and sample

interval marking core is cut along the long axis of the core. Half core for the entire length of the drill

hole is collected for submission to the laboratory.


28

Core cutting is made under the supervision of the geologist, at Stepnogorskaya Mining Enterprise

facility, a subcontractor of the drilling company. Diameter of diamond disc is 400 mm. Soft or broken material is split in half by hand.

Once the core is split into two halves, the sampler selects half core samples and places into plastic sacks

according to the sample intervals. Paper label with sample information (site, hole number, sample number, sample interval, family name of the geologist, date of sampling) is inserted into the sample

sack. Hole and sample number are signed on the sample sack. All samples from the holes are sent to the

core storage facility, where sample weighing and sample group formations for the sample preparation are performed.

These samples are weighed (scales type РН-10Ц134), with a scale accuracy of 0.05 g and are entered in the sample registration log.

Following the cutting, core boxes for the drill holes remaining core are closed and transported to the

Saga Creek Gold Company’s core storage warehouse.

Later, if necessary, metallurgical test samples are selected from the remaining core for mineralogical

and technological studies.

MCS and ACA Howe consider drill core handling and processing methodology satisfactory.

For RC drilling, a standard and acceptable approach and practice is employed. The rock chips coming

from the hammer to the cyclone through dual tube were split using a rifle splitter by the drilling

contractor under the supervision of the Saga Creek geologists. Each meter of hole was sampled. The

average sample weight was 5 kilograms.

Samples were packed in double cotton bags, properly marked and sent to Stewart Assay and

Environmental Laboratories – Kyrgyzstan.

Standards, blanks and duplicates were inserted into the sample stream submitted to the laboratory after

each 25th routine sample.

6.8.1 TRENCHING

Trenches were dug by excavator along 40 and 80 m spaced profiles to the depth of 1.5 m. The trenches were mapped and sampled manually by taking one-meter long channel samples. The samples weights

ranged between four and seven kilograms.

6.8.2 BULK DENSITY

Bulk density measurements are at present available for the low grade stockpile and pit infill material and

the old DDH (C* prefix) drilling.

These measurements were taken on-site via core and bulk sample measurements.

In 2004-2005 determination of a bulk weight and natural moisture of ore in situ was done on the basis of

paraffin-lined samples applying standard methods in the laboratory of Reaktiv LLP. The total number of samples collected was 55, and 26 of them were the samples of oxide ore from middle and lower

horizons of oxide zone and 29 samples on gold sulphide ores.

It was determined that the bulk weight of oxide ores was 2.58 t/m3, for gold sulphide ores – 2.62 t/m

3.

Estimated average values of bulk masses are in good consistency with bulk weight dependency on depth


29

diagram. Thus, for oxide ores, average value of bulk weight corresponds to the interval of depths of

160-90 m, within the limit of which the principle reserves of oxide ore are located.

In 2004 bulk samples from the surface and sub-surface parts of the mineralised waste dump were

collected. A total number of 5 bulk samples with weights from 56.8 to 61.8 kg were collected.

Weighing was done after collection and drying the samples for 2 days.

Bulk weight of re-cultivating formations in the outline of open pit was 1.73 t/m3, for technogenic

formations of the low grade stockpile it varied from 1.58 to 1.77 t/m3 with the average of 1.66 t/m

3.

Due to the fact that the bulk weight was determined in partly dried samples, adjustment to moisture was

not applied to density calculations.

Bulk density determinations are yet to be taken for recent Alhambra exploration drilling (2010-2011

drill programmes).

Bulk density for use in current resource estimations is discussed further in section 6.12.14 of this report.

6.9 DOMBRALY - SAMPLE PREPARATION, ANALYSES AND SECURITY

Between 1969-72, analytical operations were carried out in the chemical laboratories of Aksu,

Zholymbet and Tselinograd Exploration Company.

The quality of the analysis of samples collected at the Dombraly deposit in 1969-72 was systematically

checked twice a year by the internal audit system of the laboratory by repeated analysis of coded

duplicates of the original samples.

The internal assay quality audit was based on 3-3.5% of the total number of samples.

The external audit (based on 3-4% of samples) of assay operations of the Tselinograd laboratory was

carried out in the laboratory of Aksu mine, and analyses of Aksu laboratory were checked in the

laboratory of Tselinograd Exploration Company. The functions of arbitrary control were imposed on the Central Laboratory of Kazzoloto.

The total of 110 samples was sent for internal and external geological control. This data has not been

made available to ACA Howe/MCS for review.

In 2002-2005 the following types of assay works were carried out:

Fire assay of core, trench and chip samples for gold (3241 analyses);

Atomic-absorption analysis for gold of core and chip samples (2309 analyses);

Combined samples analysis.

Fire assays were done at Tsentrgeolanalit CJSC laboratory (Karaganda), atomic-absorption analysis was

carried out in Reaktiv LLP and Kvarts Chemical Analytical Laboratory.

Combined samples were analysed in Tsentrgeolanalit CJSC.

Fire assay analysis was undertaken for gold, atomic-absorption analysis for copper, lead, zinc and silver, spectral assay for 16 elements, complete silicate analysis, neutron activation analysis for platinum,

chemical analysis for total sulphur, sulphide sulphur, sulphated sulphur.

Internal geological verification of fire assays was conducted in Tsentrgeolanalit CJSC, external

geological verification – in the laboratory of Kazmekhanobr.


30

Duplicate samples were sent for internal and external verification. A total of 589 samples were sent for

both types of control compiling 5.1% of the total number of assay tests for this period. Fire assay tests were not conducted and samples for geological verification were not collected during the second half of

year 2005.

The results of external and internal geological verification of 1969-2006 show good reproducibility of results. However, the majority of the data is unavailable and without data for the submission of

reference samples (standards), blanks, or duplicates for the historical data the accuracy of values for the

1969-2006 cannot be satisfactorily verified.

QAQC data for the 2005 EW* prefix low grade stockpile trenches has been obtained, and studied in

section 6.10, Dombraly Data Validation.

Prior to 2009, samples were prepared at commercial facilities within Kazakhstan. In 2002-2004,

the sample processing was carried out in the Reagent LLP (Stepnogorsk), in 2005-2009 – in Quartz

Chemical Analysis Laboratory. Over the entire period of gold mine exploration grinding of the gold samples was carried out at K equal to 0.5.

Between 2009 and 2010 exploration works samples have been prepared in the sample preparation facility of Saga Creek Gold Company’s gold processing plant in Stepnogorsk. Sample processing is

performed using the Richard-Chechett' formula Q = kd2 at k = 0,5.

Experimental validation of the coefficient of uneven mineralisation was carried out on the field in 2005.

Subsequent to drying, samples are sent to the first stage jaw crusher, where sample material is grinded

to approximately 7 mm. Following the first stage of grinding crushed material is sent to roller crusher, where it is grinded up to 1 mm.

Sample rescreening is performed following the roll crusher through the sieve with a cell of 1 mm. Material not passed through the sieve returns to regrinding in the roll crusher.

After the second phase of fragmentation and reduction by the Jones index the bulk of crushed to 1 mm

samples enters the geological sample storage as a geological sample duplicate, while sample weighing approximately 0.75 kg is reduced to a particle of size 200 mesh (disk pulveriser).

The scheme employed for core sample processing is shown in Figure 7.

FIGURE 7: SAMPLING PROCESSING SCHEME

1,5 kg

0,75 kg

Qн=kd=0,5

d=100 ммk

2Jaw crusher

Roller crusher

Duplicate

Initial WeightSample 1,5 kg

1,5 kg

Quartering

Sample 0,75 kg

Abrasion to 0,074 мм.

Analytical sample0,75 kg

At 200 mesh

cboyd

Typewritten Text

cboyd

Typewritten Text

31


32

Batches for both atomic absorption and fire assay are selected from the sample material, plus samples for the internal and external control. Remains of analytical sample are sent to the sample storage as

analytical duplicates. Sample duplicates are stored in a specially designed secure sample storage

building of the gold processing plant. All duplicate (residue) samples are strictly controlled. Geological

sample duplicates are packed in sacks, sacks are signed, stacked by holes, and are kept on special shelves. All samples are numbered and easily accessible.

All instruments used for crushing and sample reduction are equipped with special instructions for

operators. The sample preparation workshop is kept clean. After every sample preparation, all appliances and countertops are cleaned using compressed air, but equipment is not cleaned by crushing

inert material such as granite or glass. Sample preparation methodology at the SCGC facility is

considered satisfactory.

For the 2010 exploration programmes, sample preparation and fire assay that were done by Stewart

Assay and Environmental Laboratories – Kyrgyzstan used the following methodology: samples were

crushed to -2 mm, mixed and split into two 200 gram sub-samples. One sub-sample was pulverised to –200 mesh and the other sub-sample was retained for reference purposes. A 30 gram sample of the –200

mesh material was used for fire assay atomic absorption finish.

Stewart Assay and Environmental Laboratories, as a part of ALS Group, have stringent quality

assurance and quality control procedures and does have an International Standard Organization (“ISO”)

17025 accreditation.

For trench samples (2005), sample preparation and atomic absorption analysis was completed by

Chemical and Analytical Laboratory Quartz LLP located in Stepnogorsk using the following procedure:

samples were pulverised in a jaw crusher to -1 mm, mixed and split into two 0.75 kilogram sub-samples. One sub-sample is ground to -200 mesh and the other sub-sample is retained for reference

purposes. A 10 gram sample of the -200 mesh material is used for atomic absorption. For internal

control purposes 10% of the samples are re-analysed. Most of the samples returning more than 0.3 g/t Au were re-assayed by fire assay. The external control for the results of Quartz LLP was done by

Centergeoanalit Ltd. located in Karaganda, Kazakhstan. Standards, blanks and duplicates were inserted

into the sample stream submitted to the laboratory for analysis. Both laboratories, Quartz and

Centergeoanalit, are certified in the Republic of Kazakhstan but do not have any International Standard Organisation rating.

6.10 DOMBRALY - DATA VERIFICATION

6.10.1 QA/QC ANALYSIS

The quality assurance/quality control (QA/QC) analysis comes from a combination of information

from the geological exploration reports for the project, quality control data and information and observations gathered by MCS during the project site visit.

6.10.2 INTRODUCTION

Quality control monitoring is undertaken to ensure that the chemical data used are as reliable as

possible to meet the objective of the exploration and resource development program. In advanced

exploration projects, quality control and assurance programs are designed to ensure the high integrity of data fit for the purpose of obtaining reliable and accurate, reportable mineral resource

and reserve estimates.

There are three fundamental aspects to Quality Control:


33

Accuracy - How close are the assays to the ‘true’ content of metal in the samples?

Precision - How repeatable are the values from the samples?

Sampling Errors - Sampling errors and laboratory errors that may influence the

representativeness of a sample and the representativeness of the assay result from that sample.

6.10.3 QUALITY CONTROL SUBMISSION

For an NI 43-101 compliant study, it is necessary to present a QA/QC assessment study.

The frequency of QC sample submission will vary depending upon the type and stage of the

exploration program, for example:

When dealing with a soil sampling program where the requirement from the data is

precision rather than accuracy, the submission of expensive external ‘certified’ standards

is wasteful. Routine submission of ‘In house’ standard material, along with a few duplicated sites near any known mineralisation may well suffice. Overall, quality control

samples need not exceed say 1 in 50 (2%).

When drilling a prospect which has a good chance of becoming a resource, a higher

proportion of QC samples will be required. These might comprise of ‘in house’ standards

and blanks, duplicates, and certified standards. In addition, it may be prudent to conduct

occasional but regular cross checks of mineralised samples at other ‘umpire’ laboratories

(along with QC samples).

A general industry ‘rule of thumb’ for the proportion of QC samples submitted for drill

programs is 5%, comprising of 2% ‘Certified Reference Material’ (CRM) samples, 2%

Duplicate samples and 1% Blank samples. In practice, there is a requirement to include in the sample run 1 CRM every 50 samples, 1 Duplicate every 50 samples and 1 Blank every

100 samples.

At the Dombraly project QAQC sample data is currently available for the 2005 EW* prefix

low grade stockpile trenches, the 2010 DDD* prefix diamond drill holes and 2010 RCD* and

RCW* reverse circulation drilling.

6.10.4 QUALITY CONTROL SAMPLE MATERIALS

Quality control is assessed from the evaluation of analytical results from a combination of samples:

Primary Standards – Sample of known metal content and chemical characteristics. These can

be externally sourced commercial standards (CRMs) or company ‘in house’ standards.

Blank Samples – Samples that contain none of the metal in question. These are generally

sourced in-house from barren drill spoil, however quantities of material suitable for Blank standard material can be purchased commercially (road base/gravel).

Duplicates – These are splits of drill core, RC/Aircore/RAB drill cuttings, and outcrop samples

from the same sample interval.


34

In the present project only the exploration campaigns drilled from year 2010 present the fullest

QAQC procedure analysis. During the previous campaigns no blank or standard samples were

introduced in the analysis process, only duplicate samples, based in the soviet style exploration

data check requirements introduced for low grade stockpile trench sampling.

6.10.4.1 Primary Standards

Primary standards are used to verify the accuracy of analyses reported by mineral testing laboratories. They are homogenised samples that have been analysed numerous times, usually by

definitive techniques, so their ‘true’ metal content and the inherent variability therein are known.

Two types are used:

‘In-house’ Company Standards –

In the present project ‘In-house’ company standards were not used.

Commercial (CRM) Standards –

CRM samples sourced from SGS Australia were routinely submitted for assaying with core to test laboratory accuracy. A variety of CRM’s were used, representing high and low gold grades, and

oxide and sulphide ore. CRM’s were routinely included in sample batches sent to the Stewart

Kyrgyzstan lab for fire assay.

A total of 3,263 samples were analysed from the diamond drilling program and 1,001 for RC

program.

TABLE 10. NAME AND GRADE

OF CRM SAMPLES

CRM Lab name Au grade (g/t)

- 21.600

- 13.600

- 5.530

- 1.460

- 0.520

In 2010 a set of CRM’s were purchased from Rocklabs (New Zealand), via the consultant lab GMAR Development (Nevada, USA), which are now preferred CRM sample for submission.

6.10.4.2 Blank Samples


35

It is common practice to include samples of material believed to be barren of the metal/mineral

being sought to ensure that no background drift occurs at the laboratory and that no gross contamination is taking place.

In the present study two different blanks were used, one was a standard CRM sample with a known very low grade of 0.003 ppm Au, and one an in-house made sample with a grade <0.010

ppm Au (lower detection limit of the analytical lab).

6.10.4.3 Duplicate Samples

A common method of monitoring QC is to submit duplicate samples. Duplicate samples analysis allows for the determination of the natural nugget and sample error, pulp repeats allow for the

determination of assay precision. However, and of considerable importance, it does not monitor

accuracy.

A total of 132 and 40 pulp duplicates were taken for the 2010 exploration programmes. No field

duplicates were submitted.

6.10.5 QUALITY CONTROL ASSESSMENT

Once the QC samples have been included in a batch, it is essential that the results are evaluated.

The samples sent to the lab to be analysed were divided into four batches, with a total of 4,264 (3,263 samples from diamond drilling and 1,001 from RC drilling). The QA/QC samples

submitted in each batch are detailed in Table 11.

For diamond drilling there were submitted 128 blank samples (which represent 3.9% of the total

samples analysed); 126 standard samples (which represent 3.8% of the total samples analysed);

and 132 pulp duplicate samples (which represent 4.0% of the total samples analysed).

Stewart Kyrgyzstan lab introduced, as in-house QAQC procedure, 403 lab duplicate samples (pulp

duplicates) in the DDH 2010 batch’ samples and 151 in the RC 2010.

In RC drilling there were submitted 39 blank samples (which represent 3.9% of the total samples

analysed), 41 standard samples and 40 pulp duplicate samples (which represent 4.0% of the total

samples analysed).

In the drilling programs previous to year 2010 only duplicate samples were added to the sampling

analysis process, as QA/QC methodology. Based in the soviet style, the samples duplicated were

prepared as pulp duplicates in the laboratory where regular samples were prepared, and then, they were analysed in the same analytical lab and in an umpire analytical laboratory, as an external

control.

Ninety-two grab samples were duplicated in Quartz laboratory and re-analysed in Centergeoanalit

laboratory (original samples were analysed in Quartz laboratory). (Table 12)


36

TABLE 11. NUMBER OF QA/QC SAMPLES IN THE DIFFERENT 2010-2011DRILLING

CAMPAIGNS

Program

Total Number of Samples

Analysed

Number of

Blanks Samples

Number of

Standard Samples

Number of

Pulp Duplicate Samples

Number of Lab

Duplicates

Samples

DDH 2010 3,263 128 126 132 403

RC 2010 1,001 39 41 40 151

TABLE 12. NUMBER OF QA/QC

SAMPLES USED IN PRE-

2010 EXPLORATION

CAMPAIGNS (EW*

TRENCHES).

Sample Type Number of Pulp

Duplicate Samples

Grab 92

Results from CRM’s, field duplicates and lab duplicates are assessed in the following section.

6.10.6 QA/QC ASSESSMENT

6.10.6.1 Monitoring of Standards – Accuracy

Micromine software was used to create the analyses of the standards accuracy. The known metal

content from the standard sample is added to the graphs below as a (black) horizontal line (called

Ref Value), and +/- 1, 2 and 3 standard deviation lines as Upper or lower warning limits.

There are statistical objections to using +/- 2 standards deviations as action limits, as by definition

5% of all determinations will cause ‘failure’. This is where common sense is required.

An alternative common practice in statistical assessment of standard data is to apply an arbitrary

+/-10% of the ‘true’ expected value as the action limits for a standard. The results obtained were

plotted on a graph for each element of each standard sample.

Figure 8: Legend for Standard graphs


37

Diamond Drillholes DDH 2010

Figure 9: Diamond drillhole sample graph plot of Standard sample STD=21.600 ppm Au


Au

(p

pm

)

Number of Samples

Number of Samples

Au (

ppm

)


38

Figure 11: Diamond drillhole sample graph plot of Standard sample STD= 5.530 ppm Au

Figure 12: Diamond drillhole sample graph plot of Standard sample STD= 1.460 ppm Au

Number of Samples

Au

(p

pm

)

Number of Samples

Au (

ppm

)


39


Reverse Circulation Drillholes RC 2010

Figure 14: Reverse Circulation drillhole sample graph plot of Standard sample STD=21.600 ppm Au

Number of Samples

Number of Samples

Au

(p

pm

) A

u (

ppm

)


40



Number of Samples

Au

(p

pm

)

Number of Samples

Au (

ppm

)


41

Figure17: Reverse Circulation drillhole sample graph plot of Standard sample STD=1.460 ppm Au


Number of Samples

Au

(p

pm

)

Number of Samples

Au (

ppm

)


42

The review of external standard values plotted on control charts for diamond drilling 2010 program and

reverse circulation drilling 2010 program, indicate the results for Au fall within acceptable limits of one standard deviation of the CRM, indicating analysis done by Stewart Kyrgyzstan Lab is reasonably

accurate.

6.10.6.2 Monitoring of Pulp Duplicates – Precision

Analysis of pulp duplicate precision has been studied by Scatter Plot, using Micromine software.

The precision value shown in the graphs is an indication of variability in the differences between

individual X-Y values, relative to the average X value. A perfect result has a precision of zero. Values greater than zero represent an increasing amount of deviation. For gold, in coarse duplicate,

best level of precision is 20%, and acceptable 30-40%, and for pulp duplicate best level of

precision is 10%, and acceptable 20% (Abzalov, 2008).

In the trench programme completed in 2005 duplicates were processed following soviet protocol.

The samples duplicated were prepared as pulp duplicates in the laboratory where regular samples

were prepared, and then, they were analysed both in the same analytical lab and in a different analytical laboratory, as an external control. The samples analysed were from drill chips and drill

core.

In the drilling campaigns during 2010 the methodology was changed, and the pulp duplicates

samples were analysed in the same laboratory where the original samples were analysed.

Diamond Drilling 2010 Pulp Duplicates

Figure 19: Scatterplot comparison between original diamond drillhole fire assay samples and pulp duplicates.

Original FA grade Au (ppm)

Pu

lp D

upli

cate

Au (

ppm

)

R2= 0.9967

Precision= 20.81%


43

Diamond drilling 2010 pulp duplicates has an acceptable precision result in the sample pairs. The

precision value obtained is 20.81%, which is close to the 20% acceptable value described by Abzalov (2008), and it presents a high correlation of data at R

2 = 0.9967.

Reverse Circulation Drilling 2010 Pulp Duplicates

Figure20: Scatterplot comparison between original reverse circulation drillhole fire assay samples and pulp duplicates.

Reverse circulation drilling 2010 pulp duplicates has a low precision result in the sample pairs. The precision value obtained is 75.19%, which is far from the 20% acceptable precision value described by

Abzalov (2008), and it presents a correlation of data at R2 = 0.9241.


Pulp

Dupli

cate

Au

(p

pm

)

R2= 0.9241

Precision= 75.19%


44

Grab samples previous 2010 Pulp Duplicates analysed in the original lab

Figure21: Scatterplot comparison between original grab samples fire assay samples and pulp duplicates analysed in the

original lab.

Grab samples previous 2010 Pulp Duplicates reanalysed in an external control lab

Figure22: Scatterplot comparison between original grab samples fire assay samples and pulp duplicates analysed in an

external control lab.

R2= 0.5176

Precision= 71.01%

Pu

lp D

up

lica

te A

u (

ppm

)


Pu

lp D

up

lica

te A

u (

ppm

)

R2= 0.7893

Precision= 47.01%



45

Trench duplicate samples were analysed in the same analytical laboratory as the original samples, and

they were also analysed in an external control lab. In both cases the precision value obtained is low, greater than the acceptable 20% value described by Abzalov (2008).

The precision value of 47.01% returned for the duplicate is analysed in the original lab and 71.01%

when the duplicate is analysed in an external lab. Also the correlation coefficient squared is presents a value far from 0.9-1.0, which would be the acceptable value (it is 0.7893 when the duplicate is analysed

in the original lab and 0.5176 when the duplicate is analysed in an external lab).

6.10.6.3 Monitoring of Lab Duplicates – Precision

Stewart Kyrgyzstan laboratory has a policy to insert pulp duplicates of some of the samples are

part of its in-house QA/QC process. In the present project these laboratory duplicates have been

study in the same way as the pulp duplicates inserted in each batch of samples.

Diamond Drilling 2010 Lab Duplicates

Figure23: Scatterplot comparison between original diamond drilling drillhole fire assay samples and lab duplicates.

The precision value obtained for the diamond drilling 2010 is 78.86%, which is far from the 20%

acceptable value described by Abzalov (2008), although it presents an high correlation of data at R2 =

0.9790.


Lab

Dupli

cate

Au (

ppm

)

R2= 0.9790

Precision= 78.86%


46

Reverse Circulation Drilling 2010 Lab Duplicates

Figure24: Scatterplot comparison between original reverse circulation drilling drillhole fire assay samples and lab

duplicates.

The precision value obtained for reverse circulation drilling 2010 lab duplicates is 75.19%, which is far

from the 20% acceptable precision value described by Abzalov (2008), and it presents a low correlation

of R2 = 0.9241.

6.10.6.4 Monitoring of Blanks – Accuracy

Blanks are essentially standards with zero grade. Blank sample assay data should be plot on a

control chart as described before for the monitoring of standards.

In the present study CRM samples with a very low grade of 0.003 ppm Au and in-house made

samples of 0.005 ppm Au were used as blank samples, when the detection limit of the analytical laboratory was 0.01 ppm.

In the following graphs, done for plotting blanks samples, it was used a blank value of 0.01 ppm.


Lab

Du

pli

cate

Au

(p

pm

)

R2= 0.9241

Precision= 75.19%


47

Figure25: Graph plot for DDH 2010 blank samples

Figure26: Graph plot for RC 2010 blank samples

In the diamond drilling and reverse circulation campaigns 2010 all the blank samples analysed present a result under the laboratory detection limit, which indicate no contamination in the preparation/analysis

of the samples.

Au

(p

pm

)

Number of samples

Au (

ppm

)

Number of samples


48

6.10.7 CONCLUSIONS OF QAQC STUDY

CRM Standards

The review of external standard values plotted on control charts show results fall within acceptable

limits, indicating analysis is reasonably accurate for Au, however, without the assay certificates

available at this time it is impossible to confirm the standards used are correct. Therefore some caution is needed when interpreting these results.

Duplicates

The diamond drilling 2010 pulp duplicates present an acceptable precision.

Reverse circulation pulp duplicates present a very low precision, the same as diamond drilling and

reverse circulation lab duplicates.

The grab samples taken for previous drilling and exploration campaigns from which there were analysed

pulp duplicate samples, present also a low precision, both in the scenario of duplicate samples analysed

in the original laboratory and in the scenario of duplicate samples analysed in an external laboratory.

Blank Samples

In diamond drilling and reverse circulation all the blank samples are inside the acceptable limit, indicating no contamination in the preparation/analysis of the samples.

6.10.8 RECOMMENDATIONS

Blank and standard samples indicate that the analysis laboratory procedures meet with acceptable and

standard procedures.

Diamond drilling 2010 presents a good accuracy and an acceptable precision.

RC drilling 2010 presents a good accuracy and but a very low precision. This result indicates a problem

in the sampling process, previous to the analysis in the laboratory. A revision to the sampling process

should be done.

There appears very low precision between duplicate samples introduced in the same analysis process by

the laboratory. It is recommended checks performed as to the reason this problem occurs in the analysis

laboratory.

The grab samples taken in the exploration campaigns previous to 2010 only test for precision, and the

result is that they present low precision. This result is quite common for grab samples, which sampling

procedure sometimes doesn’t follow strictly the standard procedures. There are no tests for assay accuracy.

Due to the lack of a proper QAQC procedures in the drilling campaigns prior to 2010 it is recommended Alhambra undertake a program of twinned drillholes and trenching, to verify the samples taken in the

previous campaigns are reliable for use in estimation.


49

6.11 DOMBRALY - MINERAL PROCESSING AND METALLURGICAL

TESTING

Mineral processing and recovery studies were reported in 2005 and 2006 by Kazmechanobr, the

National Centre of Complex Processing of Minerals of the Republic of Kazakhstan (Bolotova, July

2005, December 2005a, December 2005b and September 2006).

Column Leach Test Sample DTL-1

The July 2005 Kazmechanobr report describes column leach test work on technological sample DTL-1

composited from sub-samples OD-2/1, OD-3/1, OD-4/3, c OD-4/4 and OD-7/4, described as the clay

rich oxide portion of the upper level of the Dombraly gold deposit, with an average gold grade reported as 1.52 g/t Au, made up of brecciated, silicified felsic and intermediate volcanics. Drill samples were

not used to prepare this composite sample. ACA Howe can not tell whether the samples came from

surface or underground exposures of the clay rich oxide portion of the upper level of the deposit.

The sample was studied by mineralogical and size analysis, fire assay and chemical analyses, multi-

element semi-quantitative spectral analysis and multi-element quantitative spectral analysis. Gold is the

main valuable component. The average of four fire assays by Kazmechanobr was 1.96 g/t Au.

The oxidised mineralisation has free gold which is potentially amenable to cyanide leaching. Bottle roll

cyanide leaching tests were carried out with and without adsorbent (resin or carbon) and showed that the material is leached by cyanide and does not contain materials with a deleterious effect on gold

adsorption.

The hydrodynamic character of agglomerated oxide was tested under different conditions to determine

the best way to prepare the material for heap leaching.

This column leach test showed that the agglomerated, minus 50 mm oxide mineralisation of the Dombraly gold deposit can be processed using heap leaching with a predicted 74.6% recovery of gold

during commercial scale production. A study of leach tails using 12, 20, 40, 50 and 60 kg of cement per

tonne of oxide, showed that 60 kg/t gave satisfactory results. ACA Howe considers this to be a very high cement requirement.

To achieve such gold recovery, 2.45 m3

of the leaching solution was percolated per tonne of ore. Reagent consumptions were determined as 0.485 kg of sodium cyanide and 0.00012 kg of alkali per

tonne of ore. Washing solutions amounting to 0.490 m3

per tonne of ore will be needed for a complete

washing out of cyanide and alkali.

It is the opinion of ACA Howe that the material of sample DLT-1, while typical, may not be spatially

and statistically representative of the whole of the oxide mineralisation.

Column Leach Test Sample DLT-1 -add

The December 2005a Kazmechanobr report describes column leach test work on technological sample DLT-1 -add which was collected from the dump terrace of the Dombraly II gold deposit. From its gold

grade, physical and mineralogical composition, the sample is reported to be characteristic of the

oxidised mineralisation.

Sample DLT-1 -add is somewhat different from sample DLT-1. Despite the finer grain size of sample

DLT-1 -add of minus 20 mm, the yield of the minus 0.63 mm fraction in DLT-1 -add is 5% lower than that of sample DLT-1. Gold distribution in both samples is reported to be similar.


50

Gold is the main valuable component. The average of four fire assays by Kazmechanobr was 1.52 g/t

Au. Gold is in a form which allows leaching by cyanide solutions.

Tests showed that the optimal use of cement for agglomeration was 38 kg per tonne of ore which seems

high to ACA Howe. A column leach test proved that this type of material from Dombraly is suitable for

heap leach gold production. However, it is noted by ACA Howe that the combined gold recovery on resin and in the stripping solutions during the column leach test, was relatively low at 53.33%. This

indicates to ACA Howe that commercial scale recovery may be even lower; perhaps 3 to 5% lower due

to permeability efficiency and gold losses on the side of heaps. For this gold recovery, 1.31 m3 of

leaching solution per tonne of ore was used. Reagent consumptions were determined as 0.329 kg of

sodium cyanide and 0.118 kg of alkali per tonne of ore. Washing solutions amounting to 0.281 m3 per

tonne of ore will be needed for a complete washing out of cyanide and alkali.

Enlarged bottle roll tests indicated that gold recovery from sample DLT-1 -add significantly depends on

the grain size distribution of crushed ore. These tests have shown that gold recovery from ore crushed to

80% minus 0.074 mm was around 89%. Gold recovery from ore crushed to minus 2 mm was around 73%. Gold recovery from ore crushed to minus 20 mm was around 44%. It is clear that gold recovery

from sample DLT-1 -add is poor by comparison with the results of sample DLT-1.

Good gold recovery from sample DLT-1 -add depends for the most part on fine grain size following

crushing. It is clear that studies of crushing and agglomeration and leaching will be required to optimise

gold recovery from this type of material. ACA Howe suggests that this ore type may be more suitable for processing by conventional CIL technology (Carbon in Leach).

It is the opinion of ACA Howe that the Dombraly dump terrace material represented by sample DLT-1 -

add , while described as characteristic, is not spatially and statistically representative of the whole of the oxide mineralisation and that it has different mineralogical and relatively poorer metallurgical

characteristics compared with sample DLT-1 material which may be from undisturbed oxide material.

Petrographic and chemical analysis and bottle roll cyanidation of nine core samples

The December 2005b Kazmechanobr report describes the results of test work on nine small core samples from oxide, transitional and sulphide mineralisation zones. The sample numbers are DTK-1 to

DTK-3, DTK-5 to DTK-7 and DTK-9 to DTK-11, which were collected from the tailings of

processing core samples. The sampling intervals range from 25 to 327 m downhole and the weights of samples range from 0.29 to 1.92 kg. Gold grade in the samples range from 1.54 to 23.2 g/t Au.

The purpose of this work was to create a technological framework based on the results of chemical analysis and bottle roll cyanide leaching tests.

The physical composition of the samples was studied and fire assay and chemical analyses were carried

out. Gold grades range from 0.72 g/t Au to 17.50 g/t Au. Contents of copper, zinc and lead are very low. All samples contain arsenic, from 0.05 to 0.74 % As. The sulphide content increases with depth.

Standard bottle roll tests were conducted, including direct cyanidation and adsorption leaching with resin. Cyanide leaching showed high gold recovery rates from oxide and transitional samples of 94 to

98 %. Gold grade in leaching residues ranges from 0.02 to 0.34 g/t Au.

Gold recovery from sulphide samples by direct cyanidation was considerably lower at 29 to 83%, with

gold grades in the tailings of 1.36 – 4.34 g/t Au. Adsorption leaching increases gold recovery rates by 4

to 47% for different samples. This fact indicates that these sulphide ores have a “preg-robbing” effect,

i.e. the ore contains active natural adsorbent such as organic carbon which adsorbs gold leached into solution.


51

A decision was made to conduct an additional analysis of the samples in order to determine the amount

of organic carbon (as opposed to carbon in carbonate), which was found to range from 0.02 to 0.73% C.

Content of other metals in the samples is not high, and they do not impede the process of leaching and

adsorption of gold.

From its composition and leach performance, sample DTK-9, collected from the interval 25 - 30 m in

Drill Hole C 161, should be called low sulphidation mineralisation. Gold recovery by cyanide leaching

was 63.5%.

It is the opinion of ACA Howe that these nine core samples are not spatially and statistically

representative of the whole of the oxide, transitional and sulphide mineralisation and the results of this study, alone, can not be used to map the boundaries between these types of mineralisation.

Metallurgical test work on sulphide mineralisation sample DLT-2

The September 2006 Kazmechanobr report describes gravity, froth flotation, cyanidation and adsorption

leaching test work on an 85.6 kg sample of sulphide mineralisation combined from five core drill holes at Dombraly: С 162, С 322, С 323, С 402 and С 403, from downhole depths of 127 to 348 m.

Characteristic sulphidic gold mineralisation samples were combined from five drill core intercepts of 3 m to 52 m in length, which produced weights ranging from 3.1 to 43.9 kg with a simple combined total

weight of 85.6 kg in sample DLT-2. The grades of the five intercepts were 3.46, 3.62, 4.63, 5.06 and

13.06 g/t Au. The average gold content is reported as 7.19 g/t Au but it is not clear how this was

derived.

The sample rock types are described as carbonaceous sedimentary and tuffaceous sedimentary rocks

with dispersed disseminated pyrite.

The sample was crushed to minus 2 mm and only two fire assays, for gold and silver, of the combined

sample were done and gave 7.08 and 3.40 g/t Au and 5.24 and 10.16 g/t Ag, indicative of the presence

of coarse particulate gold. These assays were simply averaged to 5.24 g/t Au and 7.7 g/t Ag. Despite these reported average grades, elsewhere in the report the average gold and silver grade is reported as

6.5 g/t Au and 7.7 g/t. Chemical analysis showed total sulphur as 1.74 % S and sulphide sulphur as

1.71% S. Detrimental impurities include arsenic and carbon. The results of chemical analysis indicate a complex composition.

Gravity separation tests on shaking tables recovered 53.45% of the head grade gold into concentrate with a 3.53% yield and a concentrate grade of 61.5 g/t Au.

Froth flotation tests recovered 86.13% of the gold into concentrate with a 13.59% yield and a

concentrate grade of 27.4 g/t Au.

Direct cyanidation leach testing recovered only 17.65% of gold into cyanide solution due to the

presence of carbonaceous shale in the ore, which acts as a natural adsorbent of gold.

Cyanidation leach tests with resin adsorbent recovered 79.23% of the gold to the resin and 64.81% of

the gold was recovered from the loaded resin in gravity tails.

It is the opinion of ACA Howe that the gold ore represented by sample DLT-2 could be successfully

processed by two schemes using:

1. Gravity separation and froth flotation or 2. Gravity separation and conventional carbon- or resin-in-leach technology

and that further work is required to determine which route is economically most beneficial.


52

It is the opinion of ACA Howe that the Dombraly sulphide ore material represented by sample DLT-2,

while described as characteristic, is probably not spatially and statistically representative of the whole of the sulphide mineralisation. However, a wide range of metallurgical recovery characteristics has been

illustrated, which is a useful guide for future studies.

6.12 MCS OCTOBER-DECEMBER 2011 MINERAL RESOURCE ESTIMATES

Following positive results from drilling and in-house geological modelling work the decision was taken

to undertake block model estimations for the Dombraly low grade stockpile, pit infill and in-situ Au mineralised zones to meet with NI43-101 and JORC reporting requirements.

The following sections describe the process and decision making employed for the October-November 2011 resource estimations.

6.12.1 SOFTWARE USED

The Dombraly resource estimates were prepared using MICROMINE version 2011 3D modelling

software and Microsoft Excel 2007.

6.12.2 INPUT DATA SUMMARY

Micromine Consulting Services were provided with Dombraly zones drill, trench and pit data in

Microsoft Excel and Micromine file formats. Existing historical interpretations as polygons, wireframes

and digital terrain models were provided in MICROMINE string, wireframe and .DXF format.

Raw data used in interpretation and modelling consists of data from recent and historical

diamond drilling and RC drilling, low grade stockpile trench and RAB sampling exploration

work undertaken by Alhambra and previous explorers.

Raw data used as input to estimation consists of recent 2010 diamond, RC drill data and 2005 low

grade stockpile trench data.

It is Micromine’s and ACA Howe’s opinion that available input data is suitable for use as part of a NI

43-101 compliant and reportable resource estimations.

A summary of input sample type for the Dombraly deposit models are presented in Table 13 below.

TABLE 13. DOMBRALY MARCH 2011 RESOURCE ESTIMATE SAMPLE

SUMMARY

Deposit Area Lode/Zone Sample

Type

No of

Holes

No of

Samples

Sample

Length

Dombraly Low grade

stockpile RC/Trench 25/10 340/351 1m/5m

Dombraly Pit Infill RC 7 74 1m

Dombraly In-Situ

mineralisation RC/DDH 12/13 466/2877 Av. 1m

6.12.3 INPUT DATA

Data selected for use in the November 2011 block model estimations is contained in Micromine

drillhole databases. Each drillhole database comprises collar, lithology, survey and assay data files.

Each trench drillhole database comprises collar, survey and assay files.


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Input data file listing is provided in Appendix 2 and applicable data files and wireframes are provided in

the database which accompanies this report.

6.12.4 DATA VALIDATION

Drill hole collar, assay, survey and lithology data were processed as MICROMINE .dat files.

Data validation cross referencing collar, assay and litho (geology) files was performed in MICROMINE

to confirm drill hole depths, identify any inconsistent or missing sample/logging intervals and survey

data.

Channel collar, assay and survey were also processed as MICROMINE .dat files and validation performed to confirm hole depths between files, inconsistent or missing sample intervals and surveys.

No fatal errors were detected during computerised and visual data validation.

Data validation tables are presented in Appendix 2.

6.12.5 DESCRIPTIVE AND CLASSICAL STATISTICS

One of the most important parts of resource estimation is the study of the database, as its distribution has a fundamental influence in the estimation process. For the present study there have been taken into

account the samples from the drilling programs, and the samples from the trench sampling program. In

total 691, 74 and 3343 samples have been studied for the Dombraly low grade stockpile, pit infill and

in-situ zones respectively. The elements studied here are Au, those elements which were used to create the geological modelling and in estimation.

A Summary of the unrestricted raw Au data basic parameters for each deposit (low grade stockpile, pit

infill and in-situ mineralised zones) is presented in Table 14 below.

TABLE 14. SUMMARY OF BASIC PARAMETERS FOR DOMBRALY AU

INPUT DATASET

Deposit Number of

samples

Min

(ppm)

Max

(ppm)

Mean

(ppm) Median COV

Standard

Deviation

Low

grade

stockpile

691 0.012 48.7 0.97 0.40 2.498 2.422

Pit Infill 74 0.027 11.2 1.062 0.646 1.608 1.708

In Situ 3343 0 40.5 0.093 0.093 8.827 0.816

6.12.6 AU DISTRIBUTION

Au datasets for each deposit zone shows a strongly positive skewed distribution, with a great number of

samples with low to very low grade, and few samples with high grade. Au does not follow a Gaussian

Bell shape distribution. For estimation purposes each database’ distribution should present a Gaussian Bell shape (as would be in a ‘perfect’ distributed database). In order to conduct a study and statistical

interpretation of the dataset, and trying to achieve the Gaussian Bell shape, it was necessary to plot

element histograms using log-normal transformation. This is very common practice for gold deposit statistical analysis and estimation.

Histograms for the raw Au data for the 3 deposit zones are presented in Figures 27 to 29 below.


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Figure 27: Log normal histogram distribution for Low Grade Stockpile Au dataset

Figure 28: Log normal histogram distribution for Pit Infill Au dataset


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Figure 29: Log normal histogram distribution for In-situ Au dataset

From the histograms above it can be seen that the log-normal distribution of each deposit zone dataset

better represents a Gaussian Bell shape, although not perfect. This suggests that within the datasets

there are different grade populations. For estimation purposes the different grade populations (in each element) need to be separated as best possible for reliable robust interpolations. A first approximation

for getting the dataset divided into different populations is achieved through domain interpretation and

modelling.

6.12.7 NATURAL CUT OFF

The data for each element was carefully studied looking for the existence of natural boundaries or cut-

offs between grade populations. These natural cut-offs would indicate limits in the dataset between ‘mineralised’ and ‘non-mineralised’ samples.

Log Histograms generated for unrestricted data show grades to have a number of populations. From the

log normal histograms it is observed that there is not a clear limit between the mineralised and the non-mineralised samples. Following careful review and interpretation limits were recognised where a

natural cut off could be implied:

Low Grade Stockpile = 0.13g/t Au

Pit Infill = 0.1g/t Au

In-Situ = 0.2g/t Au

These grade values are considered as a natural boundary to mineralisation used for mineralised domain

modelling of the deposit.

In addition to statistical support of natural grade boundaries, consideration was given to the deposit

types being modelled, typical mining widths, anticipated mining methods and overall grade in selecting appropriate cut off values.


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6.12.8 DOMAIN INTERPRETATION AND MODELLING

Domain interpretation was completed from cross section and plan displays in Micromine software.

Interpretation of lode orientations done using information from discussions with Alhambra and ACA

Howe personnel on geological and deposit models, review of literature, apparent continuity and correlation displayed by exploration data and also guided by existing wireframe 3d solid models of

historical lodes, previously mined areas (historical pit shell and recent surveyed pit DTM’s) supplied to

MCS.

The interpreted profiles represent a 3 dimensional polygonal display of the current geological and

exploration model, based on the recent drill and channel data for the zones, application of knowledge

from previously mined areas, mapped and inferred structural architectures.

The boundaries to mineralised zone digitised in cross section were interpreted using best practice industry standard techniques, snapping to drillhole assay intervals, and utilising lithology where

applicable to improve accuracy of location of mineralised zone in 3 dimensions, and to reduce the

inclusion of waste within the mineralised wireframes.

Grade domain modelling was completed for Au, with minimum width of approximately 1m. Where

appropriate, geology was used in combination with grade values to assist zone interpretation.

The following grade domain models were generated across the 3 recognised deposit zones:

Low Grade Stockpile: 2 domains interpreted using an Au cut off value of 0.13g/t Au.

Pit Infill: 1 domain interpreted using an Au cut off value of 0.1g/t Au.

In-Situ: 11 domains were interpreted using an Au cut off value of 0.2g/t Au.

Other criteria for the grade domain models are:

Where constrained, lode mineralisation is extended half the distance between drillhole

and trench samples, along both strike and dip direction.

Where unconstrained, zones were extrapolated approximately 30 to 50 m along strike

and approximately 30 to 50 m in the dip direction from sample control points depending

upon grade, thickness and confidence in extrapolation of the model.

Zones were extended approximately 1 m in a thickness direction where unconstrained

i.e. where drillholes and/or channels started or finished in mineralisation.

The types, number and characteristics of the models will be subject to change as new data becomes

available and revisions of the geological model take place.

Once completed wireframes were validated both by visual examination and computer validated by use of the wireframe validation tools in Micromine.

Micromine wireframe validation process checks for open solids and any intersection triangles and

strings which may cause problems in volume and tonnage calculations.

Each mineralised domain wireframe was checked and deemed valid.

The Au mineralised domains modelled as part of this resource estimation study are:


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Low Grade Stockpile (Upper/Lower)

Pit Infill material

In situ near surface Oxide Au domain

In situ domain structure A

In situ domain structure B

In situ domain structure C

In situ domain structure D

In situ domain structure E

In situ domain structure F

In situ domain structure G

In situ domain structure X

In situ domain structure Y

In situ domain structure Z

The figures below present plan and 3d view representations of the mineralised domain models and exploration data utilised for interpretation. (Figures 30 - 36)

FIGURE 30: PLAN VIEW OF THE LOW GRADE STOCKPILE AND PIT INFILL Au

DOMAIN MODELS

cboyd

Typewritten Text

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FIGURE 31: 3D VIEW OF THE LOW GRADE STOCKPILE AND PIT INFILL Au

DOMAIN MODEL

cboyd

Typewritten Text

59

FIGURE 32: PLAN VIEW OF THE IN SITU Au DOMAIN MODELS (COLOURS)

cboyd

Typewritten Text

60

FIGURE 33: 3D VIEW OF THE IN SITU Au DOMAIN MODELS (COLOURS)

cboyd

Typewritten Text

61

FIGURE 34: CROSS SECTIONAL VIEW OF LOW GRADE STOCKPILE Au

DOMAIN MODEL

cboyd

Typewritten Text

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FIGURE 35: CROSS SECTIONAL VIEW OF PIT INFILL Au DOMAIN MODELS

cboyd

Typewritten Text

63

FIGURE 36: CROSS SECTIONAL VIEW OF IN SITU Au DOMAIN MODELS

cboyd

Typewritten Text

64


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6.12.9 DOMAIN STATISTICS

Mineralised domain statistical study has been undertaken on raw sample assay values, all lying within

the different domains modelled and described in the previous section of the report.

The summary statistics for each domain an element are presented in Table 15 below.

TABLE 15. DOMBRALY MINERALISED DOMAIN RAW DATA DESCRIPTIVE

STATISTICS

Domain Element

Number

of

samples

Min

(ppm)

Max

(ppm)

Mean

(ppm)

C.O.V

Low Grade Stockpile

Upper Au 161 0.019 12.1 1.196

1.638

Low Grade Stockpile

Lower Au 381 0.013 48.7 1.198

2.446

Pit Infill Au 70 0.133 11.2 1.12 1.553

In Situ Oxide Au 61 0.088 4.04 0.759 0.984

In Situ A Au 20 0.05 40.5 3.028 2.955

In Situ B Au 24 0.033 4.78 1.273 1.226

In Situ C Au 5 0.048 4.59 1.136 1.708

In Situ D Au 5 0.005 1.26 0.64 0.782

In Situ E Au 9 0.247 2.08 0.738 0.86

In Situ F Au 21 0.005 10.3 1.347 1.803

In Situ G Au 3 0.474 0.815 0.675 0.264

In Situ X Au 63 0.076 4.04 0.809 1.057

In Situ Y Au 9 0.143 5.52 0.934 1.858

In Situ Z Au 13 0.05 7.8 1.282 1.728

6.12.10 TOP CUTS

Top cut analysis was performed on mineralised domain raw data for Au prior to block model estimation. Top cut analysis is undertaken to assess the influence extreme grade outliers has on the sample

population of each domain. Whilst extreme grades are real, their influence in interpolation may

overstate the block grades in some parts of the deposits. Micromine distribution graphs and ranked assay data were prepared and analysed to examine the domain samples and effects of a range of top-cuts

applied to raw data and the effect these have on the co-efficient of variation (COV) and loss of data

from the domain.

For the Au element studied via classical statistics, domain histograms and domain statistics indicate reasonably well distributed log normal data and coefficient of variation close to 1. A number of

mineralised domains display minor amount of unusually high grade outliers which could affect the

grade estimation.

Following statistical analysis of domain input Au sample data, top cuts of 10 g/t Au were applied to the

low grade stockpile (lower), and in situ mineralised domain A. Top cuts of 5 g/t Au and 6 g/t Au were applied to pit infill domain and in-situ mineralised domain F respectively.

Descriptive statistics and histograms for determination of top cuts are presented in Appendix 3.


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6.12.11 COMPOSITES

Prior to estimation, samples within the mineralised wireframes contained in the database assay files were composited to a standard length for geostatistical analysis and interpolation. The

decision about composite length was determined by considering the histogram for raw sample

intervals and selecting the dominant length.

By considering the histograms of sample intervals (Figures 37 and 38) the average sample length

of 5 m and 1 m was taken to be the composite length for the low grade stockpile, and pit infill/in-

situ domains respectively. A composite assay file was created for samples within the different domains wireframes to be used in block model interpolation.

Figure 37: Histogram of sample intervals – low grade stockpile

Figure38: Histogram of sample intervals – In-Situ


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6.12.12 GEOSTATISTICS

6.12.12.1 Domain Statistics

Mineralised domain statistical study has been undertaken on raw composited sample assay values, all

lying within the different domains modelled previously. The summary statistics for each domain an element are presented in Table 16 below.

TABLE 16. DOMBRALY MINERALISED DOMAIN COMPOSITE DATA

DESCRIPTIVE STATISTICS

Domain Element

Number

of

samples

Min

(ppm)

Max

(ppm)

Mean

(ppm)

C.O.V

Low Grade Stockpile

Upper Au 131 0.1 10 1.36025

1.492

Low Grade Stockpile

Lower Au 253 0.0324 10 1.12707

1.22

Pit Infill Au 70 0.133 5 0.96599 0.96599

In Situ Oxide Au 62 0.088 4.04 0.784 0.978

In Situ A Au 12 0.05 10 1.84933 1.734

In Situ B Au 29 0.033 4.78 1.176 1.226

In Situ C Au 6 0.116 4.59 1.021 1.72

In Situ D Au 5 0.005 1.26 0.64 0.782

In Situ E Au 10 0.247 1.637 0.631 0.816

In Situ F Au 21 0.005 6 1.28695 1.36

In Situ G Au 3 0.474 0.815 0.675 0.264

In Situ X Au 46 0.076 4.04 0.7 1.262

In Situ Y Au 14 0.143 5.52 0.883 1.639

In Situ Z Au 17 0.05 7.8 1.245 1.589

6.12.12.2 Variography

Micromine Consulting conducted investigations into the production of meaningful semivariograms for

use in a linear geostatistical interpolation method (ordinary kriged), however, due to the few samples

available per domain, it was impossible to produce structured semi-variograms (it is commonly necessary to have a dataset of, at least, thousands of sample pairs per domain to undertake a variography

study).

Therefore, Micromine Consulting has not produced any variography study for grade interpolation, and used a non-geostatistical linear grade interpolation methodology. The interpolation method selected was

Inverse Distance Weighting (IDW).

6.12.13 MCS OCTOBER-DECEMBER 2011 IDW BLOCK MODEL ESTIMATION

6.12.13.1 Empty Cell Block Model

Domain restricted empty cell block models were created for each Dombraly deposit zone using

definitions which cover the extent of mineralised domains.


68

A parent block size of 5m × 10m × 5m was chosen, sub-blocking to minimum sub-block size of 1m x

1m x 1m for the waste dump. A parent block size of 5m × 5m × 5m was chosen, sub-blocking to minimum sub-block size of 1m x 1m x 1m for pit infill domain and in-situ domains.

This block size was chosen after considering the geological model, exploration grid, search ellipsoid

ranges, composite size, SMU and potential future mining methods.

6.12.13.2 Grade Interpolation

Prior to grade interpolation, domains were checked for missing intervals. Resource estimation best practice dictates missing sample intervals are assigned a value of zero for grade interpolation. For the

Dombraly mineralised domain models, no missing intervals were detected.

Mineral grade was interpolated into the block models on a zone and element domain basis.

For interpolation both the block model and composite assay file was flagged by the mineral and zone

domains and blocks within these domains assigned an interpolated grade (i.e. wireframe restricted or

closed interpolation).

Metal grade interpolation was then undertaken using the input composite assay files for all areas which

contained composite length drill and channel data.

Top cuts were applied to the input data according to statistical analysis performed and described in

section 14.10.

For each domain, the parent block Inverse Distance Weighted interpolation technique was used and

interpolation performed at different search radii, until all blocks within each domain had received an

interpolated grade. The search distances were determined by means of the evaluation of drill/trench spacing (input sample spacing), and orientations through study of geological models and domain

geometries.

Inverse distance weighting (IDW³) method of interpolation was used, which is a non-geostatistical (classical) method of grade interpolation. In this method, each input sample is weighted according to

some power of the inverse distance from the block to be estimated. Interpolation weights are only

applied to samples found within the search neighbourhood. There are no strict rules for choosing a

power; for gold a value of two or three is often used, with three most common. For iron a power of two may be appropriate. The lower the power, the more the grades are smoothed, to the point where using a

low power will produce a result which deviates only slightly from the global mean of the data. On the

other hand, higher powers will produce a result that approaches a nearest-neighbour interpolation, with the sample nearest the block contributing almost all of the weight.


69

A power of 3 was chosen for this interpolation given the deposit, grades and commodity type.

Search ellipsoid range (extent) parameters were based upon model geometries and sample density.

In the pit infill domain grade interpolation was constrained by bench. Bench height is 8 m.

Due to the extreme low number of samples and narrow nature of the transitional zone, samples for the

primary and transitional zones were combined for interpolation to inform the two zones.

The transitional and primary material zones split for the application of bulk density values and reporting purposes.

Table 17 below summarises the search ellipsoid parameters used for each Dombraly mineralised zone.

TABLE 17. DOMBRALY IDW SEARCH ELLIPSOID PARAMETERS

Domain Direction Azi (°) Dip (°) Range (m)

Low Grade

Stockpile Upper

First 035 0 50

Second 125 0 25

Third - 90 4

Low Grade

Stockpile Lower

First 035 0 50

Second 125 0 25

Third - 90 4

Pit Infill

First 000 0 75

Second 090 0 75

Third - 90 5

In Situ Oxide

First 000 0 150

Second 90 0 150

Third - 90 2

In Situ A

First 000 0 225

Second 90 35 112.5

Third 90 55 2

In Situ B

First 000 0 240

Second 090 40 240

Third 090 50 2

In Situ C

First 000 0 240

Second 090 40 240

Third 090 50 2

In Situ D

First 000 0 240

Second 090 55 240

Third 090 35 2

In Situ E

First 000 0 240

Second 090 36 120

Third 090 54 2

In Situ F

First 000 0 240

Second 090 50 240

Third 090 40 2

In Situ G

First 000 0 240

Second 090 52 240

Third 090 38 2

In Situ X

First 000 0 240

Second 090 36 240

Third 090 54 2

First 000 0 240


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In Situ Y Second 090 32.5 240

Third 090 57.5 2

In Situ Z

First 000 0 240

Second 090 33 240

Third 090 57 2

The first search radii for interpolation were selected to be equal to two thirds of the range in the

strike, dip and across dip directions of the search ellipsoid. Model blocks that did not receive a grade estimate from the first interpolation run were used in the next interpolation run, equal to the

full ranges in all directions. Subsequent search radii were incremented by multiples of the initial

ranges in corresponding directions.

When model cells were estimated using radii not exceeding the full ranges (i.e. two thirds and equal to the ranges), a restriction of at least three samples from at least two drill holes was applied

to increase the reliability of the estimates.

Detailed definition of the interpolation parameters is contained in Table 18 below.

TABLE 18. DOMBRALY INTERPOLATION PARAMETERS

Interpolation Method Inverse Distance Weighted (IDW³)

Interpolation Run

Number

1 2 >2

Search Radii 2/3 range in main directions

Equal to the range in main directions

Greater than the range in main directions

Min no of Samples 3 3 1

Max number of Samples 16 16 16

Min no of Drill holes 2 2 1

Block Discretisation 5x5x5 5x5x5 5x5x5

6.12.14 BLOCK MODEL ATTRIBUTES

Once the interpolation process and domain flagging for the orebody block model (OBM) was

complete, the resultant final block model files contain a series of attributes for each block as

outlined in Tables 19 to 21 below.

TABLE 19. LOW GRADE STOCKPILE BLOCK MODEL ATTRIBUTES

MM LOW GRADE STOCKPILE OBM OCT 2011 FINAL

Attribute Description Block Model Field

Au grade Au grade (cut) Au Fa_10

Domain Low Grade Stockpile Au

Domain

WFM (Low Grade Stockpile Au)

Domain Level DUMP (upper. Lower)

Run Interpolation Pass RUN (1, 2, 3….)

Bulk Density Density SG


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TABLE 20. PIT INFILL BLOCK MODEL ATTRIBUTES

MM PIT INFILL OBM OCT 2011 FINAL


Au grade Au grade (cut) Au Fa_5

Domain Pit Infill Au WFM (Pit Infill)

Domain Bench level Bench (1-4)



TABLE 21. IN SITU BLOCK MODEL ATTRIBUTES

MM IN SITU OBM OCT 2011 FINAL


Au grade Au grade (cut) Au Fa

Domain In situ Au Domain WFM PR (A-G, X-Z)

Domain In situ Au Domain WFM OX (OX)



Bulk density (SG) assigned to the block model cells were taken from recent core measurements as

described in section 14.6 and summarised in Table 22 below.

TABLE 22. BASIC SUMMARY STATISTICS FOR DOMBRALY BULK DENSITY

SAMPLES

Domain

No. of

sample

s

Oxide Transitio

nal Primary Comment

Low Grade

Stockpile

Pit Infill

In Situ Oxide 0 No data No Data

No Data 2.65 Average of all SG domain samples

In Situ A 13 No Data No data 2.6 Use 2.6

In Situ B 2 No Data No data 2.55 Use 2.55

In Situ C 8 No Data No data 2.53 Use 2.53

In Situ D 7 No Data No data 2.62 Use 2.62

In Situ E 1 No Data No data 2.49 Use 2.49

In Situ F 6 No Data No data 2.86 Use 2.86

In Situ G 0 No Data No data 2.65

2.65 Average of all SG domain samples


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In Situ X 0 No Data No data 2.65


In Situ Y 0 No Data No data 2.65


In Situ Z 0 No Data No data 2.65


For those domains where no density measurements were taken, the mean of all 24 sample densities was

used (2.65).

6.12.15 RESOURCE CLASSIFICATION

Classification methodology, or assigning a level of confidence to mineral resources at Dombraly have been undertaken in adherence to the Australasian Code for Reporting of Exploration Results, Mineral

Resources and Ore Reserves (JORC Code, 2004 edition) and follows the Micromine Consulting

Resource Modelling Standard Procedures (2010), and conforms to the CIM Mineral Resource and Mineral Reserve definitions referred to in National Instrument 43-101 and the Standards of Disclosure

for Mineral Projects.

Classification of interpolated blocks is undertaken using the following criteria;

Interpolation criteria based on sample density, search and interpolation parameters

Assessment of the reliability of geological, sample, survey and bulk density data

Robustness of the geological model

Deposit type

Drilling and sample density

Understanding of grade continuity

Thorough consideration of the above factors, MCS is of the opinion that a small portion of indicated

blocks occur within the low grade stockpile resource, with inferred category an appropriate resource

classification for pit infill and in situ mineralisation at this time. There are a number of minor issues relating to data verification and QAQC which need to be addressed to increase confidence of input data,

particularly Sn prior to increased resource class. Sample spacing and drillhole density within domains

is wide and the total number of assays in a number of domains remains low, along with bulk density measurements for greater accuracy of tonnage calculations.

6.12.16 MODEL VALIDATION

Global and local model validation was undertaken on the Dombraly block models prior to resource

reporting.

6.12.16.1 Global Validation

A comparison of mineralised domain raw, composite and block grade was undertaken and is outlined in

Table 23 below.


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TABLE 23. COMPARISON OF DOMAIN RAW, COMP AND

BLOCK MODEL MEANS

Domain Raw

Mean

Top Cut

Mean

Composite

Mean

Block

Mean

Low Grade Stockpile

Upper 1.20 1.07 1.13 0.92

Low Grade Stockpile

Lower 1.20 N/A 1.36 1.20

Pit Infill 1.12 0.97 0.97 0.83

In Situ Oxide 0.76 0.78 1.02

In Situ A 3.03 1.50 1.85 1.27

In Situ B 1.27 N/A 1.18 1.17

In Situ C 1.14 1.02 1.73

In Situ D 0.64 0.64 0.86

In Situ E 0.74 0.63 0.55

In Situ F 1.35 1.14 1.29 1.05

In Situ G 0.68 0.68 0.69

In Situ X 0.81 0.70 0.59

In Situ Y 0.93 0.88 0.83

In Situ Z 1.28 1.25 0.97

Comparison of mean grade is on the whole considered satisfactory for the dataset and classification

level.

Model validation also involved the cross reference of block model volume against wireframe volumes. Comparison is made between the wireframe volumes and wireframe flagged block model volume prior

to constraining by the topographical DTM. Results are presented in Table 24 below.

TABLE 24. COMPARISON OF DOMAIN WIREFRAME AND BLOCK

MODEL VOLUMES

Domain WFM Volume Block Volume % Difference

Low Grade Stockpile 910,015 865,996 4.84

Pit Infill 586,352 640,871 8.50

In Situ Oxide 118,709 118,860 0.13

In Situ A 371,327 375,892 1.21

In Situ B 248,875 249,158 0.11

In Situ C 182,874 182,898 0.01

In Situ D 380,640 380,074 0.15

In Situ E 230,786 230,614 0.07

In Situ F 573,960 574,062 0.02

In Situ G 118,215 118,270 0.05

In Situ X 345,396 347,430 0.59

In Situ Y 84,728 84,594 0.16

In Situ Z 150,052 150,876 0.55

MCS is satisfied with the global validations for the Dombraly models.


74

6.12.16.2 Local Validation

Once modelling was completed, a series of sectional slices through the block models was undertaken to assess whether block grades honour the general sense of composite drill hole grades, that is to say that

high grade blocks are location around high sample grades, and vice-versa.

A degree of smoothing is evident in block grade which is to be expected but on the whole block grades correlate well with sample grades. Local validation cross sections are presented in Appendix 4.

6.12.17 NOVEMBER 2011 IDW RESOURCE ESTIMATE REPORTING


according to best practice geological modelling techniques, current understanding of the geological

model, historical interpretations and discussion with Alhambra personnel as described in previous

sections.


dependent upon the cost of mining and processing the mineralisation and the selling price of the final

product.

The economic cut-off grade for Dombraly established using grade and block revenue factors.

Due to the early stage status of the development of the Dombraly deposits, a number of assumptions

have been made with regard to inputs to the calculation of the economic cut-off grade for reporting.

For a single product the calculation is relatively straightforward. Resources are reported using an

economic marginal cut off, determined by use of simple block revenue factor methodology and one year

trailing average gold input price.

6.12.17.1 ECONOMIC CUT OFF DETERMINATION



Inputs for oxide material are based upon actual mining cost data from Alhambra’s nearby Uzboy open

pit operation, and estimated costs for transitional and primary material taken from recent PEA studies undertaken on the nearby Uzboy deposit.







Oxide mining cost – US$1/t (waste dump and pit infill)/US$1.7/t (in-situ)



Cut off calculation is presented below for reference:

Block revenue calculation – Au grade grams per tonne x Recovery x Input gold price per gram


75

Using the Au block grade, the above Au metal price and recovery, MCS estimated the revenue per

mined block.

For a mineralised block to be considered economic it must generate higher revenue than it costs to mine.

For a block to be considered economic it must therefore generate greater than US$4.85/t of revenue for

the low grade stockpile and pit infill material, and US$5.55/t and US$8.42/t, for in-situ oxide and transitional/primary material respectively.

MCS used Micromine software to filter those blocks in the resource model with value greater than the calculated cost to mine values for economic cut-off grade determination and resource reporting.

Cut off grades used for reporting are 0.1g/t Au for the low grade stockpile, 0.2g.t Au for the pit infill

zone and 0.1g/t for in-situ oxide material, and 0.2g/t Au for in-situ transitional and primary material types.

It is MCS’ opinion that the assumptions made for input to economic cut-off grade determination and

reporting of potentially economic resources are reasonable given the current understanding of the

geology, mineralisation, anticipated mining and processing methods and comparison with similar type operations.




A summary of in situ classified inferred resources as of November 2011 for the Dombraly Deposits are

presented in Table 25 below.

TABLE 25. DOMBRALY NOVEMBER 2011 TOTAL RESOURCES

Dombraly Low Grade Stockpile Total Resource by Category and Material Type



t/m3 x 1000 m3 x 1000 t g/t g Oz

Oxide

Indicated 1.67 284 473 1.26 597,000 19,000

0.10g/t Inferred 1.67 578 963 1.07 1,033,000 33,000

Dombraly Pit Infill Total Resource by Category and Material Type



t/m3 x 1000 m3 x 1000 t g/t g Oz

Oxide

Indicated 1.73 50 86 0.97 83,000 3,000

0.20g/t Inferred 1.73 525 908 0.82 747,000 24,000

Dombraly In Situ Total Resource by Category and Material Type



t/m3 x 1000 m3 x 1000 t g/t g Oz

0.1g/t Oxide Inferred 2.63 1,043 2,700 0.99 2,700,000 87,000

0.2g/t Transitional Inferred 2.61 249 646 1.16 750,000 24,000

0.2g/t Primary Inferred 2.71 1,364 3,671 1.12 4,099,000 132,000

0.1g/t Total Inferred 2.64 2,807 7,446 1.02 7,601,000 244,000 ¹ Cut off value used here represents economic cut off determined from block revenue factor calculation methodology and input gold price of

US$1,394/Oz.

²Class represents resource category under CIM and JORC reporting guidelines.

³ Top cuts of 10g/t Au and 6g/t Au have been applied to in situ domains A and F gold assay data respectively. Top cuts of 10g/t Au and 5g/t Au

applied to low grade stockpile (lower), and pit infill domains respectively.


76

A full breakdown of resources reported by zone, are provided in tables accompanying this document as Appendix 5.

Dombraly IDW block model views are presented as Figures 39 – 48 below.

FIGURE 39: PLAN VIEW OF DOMBRALY LOW GRADE STOCKPILE BLOCK MODEL -

Au GRADE DISPLAY

cboyd

Typewritten Text

77

FIGURE 40: PLAN VIEW OF DOMBRALY PIT INFILL BLOCK MODEL - Au GRADE DISPLAY

cboyd

Typewritten Text

78

FIGURE 41: PLAN VIEW OF DOMBRALY IN SITU BLOCK MODELS - Au GRADE DISPLAY

cboyd

Typewritten Text

cboyd

Typewritten Text

cboyd

Typewritten Text

cboyd

Typewritten Text

79

FIGURE 42: PLAN VIEW OF DOMBRALY BLOCK MODELS - Au GRADE DISPLAY

cboyd

Typewritten Text

80

FIGURE 43: 3D VIEW LOOKING NE OF DOMRALY LOW GRADE STOCKPILE AND PIT

INFILL BLOCK MODELS - Au GRADE DISPLAY

cboyd

Typewritten Text

81

FIGURE 44: 3D VIEW LOOKING NE OF DOMBRALY IN SITU BLOCK MODELS -

Au GRADE DISPLAY

cboyd

Typewritten Text

82

FIGURE 45: 3D VIEW LOOKING NE OF DOMBRALY BLOCK MODELS -

Au GRADE DISPLAY

cboyd

Typewritten Text

83

FIGURE 46: 3D VIEW LOOKING NW OF DOMBRALY LOW GRADE STOCKPILE AND PIT

INFILL BLOCK MODELS - Au GRADE DISPLAY

cboyd

Typewritten Text

84

FIGURE 47: 3D VIEW LOOKING NW OF DOMBRALY IN SITU BLOCK MODELS -

Au GRADE DISPLAY

cboyd

Typewritten Text

85

FIGURE 48: 3D VIEW LOOKING NW OF DOMBRALY BLOCK MODELS -

Au GRADE DISPLAY

cboyd

Typewritten Text

86


87

6.13 DOMBRALY - ADJACENT PROPERTIES

Adjacent properties are described above in the general section on this topic.

6.14 DOMBRALY - OTHER RELEVANT DATA AND INFORMATION

None.

6.15 DOMBRALY - INTERPRETATION AND CONCLUSIONS

NI 43-101 compliant computerised 3 dimensional resource estimations for the Dombraly gold project, Akmola Oblast, Kazakhstan were undertaken between October 2011 and November 2011. The study

was undertaken by ACA Howe International Limited (ACA Howe) and Micromine Consulting Services

UK (MCS).

It is the opinion of ACA Howe and MCS that resources estimated as part of this study meet with

CIM/JORC Inferred and Indicated category classifications based upon quality of input data, modelling and estimation methodology, interpolation criteria based on sample density, search and interpolation

parameters, understanding and robustness of the geological model, drilling and sample density.

The resource estimation has an effective date of November 27th 2011 and represents a maiden NI 43-101

compliant resource estimation for the project.

ACA Howe International Limited and Micromine Consulting Services (UK), completed studies according to NI 43-101 and best practice guidelines. Resource modelling and estimations being

completed using the industry accepted Micromine 2011, 3d modelling software package.

Raw data used in interpretation and modelling consists of data from recent and historical diamond

drilling and RC drilling, low grade stockpile trench and RAB sampling exploration work undertaken by

Alhambra and previous explorers.

Raw data used as input to estimation consists of recent diamond, RC drill data and trench data.

The Dombraly project comprises low grade stockpile, pit infill and in-situ structurally controlled

mineralisation types. Mineralisation was modelled using natural cut-off grades of 0.13g/t Au, 0.1g/t Au

and 0.2g/t Au for the low grade stockpile, pit infill and in-situ mineralised zones respectively.

Several mineralised domains were modelled for resource estimation:

Low Grade Stockpile (Upper/Lower)

Pit Infill material

In situ near surface Oxide Au domain

In situ domain structure A

In situ domain structure B

In situ domain structure C

In situ domain structure D

In situ domain structure E

In situ domain structure F

In situ domain structure G

In situ domain structure X

In situ domain structure Y


88

In situ domain structure Z

Classification of resources for the low grade stockpile and pit infill are restricted to Indicated and Inferred.

Classification of in situ resources is restricted to inferred category due to the following factors which

introduce uncertainty:

o Limited number of valid drill holes and drill samples clustered in small areas.

o The number of valid drillholes are widely spaced along domain extents o A very low number of valid samples per mineralised domain

o A low number or no bulk density data for a number of domains and sub-domains

o Lack of QAQC data, and quality control issues

On working through the estimation process, it became clear that although the in situ deposit models are

coherent and robust based upon an interpretation of combined historical and recent (valid) drilling, the

domains require significant additional drill testing to increase valid input sample data numbers and sample density for both grade and bulk density determination, and improved resource block

classification.

Quality control sample data analysis and interpretation raised a number of issues with respect to assay

precision and repeatability. This could be due to nugget effect or sampling error, and will require

follow up investigation studies.

Due to these reasons the restriction and selection of resource classification currently applicable to the

deposit areas are deemed appropriate, particularly for in-situ domains.





according to best practice geological modelling techniques, current understanding of the geological

model, historical interpretations and discussion with Alhambra personnel as described in previous

sections.


dependent upon the cost of mining and processing the mineralisation and the selling price of the final

product.

The economic cut-off grade for Dombraly was established using grade and block revenue factors.

Due to the early stage status of the development of the Dombraly deposits, a number of assumptions

have been made with regard to inputs to the calculation of the economic cut-off grade for reporting.



Inputs for oxide material are based upon actual mining cost data from Alhambra’s nearby Uzboy open pit operation, and estimated costs for transitional and primary material taken from recent PEA studies

undertaken on the nearby Uzboy deposit.





89




Oxide mining cost – US$1/t (waste dump and pit infill)/US$1.7/t (in-situ)



Using the Au block grade, the above Au metal price and recovery, MCS estimated the revenue per mined block.

For a mineralised block to be considered economic it must generate higher revenue than it costs to mine. For a block to be considered economic it must therefore generate greater than US$4.85/t of revenue for

low grade stockpile and pit infill material, and US$5.55/t and US$8.42/t, for in-situ oxide and

transitional/primary material respectively.

MCS used Micromine software to filter those blocks in the resource model with value greater than the

calculated cost to mine values for economic cut-off grade determination and resource reporting.

Cut off grades used for reporting are 0.1g/t Au for thelow grade stockpile, 0.2g.t Au for the pit infill zone and 0.1g/t for in-situ oxide material, and 0.2g/t Au for in-situ transitional and primary material

types.

It is MCS’ opinion that the assumptions made for input to economic cut-off grade determination and

reporting of potentially economic resources are reasonable given the current understanding of the geology, mineralisation, anticipated mining and processing methods and comparison with similar type

operations.

At Dombraly, the historical extraction of the upper part of the oxidation zone was carried out in 1985 to 1988 when the price of gold was very low at US$320 to US$460 per troy ounce, using a gold cut-off

grade of 2.5 g/t Au. Upon completion of the extraction of oxidised material, the mine was partially

reclaimed by backfilling with the mined rock then regarded as waste. At that time the gold-bearing mineralisation left in the bottom and sides of the pit probably did not have any commercial value due to

its relatively low grade compare to the grade of the ore which was produced (6.96 g/t Au). With

improved heap leaching techniques and gold prices now about 3 times higher, the economics of

potential gold production at Dombraly are again attractive even at much lower grade, well below the historical cut-off grade.

Based on available information, ACA Howe and MCS believe that the exploration and resource development of Dombraly is progressing well and that there is scope to develop a potentially

economically viable gold resource.

Scope remains for the improved resource classification via infill drill testing, and additional resource

tonnages via step out drilling to test in situ mineralisation open along strike and depth.

Open pit production took place in 1985 to 1988 when the price of gold was very low, at US$320 to US$460 per troy ounce, using a gold cut-off grade of 2.5 g/t Au. With improved heap leaching

techniques and current gold prices, the economics of potential gold production at Dombraly are again

attractive even at much lower grade, well below the historical cut-off grade.

6.16 DOMBRALY - RECOMMENDATIONS

Several issues and sensitivities have been highlighted as part of this study and are outlined below.


90

These issues ultimately impact on the robustness and confidence of the geological and resource model

and should be considered for improved assessment, estimation of higher classification of resources and mine planning.

Data Collection

No geotechnical/geomechanical logging is taking place at drill site, and although core recoveries are

reported to be very good overall, it is best practice to perform orientation marking, metre marking,

recovery, RQD and fracture frequency logging, prior to transportation. Transportation of the core could result in the disintegration of less competent zones (which generally tend to be those of greatest

interest).

The lack of drill core orientation is an issue. Much more information could and should be collected from drill core given the structural complexities that exist at the deposit. Greater accuracy with metre

marking, RQD measurements, along with performing angle and orientation (alpha/beta) measurements

of features relative to orientation line and core axis is essential in these deposit types. The importance of oriented drill core and measurement of controlling structures will become critical as exploration and

work towards improved resource classification within the project continues.

Sampling is performed to a reasonable standard. However, due to the non-orientation of the drill core, inconsistencies are introduced into the system. Without orientation line control, the core will be

sampled randomly along its axis, and may well introduce bias to the samples. Best practice is to sample

along the core orientation line thus being consistent and selecting a sample perpendicular to the perceived strike of mineralisation and mineralised veins etc.

The sampling methodology is considered good practise in this type of deposit and is suitable for gaining a detailed understanding of lithological host rocks and controls of mineralisation.

Bulk density date is lacking from a number of areas. A greater number of bulk density samples are

required from all zones.

Analyses

Sampling preparation methodology appears reasonable and satisfactory.

Sample dispatch routines, and dispatch record sheets which have been observed and considered

consistent and of a satisfactory standard.

Sample security protocols were discussed with site personnel and demonstrated to MCS during the

project site visit. These are considered entirely satisfactory.

Checks on accuracy of analytical process and equipment for DDH appear satisfactory, however

duplicate analysis for RC and grab samples are seen show low correlation and requires follow up.

This is of concern with regard to nugget effect and/or the preparation of samples potential

contamination and introduction of error at the sample preparation stage.

It is MCS’ opinion that although overall satisfactory quality control and assurance for the current dataset

does raise some concerns and issues with regard to precision and accuracy of analysis. These issues

should be investigated further as part of the ongoing QAQC process to enable improved confidence in

updated resource estimations. Recommendations include:

Use of field duplicates

Verification twin drill holes and trenching


91

Domain modelling

Geology and controls are reasonably well understood given the available data for the particular study area, and historical models. Better understanding can be gained by improved alteration and structural

logging (oriented core), which will ultimately improve resource models and resource confidence.

Metallurgical testwork

Further metallurgical test work is required to obtain representative values across all deposit areas and

material types.

In addition MCS recommends the following work is undertaken to progress the deposit areas toward

improved classified and reportable resource and reserve block model estimations.

Further drill testing of current exploration solid models along strike and at depth.

Detailed geological, structural and grade domain modelling of the new drilling areas and

development of exploration/deposit model and concepts in the current near surface

exploration areas to assist further exploration and resource estimations.

Refined domain geostatistics to determine optimal natural grade boundaries.

Domain statistics and variography to determine optimum exploration drill spacing.

Further historical data import, modelling and verification by possible means of selected

channel and or drillhole re-sampling and analysis, twin drillhole verification, comparison of historical exploration data versus production data.

Follow up and resolve current data validation issues.

Detailed review of literature and deposit model types.

Audit of analytical laboratory and review of certification/accreditation.

Review and/or updated metallurgical test work studies.

It is recommended both the historical and current database and the wireframe models be constantly

updated and should always reflect the latest stage of the exploration so that changes or adjustments to any future exploration programs (planned drilling location), can be made immediately with best

available data and interpretations.


92

7 REFERENCES AND OTHER SOURCES OF INFORMATION

General

ACA Howe International Limited, 10 December 2009. Updated Scoping Study on the oxide,

transitional and primary resources at the Uzboy gold deposit, Akmola Oblast, Kazakhstan. For

Alhambra Resources Ltd. Bekzatov, A., 2004. Gold of Kazakhstan: Brief Overview. KAZAKHSTAN International Business

Magazine No. 1, 2004 (by Adil Bekzatov, Industry Analyst of OJSC Kazkommerts

Securities). Published on website http://www.investkz.com/en/journals/38/164.html Plyushchev, E., 28 March 2011 email. Draft exploration program. File name: EP proposal exploration

program 3-8-2011 ACA Howe.doc

Stiskin, M., Donskoy, S., Lapshina, I., and Zhunisov, Z., December 2010. Russian Metals and Mining, Russian Gold Sector: Stars Aligned. Troika Dialog private investment bank.

http://traders.net.ua/_ld/12/1244_101223-1014_MET.pdf (Map including Kazakhstan on page

30)

Dombraly

ACA Howe International Limited, 2 June 2008. Dombraly site visit notes. By geologist Galen White addressed to Elmer Stewart and John Komarnicki of Alhambra Resources Ltd. Filename:

DombralySiteVisitNotesGalenRWhite02June2008.doc

ACA Howe International Limited, 6 September 2010. Notes on the current status of the ACA Howe

estimation of waste pile, pit backfill and oxide/sulphide zone resources at the Dombraly Project. By geologist Leon McGarry. Filename:

ACAHoweDombralyProjectStatus06September2010.doc

Alhambra website, January 2011a, http://www.alhambraresources.com - exploration projects: ALH website - Dombraly Final detailed writeup.pdf. (Contains information originally reported in

Russian.) Filename: ALH website - Dombraly Final detailed writeup.pdf

Anon., 2007. Geological report with ore and gold resource calculation for the Dombraly II deposit (as of Jan. 1, 2007), Library of SCGC. (Contains information originally reported in Russian.)

Begayev, I.V., Teleshev A.A., et al., 2006. Technical and economic substantiation of the parameters of

oxide gold ores from the Dombraly II deposit as of Jan. 1, 2006, Library of SCGC. (Contains

information originally reported in Russian.) Begayev, I.V. and A.A. Teleshev, et al., 2007. Report on Ore and Gold Reserves Estimation,

Dombraly II Gold Deposit (as at January 01, 2007), 2 volumes. 159 pages of text, 56 graphical

Figures on 57 pages, 36 Tables, bibliography – 22 March 2007. (Territorial Administration “Sevnedra”, Joint Venture «Saga Creek Gold Company LLP.”, “Geos Ltd.”, North-

Kazakhstan Oblast, No. -43-109. (Contains information originally reported in Russian.)

Filename: Dombraly II report (eng)10.12.07.doc Bolotova, L.S., July 2005. Metallurgical test of Technological Sample DTL-1 collected from

Dombraly gold deposit. National Centre of Complex Processing of Minerals of the Republic

of Kazakhstan (Kazmechanobr). (Contains information originally reported in Russian.)

Filename: 1Report Dombraly DTL1(eng)July2005.doc Bolotova, L.S., December 2005a. Studies of heap leach gold recovery from ore test sample DLT-1 -

add, collected from Dombraly gold deposit. National Centre of Complex Processing of

Minerals of the Republic of Kazakhstan (Kazmechanobr). (Contains information originally reported in Russian.) Filename: 2Report Dombraly DTL1ad (eng)Dec2005a.doc

Bolotova, L.S., December 2005b. Studies of small Technological Samples DTK-1 – DTK 11 collected

from Dombraly - II gold deposit. National Centre of Complex Processing of Minerals of the

Republic of Kazakhstan (Kazmechanobr). (Contains information originally reported in Russian.) Filename: 3Report Dombraly DTK1-DTK11(eng)Dec2005b.doc

http://www.investkz.com/en/journals/38/164.html

http://traders.net.ua/_ld/12/1244_101223-1014_MET.pdf


93

Bolotova, L.S., September 2006. Laboratory studies of concentrability of sulphide ore of Dombraly -

II gold deposit (Sample DLT-2). National Centre of Complex Processing of Minerals of the Republic of Kazakhstan (Kazmechanobr). (Contains information originally reported in

Russian.) Filename: 4Dombraly LTD-2 Flot_Grav(eng)Sept2006.doc

Bubareva, N.V., Nikel V.G., et al., 2005. Report on hydrogeological geoecological studies at the

Dombraly II deposit, Library of SCGC. (Contains information originally reported in Russian.) Tarnovski, Y.I., 1953. Report on prospecting and exploration group for 1953. Geological setting of

map sheet N-42-132-A, Library of SCGC. (Contains information originally reported in

Russian.)


94

8 DATE AND SIGNATURE PAGES

CERTIFICATE and CONSENT of AUTHOR

With reference to NI 43-101, I, John Langlands, do hereby certify that: (a) I am currently employed as Principal Geologist by:

ACA Howe International Limited

254 High Street, Berkhamsted, Hertfordshire, HP4 1AQ

United Kingdom (b) The title and date of the Technical Report to which this certificate applies are as follows:

Title: Technical Report Introduction and Resource Estimation for the Dombraly Gold Deposit in

north-central Kazakhstan. Date: 22 March 2012. (c) I am a graduate of the University of Edinburgh and hold a B.Sc. Honours degree in Geology (1969)

and a Diploma in Resource Management (1980). I have been employed as a geologist for 42 years since graduation and with ACA Howe International Limited since 1980. I am a Fellow of the Institute

of Materials, Minerals and Mining (formerly the Institution of Mining and Metallurgy), a Fellow of

the Geological Society and I am a Chartered Engineer with the Engineering Council. I certify that by

reason of my education, Fellowship of the Institute of Materials, Minerals and Mining and relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

(d) I have not visited the properties which are the subject of the Technical Report.

(e) Together with the co-author, James Hogg, I am responsible for the overall structure and content of the Technical Report.

(f) I am independent of the issuer since there is no circumstance that could, in my opinion and the

opinion of a reasonable person aware of all relevant facts, interfere with my judgment regarding the preparation of the technical report.

(g) I have not had prior involvement with the issuer or the property that is the subject of the Technical

Report, other than as an independent consultant to the issuer.

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

(i) As of the date of the Certificate, to the best of my knowledge, information and belief, the Technical

Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

With reference to NI 43-101, Part 8 Certificates and Consents of Qualified Persons for Technical

Reports, 8.3, (a), I, John Langlands, address the following statement to the securities regulatory authority:

(a) I consent to the public filing of the Technical Report and to written disclosures of extracts, or the

summary, of the Technical Report, subject to other conditions of NI 43-101.

Dated this day 22 March 2012.

John Langlands, BSc, FGS, FIMMM, CEng.


95

CERTIFICATE and CONSENT of AUTHOR

With reference to NI 43-101, I, James Emberton, do hereby certify that: (a) I am currently a Senior Associate Geologist with:

ACA Howe International Limited

254 High Street, Berkhamsted, Hertfordshire, HP4 1AQ

United Kingdom (b) The title and date of the Technical Report to which this certificate applies are as follows:

Title: Technical Report Introduction and Resource Estimation for the Dombraly Gold Deposit in

north-central Kazakhstan.

Date: 22 March 2012 (c) I am a graduate of the University of Durham and hold a B.Sc. degree in Geology (1961). I have

been self-employed as a consultant geologist since 1984. Prior to that I was employed as a

mining/exploration geologist both in Australia and UK for 21 years. I am a Fellow of the Institute of Materials, Minerals and Mining (formerly the Institution of Mining and Metallurgy and a Chartered

Engineer with the Engineering Council. I certify that by reason of my education, Fellowship of the

Institute of Materials, Minerals and Mining and relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

(d) I have not visited the properties which are the subject of the Technical Report.

(e) Together with the co-author, John Langlands, I am responsible for those sections of the Technical Report describing the property, regional and local geology and mineralisation.

(f) I am independent of the issuer since there is no circumstance that could, in my opinion and the

opinion of a reasonable person aware of all relevant facts, interfere with my judgment regarding the

preparation of the technical report. (g) I have not had prior involvement with the issuer or the property that is the subject of the Technical

Report, other than as an independent consultant to the issuer.

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

(i) As of the date of the Certificate, to the best of my knowledge, information and belief, the Technical

Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.


Reports, 8.3, (a), I, James Emberton, address the following statement to the securities regulatory authority:

(a) I consent to the public filing of the Technical Report and to written disclosures of extracts, or the


Dated this day 22 March 2012.

James Emberton, BSc, FIMMM, CEng.


96

J. N. HOGG

Second Floor, Challoner House

19 Clerkenwell Close

London

EC1R 0RR

United Kingdom

Telephone: +44 (0)203 176 0080

Email: [email protected]

CERTIFICATE of AUTHOR and CONSENT of AUTHOR

I, J. N. Hogg, MSc., BSc., MAIG do hereby certify that:

1. I am currently employed as a Senior Resource Geologist by:

Micromine Limited Second Floor, Challoner House


London EC1R 0RR

United Kingdom

2. I graduated with a Bachelor of Science degree (Hons) in Geology from

Kingston University, Surrey, UK, in 1993. In addition, I obtained a Masters of

Science (merit) in Mineral Exploration in 1996 from the University of

Leicester, Leicestershire, UK.

3. I am a member of the Australian Institute of Geoscientists, and Prospectors

and Developers Association of Canada.

4. I have worked as a geologist for a total of 15 years since graduation from

university. Relevant experience includes 8 years exploration, resource and reserve development of lode gold, silver and base metal deposits in Western

Australia with Delta Gold NL, Sons of Gwalia Ltd and Newmont Australia

and 7 years as consultant resource geologist initially with ACA Howe

International Limited and later Micromine Consulting Services.

5. I have read the CIM code, and definition of “qualified person” set out in

National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101)

and past relevant work experience, I fulfill the requirements to be a “qualified

person” for the purposes of NI 43-101.

6. Together with the co-author, John Langlands, I am responsible for the overall

structure and content of the Technical Report.

7. I have not conducted a site visit to assess data collection methodologies,

auditing and data verification exercises for the purpose of this resource

estimation report. This was undertaken by Mr. Evgenij Zhuravlyov, Senior Geologist, Micromine Consulting Services (Kazakhstan), between the dates

12th and 14

th August 2011.

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


97

Report, the omission to disclose which makes the Technical Report

misleading.

9. I am independent of the issuer applying all of the tests in section 1.4 of

National Instrument 43-101.

10. I have no prior involvement with the project.

11. I have read and am familiar with CIM code and National Instrument 43-101 and Form 43-101F1. The Technical Report has been prepared using those

reporting guidelines.


Reports, 8.3, (a), I, James Hogg, address the following statement to the securities regulatory authority:

(a) I consent to the public filing of the Technical Report and to written disclosures of extracts, or the summary, of the Technical Report, subject to other conditions of NI 43-101.

Dated this day 22 March, 2012.

“J. N. Hogg”


98

M.SOSTRE

Second Floor, Challoner House


London

EC1R 0RR

United Kingdom

Telephone: +44 (0)203 176 0080

Email: [email protected]

CERTIFICATE of AUTHOR and CONSENT of AUTHOR

I, M. Sostre, MSc., BSc., AUSIMM do hereby certify that:

1. I am currently employed as a Senior Resource Geologist by:

Micromine Limited

Second Floor, Challoner House 19 Clerkenwell Close

London

EC1R 0RR United Kingdom

2. I graduated with a Bachelor of Science degree (Hons) in Geology from

Zaragoza University, Spain, in 1998. In addition, I obtained a Masters of Science (merit) in Mining Geology in 2009 from the Camborne School of

Mines (University of Exeter), UK.

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

(AUSIMM).

4. I have worked as a geologist for a total of 10 years since graduation from university. Relevant experience includes structurally controlled lode gold

system modeling and geostatistics whilst spending 2 years as consultant

geologist with Wardell Armstrong Limited and 4 years resource consultant with Micromine Consulting Services.

5. I have read the CIM code, and definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and in spite of my education,

affiliation with a professional association (as defined in NI 43-101) and past

relevant work experience for this particular deposit type, I do not fulfill the

requirements to be a “qualified person” for the purposes of NI 43-101 at this time.

6. I am responsible for the content of section 6.10 under the supervision of qualified person, Mr. James Hogg MAIG, Senior Resource Geologist,

Micromine Ltd.

7. I have not conducted a site visit to the property.

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.


99

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

10. I have no prior involvement with the project.

11. I have read and am familiar with CIM code and National Instrument 43-101

and Form 43-101F1. The Technical Report has been prepared using those

reporting guidelines.


Reports, 8.3, (a), I, Marta Sostre, address the following statement to the securities regulatory authority: (a) I consent to the public filing of the Technical Report and to written disclosures of extracts, or the


Dated this day 22 March, 2012.

“M. Sostre”


100

APPENDIX 1. MCS DOMBRALY

SITE VISIT REPORT


101

MICROMINE CONSULTING

REPORT ON PROCESS QUALITY CONTROL

AT THE DOMBRALY AND SHIROTNAIA DEPOSITS

Almaty, 2011


102 Authors of the Report on Process QA/QC at the Dombraly and Shirotnaia Deposits of Sary-Arka

licensed territory of Saga Creek Gold Company LLP visited the stated deposits for the purpose of inspection of

conducted geological exploration works, their scope and quality. The deposits are explored on the basis of issued

licenses No.719ДД and No. 1029Д. The Dombraly and Shirotnaia gold deposits are located in North Kazakhstan Oblast of the Republic of Kazakhstan in the region with developed infrastructure. The major gold fields known

here are Vassilkovskoe, Stepnyak, Aksu, Bestobe, Zhelombet, also there are uranium, tin, industrial diamonds,

titanium and zircon deposits. Underexplored deposits of Bailyusty, Severnoe Bailyusty and Kimaly are located nearby the Dombraly and Shirotnaia deposits.

The Dombraly deposit was discovered in 1966 and from 2002 it has been explored by Saga Creek Gold Company LLP. Geological-economic appraisal (together with Geos LLP) based on the results of estimation by alternatives

of oxide-bearing gold ores was carried out in 2006. Feasibility study of evaluation conditions was approved by the

State Reserves Committee of the Republic of Kazakhstan (the RoK SRC) (Minutes of Meeting No.496-06-K

dated 21.03.2006). Obtained technical and economic indicators prove the possibility of cost-effective development of gold-sulphide ores together with technologic mineral formations by open-cut mining. Oxide

reserve mining in solid of the open pit was considered to be economically impractical due to a high strip ratio and

insufficient content of gold in oxide ore. However, with current prices on gold this position will be reconsidered. Total reserves in subsoil of oxide ores, re-cultivated mined rock in open pit and industrial dump of С2 category

amounted to: ore – 4254.1 thous. tons, gold – 6478.8 kg, average gold content – 1.52 g/ton. Reserves of gold-

sulphide ore to the depth of 300 m on С2 category amounted to: ore – 575.6 thous. tons, gold – 3119.6 kg, average gold content – 5.42 g/ton. Geological exploration works are still ongoing.

The Shirotnaia deposit is located in Akkol region, 3 km north-east from Aksu settlement. Most of the deposit is

located in stream-valley and is bridged over with recent sediments. Exploration works in 2002-2011 were carried out on the territory of the whole deposit by trenching, drilling core holes, RCC holes and air drilled holes. These

works resulted in detecting of 2 thick zones of mineralization including the group of large ore bodies and lenses.

On the south-west side, 520 m from the main body, there is a zone of a column-shaped mineralization with inclination 50-45

0 north-west. Exploration of main ore bodies was done by core holes to the depth of 120-140 m

(18 holes). North-east flank with length of up to 2 km is the most prospective for discovering new ore bodies.

Four bodies of ore mineralization with length of 1100 m and thickness of 5-40 m with average gold content of

0.8-1.1 g/ton were discovered as a result of exploration drilling works on this flank.

As of 01.01.2010, according to the preliminary estimation, gold reserves in the area of oxidation on С2 category

amounted to 3900 kg with average gold content of 1.2 g/t. In oxide ores on P2 category the gold amounted to 40,000 kg with average content of 1.2 g/t. The exploration works are still ongoing on the site.

The results of analyses on the main part of holes samples for this year haven’t been obtained yet.

The results of works are stated in corresponding reports and are done using Microsoft Excel, Word, MapInfo,

Access, Corel Drаw. Interpretation of geological exploration data on the Dombraly and Shirotnaia deposits was

carried out by Micromine program.

The authors are thankful to the management and the chief specialists: John Komarnitski (Alhambra company) and

Alexander Miroshnichenko, Yevgeniy Plyuschev, Stepan Trofimov (Saga Creek Gold company) for warm welcome and arranging details allowing visiting the Dombraly and Shirotnaia deposits on August 11-15, 2011, as well as seeing

their infrastructure and collecting the required geological data.

1. Brief Geological Structure of the Contract Area

The contract territory (Licenses No.719ДД and No.1029Д) covers north-east part of Kokchetav-North-Tian-Shan

Caledonian mosaic fold system. In the northern part of Kokchetav-North-Tian-Shan fold system there are four Caledonian structured formation megazones with almost north-south direction: Ishim-Karatau megazone

(western), Kokchetav-Ulutau anticline megazone (eastern), Stepnyak-Zhaksykon syncline megazone (central),

Yerementau-Boschekul anticline megazone (from the east).

The contract territory is mostly located in the central and north parts of Stepnyak syncline. From the west and

from the north, the part of Kokchetav block mass anticline is considered the part of the contract territory

(Kokchetav and Shat anticline) and from the east – the western part of Ishkeolmes anticline is also considered the part of the contract territory.


103

Figure No.1

Location map, scaled 1:1000000


104

Kokchetav block mass is divided by geologic and geophysical data into four structure blocks:

Shatskiy, Zagradskiy, Borovskoi and Zerendinskiy.

The Caledonian structures formed by the mid-Ordovician deposits of Stepnyak strata lay in the

structures of ancient folded body of Kokchetav block mass with dramatically different tectonic layout.

These mid-Ordovician deposits form rather small brachiform synclines. The largest ones have sub-lateral extension. Superposed folds made by glomerations and sand-rock of upper Ordovician also

have the same characteristics.

Stepnyak syncline is a complex late-Caledonian folded structure. It is formed by slightly

metamorphized deposits of mainly middle and upper Ordovician. There are several syncline and

anticline zones in the synclinorium.

Hercynian structural level is represented by wide and flat geological basins laid over consolidated

Stepnyak synclinorium and Ishkeolmess anticlinorium. Geological basins are formed by thick mass of

the volcanic rocks of the Lower-Middle Devonian, the red beds of the Givetian-Frankish age, and by carbonic rock.

2. General Information on the Dombraly Deposit

The Dombraly gold field is located within a sheet N-43-109-a. Geographical coordinates of the field center – N 52°55 ́and E 72°05 .́

The nearest railroad station “Aksu” is located in 70 km to the south from the field. The nearest

settlements in the area of the deposit are Zolotaya Niva (15 km west), Projektor (20 km east) villages, Komsomolskiy settlement (42 km north-east), Valikhanova settlement (40 km west), population of

which is mostly work on personal farms. All settlements are interconnected by country roads that are

only suitable for vehicles at summer. The terrain of the deposit area is plain with highly compressed ridges, ranges, wide drainless hollows and lake degradations. Sea level is 220-325 m with local

difference in elevation being 1-5 m, rarely 8-10 m.

The area exposure is very poor. Segmental rock outbreaks of effusive rock are rare. The whole area of works is blocked by poorly consolidates sediments of the Quarternary age, thickness of which on the

average does not exceed 10 meters. Hydrographic network of the area is underdeveloped and is

represented by rare dried river channels and Kizdyn-Karassu and Karassu rivers.

Rivers’ water schedule is of a seasonal nature and their activity depends on the spring thaw and

periodic rains. During the summer the rivers look like the chain of narrow, shallow and isolated from each other pools. River beds are formed by clay loam, sand clay, clays and rarely by sand and gravel

material. River valleys are meandering with insignificant slopes (0.001-0.002), smooth slopes, poorly

defined terraces above flood-plain and flood-plains. River valleys are usually swamped.

The climate of the region is extremely continental. The coldest months are January and February with

average monthly temperatures varying from -17, -20 up to – 35.4. The highest average monthly

temperatures (+18, +22) are in June-July and reach up to 35.3.

Average annual precipitation varies from 250 to 300 mm. Long-term average annual precipitation is

268 mm. Prevailing wind direction at summer is west and south-west; at winter – north-west, west and

rarely north-east and east. The territory can be classified as sheep fescue/feather-grass steppe of the North Kazakhstan as per the nature of vegetation. Xerophilous narrow-leaved gramineous plants like

feather grass, Stipa capillata and sheep fescue form a rather monotonous background. In low places

carpet plants become denser, gramineous plants are replaced with miscellaneous herbs. Birch and aspen groves with bushes grow in swamped low lands.


105

Soil cover is represented by reddish brown and light-gray chestnut, loam and sandy loam soils with

thin humus horizon (0.1-0.5 m). Fauna is defined by the peculiarities of the landscape. Mammals are mostly represented by rodents: prairie dogs, hamsters, field mice, marmots and such animals as hares,

wolves and foxes. Large mammals (roedeers and roars) are rather rare. There are enormous numbers

of mosquitoes, midges, horse-flies, gadflies, flies and mites. There are a lot of birds are well. Most of

them are of passerine and sandpiper families.

Photo 1. Open cut at the Dombraly Deposit

Photo 2. View of the open pit from the low grade stockpile

Electricity for a future mining processing plant will be supplied via LEP-35 kV (power line)

constructed by the subsoil user from Zolotaya Niva village (15 km) to the deposit.

Supply of drink and process water for the needs of processing facilities will be ensured in required

volumes from the wells drilled on the territory of the deposit.

Explored fields of industrial minerals are unknown in this area. Surroundings of the deposit to the

radius of 20-25 km are prospective for exploring deposits of non-metal rock and industrial minerals


106

deposit required for arranging ore processing production by heap leaching method: waterproof clay

sand and gravel mix.

Large-scale geological exploration works in the area of Dombraly deposit were started in 1952 by

searching party of the Kazzoloto trust of quartz-lode deposit of Severnoe gold field (Severnoe

Bailyusty). Evaluation of this site was also conducted by this party in 1966-1968, and in 1969-1972 by Tselinograd Exploration Company of the Central Kazakhstan Territorial Geological Administration.

During the exploration period oxide zone was explored by mine opening on surface (pit-holes, ditches, test pits, bell-pits) and by underground mining workings (shaft, roadways and crosscuts to it) on the

horizon 30 m from the surface and down to 170 m by core holes. During exploration period,

technological survey of oxide ores for developing a washing process layout was done. Reserves estimation was approved by the Territorial Reserves Committee of the Central Committee of

Production Geological Association (Minutes of Meeting, No. 3-411, dated 27.02.81). Estimated

reserves were assigned to the balance of Kazzoloto industrial complex, except the reserves of С2

category.

In 1985-1988 the deposit was mined to the horizon of 60 m from the surface by Enbek prospector’s

team of Kazzoloto Mining and Processing Complex.

For the period of operation, 140 thousand tons of ore and 949 kg of gold with content of 6.96 g/t were

extracted.

In 2002-2005 geological exploration works at the Dombraly deposit were re-started by JV Saga Creek

Gold Company LLP. The purpose of the works of this period was re-estimation of oxide ore reserves

in pillars with account to the gold in the re-cultivated rock mass in the open pit for determining the opportunity for their processing by heap leaching technology. During the works on preliminary

economic-geological evaluation it became clear that it was economically impractical to process all

oxide ore reserves due to high overburden ratio and insufficient gold content in the oxide ores. Cost efficient development is possible only for some part of oxide ore reserves (by reducing the depth of

final open pit). At that, the reserves of oxide ores for development will be decreased by 33.8%.

Complete development of gold reserves in oxide zone is only possible by taking dump technogenic

mineral formations to heap leaching processing. Reserves estimation in technogenic mineral formations was carried out in 2005. Geological exploration works on estimation of gold-sulphide ores

were conducted together with re-estimation of the reserves of gold in oxide area.

Volumes of the reserves as of 01.01.2008 of the Dombraly deposit are shown below:


107

Reserves allocation at the Dombraly deposit

Table 2.1

Estimation of oxide ores in situ and in technogenic formations of the Dombraly deposit was done on

the basis of the results of geological exploration works carried out by the Dombraly Geological

Exploration Crew of Tselinograd Exploration Company of TsKTGU in 1969-1972 and by JV Saga Creek Gold Company LLP in 2002-2005 on the basis of estimation conditions approved by the RoK

State Reserved Committee (Minutes of Meeting, No.496-06-K, dated 21.03.2006). Estimation

parameters accepted by the analogy to estimation conditions for the gold-sulphide ores of the Uzboy

deposit (underground development method) were used for estimation of the gold-sulphide ore reserves.

Chemical analyses were made at Kvartz Chemical Analytical Laboratory LLP and Tsentrgeolanalit CJSC. Technological survey was done by Kazmekhanobrom. Such works as: field survey,

hydrogeological, engineering-geological works and environmental monitoring, were conducted by the

Karagandy Branch of Azimut Company.

On the balance Extracted in 1985 - 88

Re-estimation of the reserves as of 01.01.2008

Category Ore,

thous.to

ns

Average

content,

g/t

Gold, kg Ore,

thou

s.tons

Aver

age

content,

g/t

Gold

, kg

Category Ore,

thous.to

ns

Avera

ge

content, g/t

Gold, kg

С1 87.8 10.82 949.9 С1

С2, including С2,

including

In situ ore In situ ore 1451.9 2.05 2981.7

Re-cult. mass Re-cult.

mass

1060.3 0.90 951.1

Low grade

stockpile

Low grade

stockpile

1742.0 1.46 2546.0

total С2 34.9 8.63 301.4 total С2 4254.1 1.52 6478.8

С1+С2 122.7 10.20 1251.3 140 6.96 975 С1+С2 4254.1 1.52 6478.8

С2 С2 575.6 5.42 3119.6

С1 87.8 10.82 949.9 С1

С2 34.9 8.63 301.4 С2 4829.7 1.99 9598.4

С1+С2 122.7 10.20 1251.3 140 6.96 975 С1+С2 4829.7 1.99 9598.4


108

2.1 Geological Structure of the Dombraly Deposit

The Dombraly II deposit is located in the Stepnyak synclinorium close to its connection to

Ishkeolmess anticlinorium. It is referred to the central part of the Bailyust synclinal zone complicating

the Stepnyak synclinorium with north-north-west trending in the area of the deposit. The total length

of the structure is approximately 100 km with width of 11 km. The deposit is located within the Dombraly tectonic zone of north-north-west trending. Within this structure there are several gold ore

occurences: to the north of the deposit - Kimaly gold polymetallic ore occurrence and small quartz-

lode deposit of gold Severnoe Bailyusty, to the south – Bailyusty gold field and other gold ore occurrence.

2.1.1 Stratigraphy

Igneous-sedimentary rocks of Maylisor, Mayatass (Bailyust) and Shat (Koksengirsor) series of upper

Ordovician take part in forming the geological structure of the deposit.

Maylisor series (upper-Caradoc Sub-stage – Ashgillian Stage).

Diabase-spilite strata of Maylisor series forms the western wing of synclinal fold. It is formed by dark-green-gray spilites and diabases with interlayers of tuffs and tuffaceous sandstone. Diabases in form of

thin dikes are found in the layer of spilites.

Mayatass (Bailyust) series (upper-Caradoc Sub-stage – Ashgillian Stage).

Deposits of these series form the core part of the Ordovician synclinorium. The horizon of ash-gray

tuffaceous sandstones and silty rocks with lens of psammite and psammite-psephitic tuffs of average

composition lies in the base of the series. It is separated from the overlying horizon by interlayer of basal conglomerates. Polymictic sandstones, silty rocks, argillites of gray, green and lilac color with

interlayers of silicified and carbon-bearing silty rocks and lens of marmorized limestone are developed

in the upper part of series cut.

Shat (Koksengirsor) series (upper-Caradoc Sub-stage – Ashgillian Stage). Deposits of these series

form core parts of small synclinal folds of sub-meridian and north-west trending. They are mostly

represented by tuffs of andesitic composition interlaying with coarse-grained tuffaceous sandstone.

Mesozoic Weathering Crusts. Weathering crusts developed throughout the whole territory of the site,

they are concealed under the Cenozoic deposits. Areal and fractured-linear types of crust can be observed. Areal weathering crusts in section are characterized by subhorizontal occurrence of

subzones. In the linear-fractured saprolites the subzones occur aslope matching the attitude of tectonic

dislocations, on which they develop.

As per the degree of changes in the section of a crust the following five are weathered (from top to

bottom):

Structureless clayish weathered crust composed from reddish-yellow, light-gray-green, light-brown, grayish- and pinkish-red clays, sometimes, with inclusion of ferrous nodules. Indistinct relics of

structures and textures of parent rock can be rarely detected in the lower part. Possibly, it corresponds

to the zone of colored kaolinites by its composition.

Structured clayish, loamy, arenaceuos weathered crust (corresponds to the upper part of intermediate

weathering) is represented by clays, and with high content of quartz in parent rock it is represented by clay loam and sand clay with usually distinct relict structure and texture of parent rock. Thickness is

from 4 to 9 m, in the areas of tectonic jointing it increased up to 18 m.

Rubble-clayish weathered crust (the volume of break stone of weathered parent rock being less than 50%). The most of the rock volume is clays, clay loam and sand clay. It corresponds to the lower part

of intermediate weathering area. Thickness is from 1.5 to 8-12 m, in the areas of jointing – up to 17 m.


109

Rubble-clayish weathered crust (the volume of break stone of weathered parent rock being more than 50%), corresponds to the upper part of rock breaking area. Thickness is from 0.2 to 13 m.

Intensively weathered-fractured rocks corresponding to the lower part of rock breaking area.

Thickness is from 6-10 m to 19-26 m. Total thickness of weathered crusts throughout the territory varies from 21 to 120 m.

Selective concentration of gold in any subzone of weathered crust hasn’t been detected.

The previous exploration works detected non-availability of a significant gold migration in the process

of crust formation. There was only coarsening of gold with insignificant concentration of oxide ores comparing to the initial condition. Concentration did not exceed 5-10% and occurred by partial

removal of parent rock material during crust formation.

Cenozoic Group

Deposits of the Cenozoic group are represented by eluvial formations covered by clayish deposits and

topsoil.

Eluvial formations developed on the Ordovician are detected everywhere, lithologically represented by

sand clay, clay loam, clays, rubbly rock and argillite. They were uncovered on the depth of 0.20-2.00

m, uncovered thickness reaches 118.50 m (hole 19).

Amongst clayish rock the most frequent is clay loams, both – in plan view and by depth. They are

uncovered on the depth of 0.2-11.20 m, maximum uncovered thickness reaches 41.60 m (hole 19).

Sand clays and clays are distributed sporadically. Clayish soils are multicolored, moist, hard, solid, ferruginous; in the upper part of geological section they are with fractures filled with quarternary clay

loams, often with semi-rock and rock chippings of up to 30-40%.

Rubbly rock with clay loam filling reaching up to 30% is of dark-violet color, was uncovered by hole 20 at the depth of 37 m, uncovered depth – 7.5 m.

Argillites are dark-violet, gray with bluish shade, crumbling, rather frail, with low ferruginization and interlayers of clay loam in the beginning of the interval. Uncovered by holes 19 and 20 at the depth of

42.6, 44.5 m. Maximal uncovered thickness is 47.8 m (hole 19).

Topsoil is almost everywhere. Its thickness is 0.15-0.30 m.

2.1.2 Subvolcanic Magmatism

Subvolcanic intrusions of the upper Ordovician. Intrusions are represented by several separate bodies

of andesite porphyrites, dolerite and dikes. Subvolcanic bodies of andesite are logged on the south flank of the ore field. In plan view they have lenticular shape. Their size on long axis is 800-900 m,

and on short – 200-400 m. They are characterized by greenish-brown color, porphyric structure and

small/medium-grained ground mass. Porphyric buildups are numerous and large (up to 5-10 mm), they are represented by plagioclase and hornblende.

Subvolcanic body of dolerites is located 1 km north-east from the deposit. It has lenticular shape and

sizes of 480x180 m.


110

Dikes of diabases are localized within the ore area of the deposit and are related to the tectonic areas of

sub-meridian trend. They are not detected often. These dikes are characterize by insignificant thickness (from 0.5 to 5-10 m), length (first dozen of meters and are not affected by hydrothermal

metasomatic changes.

2.1.3 Tectonics

Synclinal limb containing the deposit is tectonized by two systems of fractures. Thrust fault of northeast trending is the basic disjunctive structure and with a dip it moves east, southeast under 60-70

degrees. The thickness of the brecciation zone on the south flank of the deposit varies within 3-6 m.

The group of sub-parallel crushed zones of north-west trending is ore-bearing. Dip of these zones varies within 45-65 degrees to the east. Crushed zones were detected rather clearly by south and north

trending by mine opening, holes and geophysical methods.

Both fault systems are accompanied by numerous cleavages and shear cracks filled with quartz. The net of echelon small cracks determining the stockwork structure of ore body formed in the areas of

fault junctions.

2.2 Ore Body Characteristics

The main ore body and 7 smaller lenses are detected within the area of the Dombraly field oxide zone with the cutoff grade of gold being 0.3 g/t.

The main ore body includes 73.6% of the total gold reserves of the deposit, it is explored by drill holes

and underground mining and outlined on the surface alone the strike. It is not outlined by the slope, explored to the depth of 300 m. To the depth of 60 m it was explored by open mining in 1985-88. In

view plan, the ore body mostly have irregular lens form stretched to sub-meridian direction with sizes

on long axis up to 300-500 m, on short axis - 50-60 m. From the side of a hanging layer, lenses 1, 2 and 3 being its apophysis are connected to the ore body. From the north, the ore body sharply wedges

out between profiles 32 and 36. It is possible that this sharp wedge of the ore is associated with

dislocation of ore body by fault plane.

In section, the ore body has asymmetrical lens form and at the bottom it is split into several branches.

In the axial part of the ore body, there are the series of quartz veins and zones of stockwork

silicification that on the flanks of ore body also split.

The dip of the ore body is eastern, rather flat (60-45 degrees), decline is to the north under 20 degrees

in the oxide zone and 45 degrees on the lower horizons. Earlier explored ore bodies 1, apophysis and 2 with cutoff gold grade of 3.0 g/t are localized within the outline of the main ore body and are

controlled by the zones of stockwork silicification.

The Parameters of the main ore bodies of the Dombraly deposit with cut-off grades of gold 0.3 g/t are shown in Table 2.2.


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Body Ore Body Parameters Reserves

Explored length, m

Projected width in oxide zone , m

Average thickness

in oxide

zone, m

Ore, thous.

tons

Average gold

content,

g/t

Gold, kg

% from total

reserves

Main body 350 390 (90) 24.6 1191.0 3.77 4487.5 73.6

lens 1 165 73 4.5 79.1 1.00 79.2 1.3

lens 2 350 60 (45) 9.9 158.7 2.68 424.6 7.0

lens 3 330 90 (30) 8.1 228.9 2.53 578.3 9.5

lens 4 180 110 (30) 6.9 129.4 1.10 142.5 2.3

lens 5 70 180 3.4 83.1 2.49 206.8 3.4

lens 6 90 60 2.1 29.7 1.04 30.8 0.5

lens 7 100 130 4.4 127.6 1.19 151.5 2.5

Technogenic mineral formations (TMF) of the deposit are represented by overburden rocks: low grade

stockpile and re-cultivating mined rock in the open pit. The low grade stockpile is located 170 km south of the open pit on the southeast extension of gold ore mineralization zone with reserves of 1.7

million tons. The low grade stockpile is not registered in the state records. The industrial value of

TMF was determined by the geological exploration works carried out by Saga Creek Gold Company

in 2002-2006. The only commercial element in the low grade stockpile is gold. The area of the low grade stockpile base is 123095 m

2. The volume of the low grade stockpile is 1049.4 thousand m

3.

Re-cultivating mined rock fills the south part of the open pit. It has double deck structure. The south part of the open pit from profile 16 is filled completely to altitude of 218.6 m. The north part of the

open pit is filled partially due to the naturally occurring sloughing of the open pit sides. With regard to

the TMF reserves of the open pit it is rather small (1.06 million tons). The volume of the re-cultivating

mine rock is 612.9 thousand m3.

2.3 The Group of the Deposit Complexity

The deposit is classified as the third group of geological structure complexity according to The Classification of Reserves and Possible Resources of Solid Mineral Deposits (the RoK State Reserves

Committee, 2001).

The low grade stockpile of the deposit is classified as a second group (average complexity) on the

complexity of exploration. Methodical recommendations on analysis and evaluation of TMF (1995)

suggest air-drilled holes for the category С1 50 х 100 m. This low grade stockpile was explored by air-

drilled holes 40-60 x 50 taking into account high value of gold grade variation (224.1%).

2.4 Estimation of Forecast Resources

The increase of reserve can be achieved on the flanks and deep horizons of the deposit. Currently, this

area of southeast flank is taken for the dump and it is impossible to carry out geological exploration

works here. The forecast resources of Р1 category of the main ore body were estimated by the geological blocks. Outline of the Р1 category resources fixed to the C2 category reserves block. The

depth of resources evaluation was limited to 500 m from the surface. The average content of gold and

average horizontal thickness were determined on the basis of the data of sampling drillholes No.401,

402, 403. For the lens of ore 4 forecasted resources of Р1 category were estimated to the depth of 300 m. Determined on the basis of the data of sampling drillhole No.23. Forecast resources for Р1 category

are 3072.6 kg with average content of 6.02 g/t.

2.5 Geological Exploration Work Methods


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2.5.1. Survey Works

Survey of the Dombraly deposit was conducted to the scale of 1:1000. During this survey all old and

new geological exploration working were re-connected. Survey works were conducted by the

specialized division of the Karagandy Branch of Azimut Enegry Services JSC.

Division of the basic geodetic network to areas was done using nearest earlier set 3, 4 order state

triangulation points and the state leveling network. The total of 4 points were surveyed and used

further as the base ones for the next work stages.

Transfer of coordinates from WGS-84 system to a nominal coordinate system was done with Trimble

Geomatic Office, Geocalc software.

Field survey works were done with equipment of GPS Trimble4600 in PP-Kinematic mode. GPS

receiver Topcon GP-R1D was used as a base station. Processing of field survey data was done with

GPSurvey and Trimble Geomatic Office software with conversion to the nominal system of coordinates and to the Baltic Height System. Futher, this set of pickets was loaded to CREDO_TER

program for creation of the digital surface model.

In the zones closed from satellites electronic tachometer Topcon GTS-302 with field terminal FS-5

was used. Tacheometric survey processing was done in CREDO_DAT program.

Topographical drawings were prepared in CREDO_TER program with further export to AutoCAD

and MapInfor. Horizontals were drawn with the interval of 0.5 m.

2.5.2 Stages of the Deposit Exploration

The first stage of exploration and evaluation of oxide zone reserves of the Dombraly deposit was done

in 1969-1972. The program of works included topographic survey, preparation of regional and local maps, trenching, delving, mines and underground workings in horizons, core drilling and

technological sampling. Deposit exploration was done from the surface with ditches and bell pits

every 12-25 m, on the horizon of 30 m – by mine roadways and crossways from them every 20-25 m

and holes by 20x40 and 40x60-80 m grid down to 100-180 m.

The volumes of geological exploration works for this period are provided in the tables below:

Table 2.3

The volumes of geological exploration works for the period from 1969 to 1972

Type of works Meas.unit 1969-1972

Prospecting and exploration

drilling

r.m. 3785.6

Prospect mapping drilling r.m. 1840

Mechanical trenching r.m. 5416

Bell pits r.m. 232.8

Prospecting holes with entry of 9

m2

r.m. 32.2

Cutting crossway of 4 m2 720


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Table 2.4

Summary table on volumes of sampling by type of work

Types of work Meas.un. 1969-1972

Collecting core samples from holes r.m. 3732.8

Collecting pit samples from holes pcs 425

Collecting trench samples from underground workings r.m. 1954.7

Collecting trench samples from ditches r.m. 887.2

Collecting pit samples from ditches pcs 63

Reserves of the deposit oxide zone were evaluated to the depth of 170 m from the surface. Two ore

bodies were detected in the oxide zone. Reserves of the deposit oxide ores were approved by the

Territorial Reserves Committee of the Central Committee of the Production Geological Association (Minutes of Meeting No.3-411 dated 27.02.81).

The following reserves were approved with the cutoff grade of gold being 3.0 g/t:

for С1 category –87793 tons of ore, 949 kg of gold with average content being 10.8 g/t of gold.

In 2002-2005 geological exploration works at the deposit were conducted by Caga Creek Gold

Company. The main purpose of geological exploration was clarification of the lower border of the oxide zone, detection and evaluation of gold-sulphide mineralization, exploration of technogenic

mineral formations of the low grade stockpile and re-estimation of the oxide zone for detecting an

opportunity for processing the remaining oxide ores by heap leaching method. Moreover, additional exploration of poorly studied oxide zone flanks was carried out. Digital database including all old and

new data for geological-economical evaluation of the deposit oxide zone was created. Core drilling

with Longyear bullet was applied. There were 10 holes drilled with total volume of 2394 m.

Technogenic mineral formations were explored by air drilling and ditches.

Scope of works for 2002-2005 is shown in table 2.5.

Table 2.5

Scope of works 2002 – 2005

Type of works Site Meas.un. Volume

number Total

volume

Core drilling Open pit area m 10 2394

Air drilling Open pit flanks m 47 1004

Air drilling Open pit m 49 424

Air drilling Low grade stockpile m 110 1284

Ditches Low grade stockpile m3 3656.7

Trenching sample 1742

Slurry sampling sample 1361

Core sampling sample 1958

Technological sampling sample 3

Composite sampling sample 61

Metallurgical sampling sample 11

Pillars for determining bulk weight sample 5

On the basis of data from the first period of exploration and works of Saga Creek Gold Company in

2006, the RoK State Reserves Committee approved feasibility study of evaluation conditions (Minutes of Meeting No.496-06-K dated 21.03.2006). Total reserves (cutoff grade – 0.3 g/t) in re-cultivating


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mine rock in the outline of the open pit and in technogenic formations are: ore – 3797 thous.tons, gold

– 5629 kg, average gold content – 1.48 g/t. In 2007-2011 the deposit was explored again by drilling exploration holes and trenching.

2.5.3. Methods of the Deposit Study

In 1969-1972 the oxide zone was explored by trenches, pits with cuttings, mine with the range of

underground workings. All workings were planned in exploration lines oriented transversely to

potential length of the ore body.

From the mine with section of 9 m2, on the horizon of 30 m, the ore drive with section of 4 m2 was

made on the length of the main ore body in the south and north direction with total length of 220 m. From the ore drive transversely to ore bodies seven crossways were made of 4 m2 in section with the

length from 25 to 50 m for outlining ore bodies I and II every 25-50 m. Trenches were planned on the

area with thin layer of overburdens (0.5-3 m) laying on a clay-like weathered crust of the Paleozoic

ores. The depth of the trenches is 3.2-3.5 m with the width of 1 m. The distance between the trenches is 15-20 m and it was determined on the basis of the instruction requirements for exploration works on

the deposits of this type.

Photo 3. Drilling hole on the Dombraly site

Prospect drilling was done on profiles through 40-60 m under 80 degrees angle and magnetic azimuth

of 260 degrees transversely to the strike of ore-bearing structures and ore zones and downward by 40-

60-80 meters.

Drilling of core holes was done to the diameter of 127 mm to the depth of 5-6 m and further drilling

was done to the diameter of 108 mm to the depth of 50-60 m, and after that the hole had diameter of

90 mm to the bottom.

In 2002-2011 additional exploration of the oxide zone was done by air-drilled holes with direct

circulation, core holes with Longyear application and trenches. Exploration core drilling was done by stationary aggregates of drilling rigs ZIF-650 and SKB-5. The holes were drilled under the angle of 60


115

degrees. Drilling was done with diameter of 108 mm to the depth of up to 5 m. After installation of a

collar pipe the drilling continued by the double core equipment Longyear. The diameter of drilling was 76 mm, diameter of core was 40-42 mm. Average yield of core was not less than 90%. Core was

marked and stored to the standard boxes.

Air-drilled holes were designed for studying oxide gold mineralization in subsoil and in technogenic formations above ground water level. The diameter of drilling was 76 mm. Yield of slurry material

was 59-100%. For confirming the quality of air drilling in the technogenic formations on the dump

surface 8 trenches with the depth of 2.5 m were made in the holes profiles. Trenches were tested by the vertical trench samples with the length of 2 m. The results of comparing trench samples of the

walls of trenches and upper intervals of air drilling show satisfactory matching.

2.5.4. Geological-geophysical conditions of the deposit

Geological exploration of holes was done in 1969-72 with application of the following complex of geophysical methods:

electric logging in the exploration holes with the depth of 200 and more meters in modification of

resistivity logging, spontaneous polarization and electrode potential; measurement of natural radioactivity of the holes sections;

measurement of holes crookedness;

induced polarization logging; SP Exploration;

Electrical correlation method on constant current.

All works have been carried out in accordance with the requirements of instructions and projects. Measurement of holes crookedness was done by inclination compasses of ISh-4T, MK-2, UMI-25

type with the rise of point space being 10 m. Works by holes geophysics method were conducted using

VPS-63 station.

The oxide zone does not stand out in magnetic and gravitation fields. Ore zones stand out only visually

(hydrothermal-metasomatic changes) and by the data of sampling. In order to determine the thickness

of oxide zone, it is possible to use VES survey (and other modifications).

Down-hole survey has been carried out (inclination compass MI-36) with spacing of 20 m since 2005.

2.5.5 Sampling and Sample Handling

During the first stage of field works (1969-72) systematic sampling of mine workings was carried out

in the process of drilling those workings. Trench samples with the length from 0.1 to 1-1.2 m with

trench section of 5x10 cm were collected.

In the mine the samples were collected from all four walls, moreover, continuous vertical sectional

sampling (the length of section being 0.2-1 m) was carried out on the eastern and western walls and

horizontal trench samples were collected every 1-1.5 m vertically on the south and north walls. Detected quartz veils with thickness of more than 15 cm and their selvages were sampled separately.

In the roadways workings were sampled in 1-2 m of drilling by horizontal sections. In each section 2-3 trench samples with the length of 0.3-1.2 m were collected.

Collection of trench samples from ditches and bell-pits was done using the same method as from

underground workings and sampling pit (of mine) with the only difference that the samples from host rock without visible traces of hydrothermal changes on trenches and bell-pits were collected by pit

samples with the length of 3-5 meters.


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Sampling of the ore body in the open pit was done by trench method. Trenches were located transversely to the ore body each spaced 10 meters along the strike, the length of one sample varied

from 0.1 to 1.0 m with trench section of 6x8х10 cm.

Slurry sampling was done with the air-drilled holes. Only ¼ of the whole material was collected to the sample. Separation of samples was done in preventer.

In order to justify representativeness of slurry sampling of unconsolidated technogenic formations, the validation of upper interval of air-drilled holes sampling drilled at the dump by vertical trench samples

in the walls of ditches was carried out.

The relative value of a particular error is 25.42% and this is within the allowable limits. Comparison

of the sampling results of air-drilled holes and ditches applying the Student’s test surely proves non-

availability of a systematic error.

The trench sampling of technogenic formations of the dump was done on the wall of a ditch. The

section of the trench was 10x3 cm. The length of the section was 2 m. The location of the trench was

vertical.

Core sampling was done in core drilling holes. The length of samples was from 0.8 to 1.5 m (average

of 1 meter) depending on the lithological differences and the degree of hydrothermal changes. Core material of a sample was divided into halves by the long axis with a diamond saw. The weight of a

sample was approximately from 1.3 to 2.8 kg with core diameter being 40-42 mm. The samples were

sent to Kvarts Chemical Analytical Laboratory (CAL) for sample preparation. The rest of the core

material was stored in wooden boxes at the core warehouse of Sage Creek Gold Company (Stepnogorsk).

Two technological samples No.1 and No.2 (with weight of 600 kg each) from the underground workings from the horizon of 30 meters were collected in 1969-72 at the Dombraly deposit for

technological analysis on gold extraction from the ores and its material composition.

The parameters of technological samples are shown in table 2.6.

Table 2.6 Summary Table of Sampling Volumes

Types of Works Meas. un. 1969-1972 2002-2005

1 2 3 4

Collecting core samples from holes r. m. 3732.8 1958

Collecting pit samples from holes pcs 425

Collecting chip samples from air-drilled holes pcs 1361

Collecting trench samples from underground

workings

r. m. 1954.7

Collecting trench samples from ditches r. m. 887.2 1742

Collecting pit samples from ditches pcs 63

Collecting pillars pcs 3 5

Collecting composite samples pcs 5 61

Collecting samples for metallurgical sampling pcs 11

Field processing of a sample meant its crushing to pieces with the size up to 50 mm across and

packing it into boxes. Both samples were analyzed in the Central Laboratory of Kazzoloto trust.


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In 2004 the laboratory sample DLT-1 of 525 kg was collected from the open pit for determining the possibility of processing oxide ores by heap leaching method. In 2005 the additional laboratory sample

DLT-1 of 500 kg was collected for studying technological characteristics of the dump. Technological

analysis was carried out in Kazmekhanobr (Almaty).

For clarifying the lower border of the oxide zone and technological characteristics 11 samples of 1 up

to 2 kg (metallurgical sampling) were collected in 2005 in core holes. These samples were sent for

analysis to Kazmekhanobr. Material from the tailings of core samples processing was used for forming these samples.

In order to determine the bulk weight and moisture content in the ore of the Dombraly deposit, 35 paraffin-lined samples from underground working and holes were collected in 1969-72 and sent for

analysis to the laboratory of the Aksu deposit. Two monoliths with dimension of 50x50x50 cm were

taken for the same purpose from crossways III, IIIa and IVa and sampling pit of 1x2 m2 in section was

drilled in ored body I between ditches No.28 and 32a.

In 2004 pillars from re-cultivating mass and technogenic formations of the dump were collected. The

total number was 5 pillars from 56.8 to 61.8 kg.

Crushing of samples was done using mainly two approaches. First – irregularity coefficient in

Richards-Chechautte equation is equal to 0.2. This approach was applied to samples collected from deep prospecting holes, original weight of which did not exceed 0.5-1.5 kg, and samples collected

from weathered crust and from ores without visible ore mineralization with weight of up to 6 kg.

Samples collected from ore intervals, quartz veins, mineralized zones were processed using the approach with irregularity coefficient being 0.8 (with highly uneven distribution of components).

In 2002-2005 coefficient of 0.5 was accepted by analogy with operating gold fields of Kazakhstan oxide zones with uneven gold distribution (Central Mukur, Miyaly, Zhaima, Dzherek, Mizek, Uzboi,

etc.). The final diameter of crushing was 0.074 mm. The samples were processed in Kvarts Chemical

Analytical Laboratory.

2.5.6. Assay Works

Assay works were carried out in chemical laboratories of AKSU, Zholymbet and Tselinograd

Exploration Company in 1969-72.

The quality analysis of samples collected at the Dombraly deposit in 1969-72 was systematically

(twice a year) checked by the internal audit of the laboratory, where the main analyses were carried

out, by repeated analysis of coded duplicates of the original samples.

Internal audit of samples assay tests was carried out to the volume of 3-3.5% from the total number of

samples in the laboratories carrying out main assay works.

External audit (3-4%) of assay works of Tselinograd laboratory was carried out in the laboratory of

Aksu mine, and analyses of Aksu laboratory were checked in the laboratory of Tselinograd

Exploration Company. The functions of arbitrary control were imposed on the Central Laboratory of Kazzzoloto.

The total of 110 samples was sent for internal and external geological control.

In 2002-2005 the following types of assay works were carried out:

Assay of core, trench and chip samples for gold (3241 analyses);


118

Atomic-absorption analysis for gold of core and chip samples (2309 analyses);

combined samples analysis.

Assay test was done in Tsentrgeolanalit CJSC (Karaganda), atomic-absorption analysis was carried

out in Reaktiv LLP and Kvarts Chemical Analytical Laboratory.

Combined samples were analyzed in Tsentrgeolanalit CJSC. Assay tests for gold, atomic-absorption

analysis for copper, lead, zinc and silver, spectral assay for 16 elements, complete silicate analysis,

neutron activation analysis for platinum, chemical analysis for total sulphur, sulphide sulphur, sulphated sulphur were carried out on samples.

Internal geological verification of assay tests was conducted in Tsentrgeolanalit CJSC, external geological verification – in the laboratory of Kazmekhanobr. Samples collected from one duplicate of

prospecting samples were sent for internal and external verification. The total of 589 samples was sent

for both types of control compiling 5.1% of the total number of assay tests for this period. Assay tests

were not conducted and samples for geological verification were not collected during the second half of year 2005.

The results of external and internal geological verification of 1969-2006 show good reproducibility of results.

2.5.7. Determining bulk weight In 2004-2005 determination of a bulk weight and natural moisture of ore in situ was done on the basis

of paraffin-lined samples applying standard methods in the laboratory of Reaktiv LLP. The total

number of samples collected was 55, and 26 of them were the samples of oxide ore from middle and

lower horizons of oxide zone and 29 samples on gold sulphide ores.

It was determined that the bulk weight of oxide ores was 2.58 t/m3, for gold sulphide ores – 2.62 t/m

3.

Estimated average values of bulk masses are in good consistency with bulk weight dependency on depth diagram. Thus, for oxide ores, average value of bulk weight corresponds to the interval of

depths of 160-90 m, within the limit of which the principle reserves of oxide ore are located.

In 2004 pillars from re-cultivating mass and technogenic formations of dump were collected. The total number of 5 pillars with weight from 56.8 to 61.8 kg was collected. Weighting was done after samples

collection and drying in drying cabinets for 2 days.

Bulk weight of re-cultivating formations in the outline of open pit was 1.73 t/m

3, for technogenic

formations of the dump it varied from 1.58 to 1.77 t/m3 with the average of 1.66 t/m

3.

Due to the fact that the bulk weight was determined in preliminarily dried samples, adjustment to

moisture wasn’t applied to reserves calculation.

2.6. Ore material composition Two native varieties of gold-containing ores (oxide and primary) are allocated at the Dombraly

deposit.

Oxide zone ores are represented by brown, yellowish-brown, rarely by bright-red quartz-ferrous-

clayish saprolites on igneous-sedimentary rock.

Sulphide minerals are fully oxidized and leached. Oxide zone is characterized by the large amount of

secondary iron minerals: limonite, hematite, goethite, hydrogoethite and others. The ores in most cases

are pored and cavernous, it is usual for them to have the blebs of sulphide leaching out.

Material composition of oxide ores is shown in table 4.3.


119

Gold distribution in oxide ores is uneven. Gold content reaches 10-70 g/t in rod quartz veins and

silicification zones. In highly oxidized, disintegrated and ferruginized selvages the content of gold reaches 0.3-7.0 g/t.

Content of silver in oxide ores varies from traces to 5-10 g/t.

Besides silver, there is a small amount of zinc and arsenic in the oxide ores. Arsenic content reaches

0.15%, zinc – 0.34%. The leading positions in oxide zones are reserved for iron minerals:

hydrohematite, hydrogoethite, limonite, and less of jarosite. Gold nuggets are occurring in polished thin section. There are some barren minerals like quartz, barite, montmorillonoid, illite, sercite,

kaolinite.

The gold occurs in quartz (mostly) in host rock of oxide zone and in crystal lattice of pyrites in finely

dispersed condition. Relations of gold with other minerals were not detected.

The form of gold in quartz and in side weathered ores is characterized by a wide variety: sponge, vein-like, cloddy, dendrite-like, sheet-like, scaled grains adapting to intergranular space in quartz, in

microcracks, etc. Sponginess of gold grains is due to fine grain coarseness of quartz. The value of gold

discharge varies from 0.1 mm to 0.2-0.5 mm, the chemical composition of gold is noted for its changeability. Gold fineness variation limits is 800-950. The major part of gold (80% from the amount

of analyzed gold grains) is rated 800-825, the rest – 900-950 and even 990 (as per data of atomic-

absorption analysis (Central Geological Research Institute for Nonferrous and Precious Metals, 1970). There are no intermediate values between rates 800 and 900. In such manner, two types of gold (fine

gold and medium-quality gold (800-850) are recognized by the chemical composition. According to

the data of microspectral analysis, gold contains silver 5-20%, traces of lead and iron. The gold is of

coarse-grained structure (0.04-0.4 mm) with narrow penetration twins.

Gold-sulphide ores are characterized by impregnated, spotty, stringer-porphyry and pocket-

impregnated textures. Textures depend on replacement of brecciated sedimentary rock by quartz-feldspathic aggregare by filling space between fragments of fine-grained sulphide mass. Ore structure

is idiomorphic-grained formed by prismatic grains of arsenical pyrite and cubical grains of pyrite.

Sulphide content in ores varies from part of a percent to 15-18%.

The main ore minerals are pyrite and arsenical pyrite. There are affluent quantities of gold, galenite,

sphalerite, copper pyrite and hematite.

The main barren mineral is light-gray compact, fine-grained and anisometric quartz. Stripes of

potassic feldspar and impregnation of sulphides are frequent in quartz.

There are two mineral associations: early gold-quartz association and later gold-sulphide association.

In the latter, there are pyrite-arsenic pyrite and gold-sphalerite-chalcopyrite, late.

2.7. Natural and Commercial Types of Ores

There are two natural and commercial types – oxide gold-bearing and gold-sulphide ores at the

deposit. Oxide ores are developed from the surface to the depth of 60-140 m. The depth of oxide zone is primarily defined by intensiveness of tectonic processes in ores. Oxide zones on the flanks of the

deposit and in general on the territory of ore field is 60-80 m. The lower border of oxide zone of the

central part of the deposit goes down to 140 m (the area of profiles No.24-36).

Analysis on leaching out gold from the ore of the Dombraly deposit with column tests allowed making

the following conclusion:

the ore of the Dombraly deposit is suitable for heap leaching processing; expected commercial extraction of gold from granular ore of the Dombraly deposit will be from 45.18

to 74.60%;


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The ore sample DLT-2 of the Dombraly deposit related to sulphide gold-containing ores with

admixture of carbonaceous shales contain 6.5 g/t of gold and 7.7 g/t of silver, total sulphure – 1.74%, sulphide sulfur – 1.71%. Arsenic and carbon are the harmful impurities in the ore. On the basis of

rational analysis it is seen that the gold in the ore is mainly available in free form and in aggregates –

81.63%.

Testing on gravity washing of the ore was done on the concentration table KC-30. The results of the

gravity washing of ore of 65 grain size and 85% gr. – 0.074 mm show that gold can be extracted to

gravity concentrate. The results of chemical analysis show content of arsenic in gravity concentrate on the level of 3.5%, and stibium of < 0.005%.

Two alternative of flotation diagram were tested: with extraction of carbonaceous shales and further sulphide flotation and a regular sulphide flotation.

Ore flotation testing was done in laboratory conditions in a mechanic flotation machine.

Sorption leaching is the effective method of processing gold-containing carbon-bearing ores. Anionite

AM-2Bwas used for analysis of sorption leaching of gold from ore.

Gravity tailings of 65% grain size gr. -0.074 mm were additionally crushed to 85% gr. -0.074 mm for

carrying out tests on sorption leaching of gold.

The following tablulation shows extraction of gold from the ore by different methods:

Ore Washing Gold Recovery %

Gravity washing 53.45

Ore Flotation 86.13

Gravity-flotation of gravity tailings 85.25

Ore cyanide leaching 17.65

Sorption leaching 79.23

Sorption leaching of gravity tailings 64.87

Gravity – Sorption leaching of gravity tailings 83.65

Sorption leaching of gravity concentrate 92.37

Gravity – Sorption leaching of gold from gravity tailings with account of processing

gravity concentrate by Sorption leaching method

79.57

Different alternatives of ore washing were tested in the course of analysis. By gravitation method,

53.45% of gold is extracted to gravity concentrate from the ore, with gravity concentrate yield being

3.53% and gold content in it – 61.5 g/t. By flotation method, 86.13% of gold is extracted from the ore. Flotation concentrate yield is 13.59% with gold content in it – 27.4 g/t.

Direct leaching of ore allowed extracting only 17.65% gold to cyanide solution due to availability of carbonaceous shales being the natural sorbents in the ore.

Sorption leaching extracts 79.23% of gold to resin from the ore and 64.81% of gravity tailings.

Combined methods of ore washing ensured gold extraction: gravity – Sorption leaching – 83.65%;

gravity – sorption leaching of gold from gravity tailings with processing of gravity concentrate by

sorption leaching – 79.57%.

Gold-containing ore DLT-2 of the Dombraly deposit can be washed by two methods – ore flotation

and sorption leaching.


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2.8. Flowsheet for processing oxide ores Flowsheet for processing oxide ores was prepared on the basis of results of technological survey of ore

(sample DLT-1) and tests of technogenic formations of the dump of the Dombraly deposit (sample

DLT-1 additional).

Process flow diagram: two-stage crushing with screening before the second stage;

agglomeration using Portland cement;

heaping sintered ore by radial stacker; leaching out gold from the ore by cyanide solution;

sorption of solved gold by resin with output of concentrated sorbent and barren solution, to which

required reagents are added for further return to processing; desorption of gold from resin, electrodeposition of gold from desorbent;

preparing Dore gold.

The output commercial product of the process is a gold-containing Dore bar, base gold of TS 98 RK-1-93 sent to a gold refinery for metal separation.

2.9. Hydrogeological conditions

Engineering-gological, hydrogeological and environmental surveys at the Dombraly deposit were

conducted by Karagandy Branch of Azimut Energy Services JSC under the agreement with JV Saga Creek Gold Company LLP in 2005.

The landscape of the deposit is an elevated hilly-undulating plain. On the top this plain is formed by

thin cove of the Quarternary eluvial-diluvial formations. Weathered crust is distributed widely across the site. It consists of loamy, rubbly-loamy and rubbly parts with the regular increase of fragmentary

material content in deeper parts and transfer to fractured (exogenous) zone. The total thickness of the

weathered crust varies from 20 to 60 m on area and up to 120 m in the zones of tectonic disruptions. Water in the weathered crust is related to rubble and grus material.

Besides availability of the weathered crust, lithology of water-bearing masses defining the nature of

rock jointing, filtration properties, quantity and composition of soluble salts released during weathering process also influences the quality and quantity of underground waters. Water occurs in

upper rock jointing of the Paleozoic formations that often combine with the interstitial waters of

weathered crust into one hydraulic system mixing by circulation and, in such manner, become the unified complex of weak pressure interstitial water. The depth of water circulation is defined by the

depth of cracks and the nature of rock jointing. In effusive rock, jointing spreads to the depth of 30-35

m, and in sand-rock – up to 50 m. Static levels of underground water occur on the depth from 15 to 26 m depending on hypsometric holes location. In general water content in the rocks is low. Maximal

flow rate is noted in the holes located within tectonic zones and it is 0.8-1.0 l/s with corresponding

lowering to 17.2-9.5 m and it is 3-8 l/s with lowering to 8.7-8.5. On the basis of pumping-out data,

rock filtering coefficient varies from 0.06 m/day (hole 3, 29) to 0.54 m/day (hole 27, 28), average – 0.06-0.20 m/day. In holes 16э and 35э pumping out flow rate was correspondingly – 8 and 3 l/s, with

lowering to 8.5 and 8.7 m filtering coefficient was 4.63 and 2.4 m/day. The holes are located in the

tectonic disruption zones and that’s why such values of coefficient are not representative of the site and are not taken into consideration during water inflow estimation.

Underground waters mostly feed from precipitation. Feed areas are located outside the work site on the elevated areas, because bottoms of liars and sais are formed by rather thick clayish formations with

weak water filtering capacity. Local underground water-flow goes on upper zone of rock fracture,

regional underground water-stock – on large areas of tectonic disruptions. Underground water flows in

local base level of erosion – Kop’s tract – located 3 km from the open pit.


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Regarding the level of mineralization, ground waters being of sporadic distribution are slightly salty

(1.3-2.2 g/dm3). Underground water of rock fracture zones is fresh with mineralization from 0.4 g/dm3 (hole 27) to 0.8 g/dm3 (hole 35) and slightly salty with mineralization of 1.4 g/dm3 (hole 29).

By chemical composition, fresh water is hyrdocarbonate, chloride carbonate and hydrocarbonate-

chloride calcic-and magnesium-sodium with total hardness of 3.8-5.8 mg-eq, neutral. Slightly salty water is hydrocarbonate- and sulphate-chloride magnesium-sodium and sodium, with total hardness of

7.9-1.10 mg-eq, neutral.

Spectral and X-ray spectrum analyses in dry residue of water and chemical analysis of water did not

detect excess of any chemical elements (according to Maximum Allowable Concentration for potable

water in terms of total mineralization).

2.10. Mining and geological conditions

Geological survey works were conducted at the deposit. Design phase – engineering documentation. The program of geological survey works provided drilling of 24 holes (No.1-24) with the depth from 6

to 70 m. During the process of field works, additional holes 1a and 21a were drilled for more detailed

analysis of ground in the places of intense landscape change, where there was the probability of geological structure change.

Holes were drilled by spud method by UGB-50M machine, d=132 mm, without hole casing. Holes 19 and 20 were drilled by core drilling with URB 2A-2 machine, d=132 mm, to the depth of 70 m. Total

metreage of drilled holes was 301.5 running meters. Monoliths and samples were collected from all

holes.

Prepared the map of filtration fields of heap leaching area. Analysis showed that the main section of

the area was formed by weakly permeable grounds (Coef.=0.005-0.3 m/day). Clays with filtering

coefficient being <0.005 m/day (water-resistant) are of sporadic distribution in plan view, and in depth. This is why it will be required to lay additional hydro insulating layers during construction of

PKV.

Topsoil is available almost everywhere. Its thickness varies from 0.15 to 0.3 m. Humus content as per lab analyses data is 0.7-7.7%.

All field, laboratory and data processing work was done in accordance with the regulatory documents effective on the territory of the Republic of Kazakhstan (the RoK SNiP (Construction Rules and

Norms) 1.02-18-2004; the RoK SNiP 5.01-01-2002; SNiP 1.02.07-87; SNiP 2.02.01-83*; SNiP

2.03.11-85; the RoK SNiP 2.04-01-2001; SNiP 2.01.07-85*; GOST 25100-95; GOST 9.602-89; "Guidelines for Designing the Foundations of Buildings and Constructions” to SNiP 2.02.01-83; SNiP

IV-2-82, etc. ).

Soil characterization was done in accordance with table Б.1=3 of GOST 25100-95.

As per the conditions of drilling in soil, according to SNiP IV-2-82, topsoil is related to group I-II,

Quartenary clay foam – group II-IV, sand clay and clay foam – II-IV (§ 33g), clays – II-IV (§ 8d), rubbly soil –groups IV-V (§ 31а), argillites – group V (§ 31b) depending on the method of

development.

2.11. Methods of opening and field development

The method of the oxide zone development – open pit mining.

The borders of open pit mining are set with account to inclusion to mining of all commercial reserves

of the oxide zone.


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For open pit mining to the horizon of 90 m, opening by internal entry ditch that transforms into stationary transport access tracks on the way down is provided.

Slopes are under the angle of 65 degrees, open pit walls – 51 degree, the width of berms is 5 m, of

ditches and access tracks – 15 m.

Development system – transport system by descending horizontal layers with transporting overburden

to external dump, ores – to ore site and partially to the batches of agglomeration complex.

In near-contact zones, in order to reduce losses and dilution, selective mineral extraction with storing

offgrade ore in out of balance ore dumps is provided.

Ore and overburden rock hardness coefficient as per the classification of M. M. Protodiyakonova is 8-

15 (SNiP group – VII-IX), blastability grade – V.

2.12. Method of Reserves Estimation

Reserves estimation was carried out in 2008 by the method of vertical geological open cuts separately for oxide and primary gold-containing ores, re-cultivating mass within outline of open pit and

technogenic formations of the low grade stockpile.

Geological peculiarities of the deposit, methods of its exploration and development were taken into

account when choosing a method for reserves estimation. The following was aiding to estimation:

drilled holes and workings located on 10 profiles in latitudinal direction (53 holes, 19 bell-pits, 8

ditches, 24 crossways, 206 air-drilled holes, 61 line of trench sampling, 7089 samples). Interval between profiles is 40 m, except the interval between profiles 0-7 (80 m). Directly for estimating

reserves were used 40 holes, 3 ditches, 4 crossways, 17 lines of trench sampling, 163 air-drilled holes

and 1422 samples, with which the reserves of ore were blocked.

Estimation of oxide ore reserves was done using evaluation conditions approved by the RoK State

Reserves Committee (Minutes of Meeting No.496-06-K dated 21.03.2006):

cutoff gold grade in sample – 0.3 g/t; minimal thickness of ore bodies included in reserves estimation – 2 m; with lesser thickness, but high

gold content the appropriate GT should be used;

maximal thickness of barren rock layers or offgrade ores used in reserves estimation – 4 m; estimation of poor ore reserves being re-cultivating rock within the outline of earlier developed open

pit should be done by total mass.

Nominal conditions accepted by analogy to evaluation conditions of primary gold-containing ores of the Uzboi deposit (underground mining) approved by the RoK State Reserves Committee were used

for estimating reserves.

Outlining oxide ore bodies was done on the basis of the data of ditches, mines, roadways, crosscuts, sampling pits, actual mining horizons and holes to the depth of oxide zone. The lower border of the

oxide zone was drawn upon detection in the rocks of carbonaceous material relics and first traces of

sulphide. Border of the oxide zone is shown on a layout mainly on the basis of primary documentation data. Outlines of ore bodies were reconstructed on estimated section of 1:500 scale and assay plans of

1:500 scale. Ore bodies border were drawn through a final sample with cutoff gold grade of this

alternative. Between ore and barren roadways the outline was drawn on the half of distance between them, but not more than 25 m. In case of non-availability of outlining roadways and thickness of ore

intersection more than 2 m, the outline was drawn to the distance of 25 m downwards (along the

strike), with thickness 2 m and less – to the distance of 12.5 m.


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Estimation of re-cultivating rock reserves within the outline of the open pit and technogenic

formations of the low grade stockpile was done by the method of vertical sections. Reserves are classified under C2 category.

The main exploration minings were air-drilled holes. Representativeness of chip sampling of

technogenic formations was confirmed by trench sampling of ditch walls. Sampling was done only of the upper above-water part in re-cultivating formations of the open pit. The results of holes sampling

are considered to the whole volume of re-cultivating formations. The bulk weight of re-cultivating

formations was 1.73 t/m3, for technogenic formations of the low grade stockpile – 1.66 t/m3. 3. General Information of the Shirotnaia Deposit

The Shirotnaia deposit is located in the southeast part of Alhambra Resources licensed site, 8 km north from Stepnogorsk and 100 km from its operating open pit – Uzboy

.

Photos 4 and 5. Landscape of the Shirotnaia deposit

3.1. Geological Structure of the Shirotnaia Deposit

Volcanic formations of the Mid-Ordovician period and Quarternary alluvial formations form the

structure of the deposit. Mid-Ordovician is represented by volcanic-sediment deposits, mainly by tuffs of average compostion

with lens-shaped bodies of diorites of upper Ordovician age, quartz andesite, volcanic brecciates,

siltstones, fine-grained sandstone with rare layers and lenses of limestone, tuffs and quartzrock bodies. Rocks of the Mailysor late-Ordovician sub-volcanic complex represented by andesite-dacitic

porphyrites are quite frequent.

Ore mineralization is located in the Aksu zone of disruptions of north-east trending with 45о

trend azimuth and is clearly seen on the joining section and intersection of splits of different orientations.

3.2 Methods of Geological Exploration Works

In 2002-2004 in the north-western part of the deposit area, ditches in profiles every 40-80 m were

drilled, as well as 5 profiles of air-drilled holes with drilling step – 5 m. On the basis of the results of these works, 4 ore bodies were detected with cutoff gold grade of 0.5 g/t and the range of small lenses.

The length of ore bodies was 240-540 m, and thickness from 1 to 15-20 m. Gold reserves were

estimated to the depth of 30 m and amounted to 1500 kg with average content of 1.62 g/t.

In 2005 prospect-evaluation survey was carried out in south-east part of the deposit area within a

swamped valley. Works were conducted by digging ditches and drilling RCC holes. Ditches were

done in profiles every 40 m, their depth did not exceed 1.5 m due to high level of underground water. The volume of ditches digging was 2097 m, the number of collected trench samples – 2023. RCC


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holes were drilled in the lines of ditches profiles by grid 80x10-20 m to coming to transition zone

rocks. The volume of drilling was 3797 m, the number of collected core samples – 1864. The depth of the oxide zone was 30 m.

As a results of 2005 works, large sub-lateral ore body of complex morphology. Along the strike it was

traced for 700 m. In the central part this ore body’s thickness increases to 97 m to the length of 160 m,

on the flanks the thickness of ore body decresed to 2-9 m. Ore body is inclined 60-75 to the north-west. Three band-like mineralization zones with thickness of up to 6-10 m were detected on the

bottom layer close to the main ore body.

In 2006 prospecting and evaluation works were continued on the south-west flank of mineralization

and its central part. Ditches were drilled on the south-west flank every 40 m, their average depth –

1.8m. Volume was 3845 m, the number of collected samples -3711.

RCC holes were drilled in the hanging wall and bottom of the main ore body by grid 120-80x10 m to

exit to solid rock. Drilling volume was 197 holes with the depth of 27.6 m, total length of 5443.5 m,

total of 2083 samples was collected.

As a result of these works, earlier stated reserved of the central part of mineralization and its south-

west were confirmed. On the south-east flank 520 m away from the main body, columnar

mineralization with north-west inclination to 50-450 was discovered.

In 2007 main prospecting works and depth exploration of the main ore body were conducted.

The purpose of the prospecting works was to discover gold mineralization zoned to the depth of 4 m

on the north-east and south-west flanks. Deep-earth geochemical survey of the territory of 27.7 km2

by grid of 200x50-20 was conducted here. During this period 818 small-depth holes with the length of 3760 m, 1120 samples were collected.

Exploration of the main ore body and south-west flank of the mineralization was done by core holes to

the depth of 120-140 m. There were 18 holes drilled to the length of 2117 m, 2001 sample was collected.

Prospective areas on the north-east flank were discovered on the basis of prospecting works of 2007. Drilled holes opened columnar main body on the depth within 100x80 m with north-west inclination

to 60-7- degrees.

In 2008 works were carried out in three stages:

First stage – continued prospecting works on the east, west and south-west flanks by deep-earth

geochemical logging by grid 500-1000x50 m on the area of 45 square meters. There were 1893 holes drilled with total length of 8723 ml 2995 samples were collected. On the basis of these works’ results

secondary haloes of gold were built.

Second stage – conducted drilling works on secondary gold haloes. The latter were traced to the

distance of 1800 m from the main ore body to north-west by grid of 200x10 m, drilling volume was

20891 m, 10449 samples were collected.

Third stage – exploration works (air drilling) was carried in discovered zones and ore bodies by grid of

40x5 m along the strike to the distance of 500 m. The volume of drilling was 7218 m, collected 3609

samples.


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Photos 6 and 7. Trench in the stream valley. Head of air-drilled hole in 2011.

From 2009 and up until now additional exploration of the deposit on its flanks has been done. There

are no results of drilling works for this year.

Photos 8 and 9. Geodesic marker. Old ditch on the south part of the deposit.

Approximate reserves estimation on the oxide zone of the deposit on С2 category was conducted by

the Geology Department of the company and is (as of 01.01.2010): ore – 3250 thous. t. ;

gold – 3900 kg;

average gold content – 1.2 g/t. Oxide ores on P2 category:

ore – 33 300 thous.t.;

gold- 40 000 kg;

average content- 1.2 g/t.

4. Visiting the Dombraly and Shirotnaia deposits

Geological exploration works on the deposits are carried out according to the conventional Soviet

geological methods. Documentation of the mine workings and holes is carried out directly at the mine sites. Sampling and processing samples are conducted on the territory of the Saga Creek Gold

Company’s gold processing plant.


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4.1 Core documentation

Core documentation is carried out during the core drilling process on a daily basis. Geological information records are entered into the documentation logs with compulsory requirement to fill in the

table form with drilling intervals (runs), fixing the core length, % core yield, core sketches (separate

fragments), the geological characteristics of the core by the layers of rocks and ores, sampling

intervals and sample numbers. During performing geological task of the hole, the hole is closed by the district geologist, and check measurement of the hole is taken.

After drilling is finished, Case (Passport) for the hole is composed, which includes geological documentation logs, drilling logs (conducted by the drilling foreman with filling in the drilling

parameters and possible geology-technological complications), acts of initiation and completion of

hole drilling, the act of survey measurement, the act of geological research (if conditioned by the project), the geologic column for the hole. Photos of core holes are taken in the position “box on the

ground”. Hole documentation on the area of the deposits and fields is not produced in electronic form.

4.2. Core cutting Following the geological documentation and the selection of ore zones of mineralization, core holes

are directed to the cutting along the long axis of the core. Only those ore intervals visually highlighted

by the geologist are subject to the cutting.

Core cutting is made on the machine produced in Russia under the supervision of geologist, on the

territory of the Stepnogorskaya Mining Enterprise, a subcontractor of the drilling works. Diameter of diamond disc is 400 mm.

Photos 10 and 11. Cutting machine and core receiver

Core is split into two halves. Detrital material out of cleavage zones with the fragment size smaller

than the diameter of the core is split in half by hand. Following the cutting, core boxes filled with the sample material are delivered to the Saga Creek Gold Company’s core storage.

4.3 Sample preparation

Core boxes are laid out in a special order on the paved area of the core storage.

Geologist marks sample intervals, putting sample labels along sample borders that were strictly

identified by the documentation. Sampler selects samples into the plastic sacks strictly according to the sample intervals by selecting core halves. Paper label with sample information (site, hole No.,

sample No., sample interval, Family name of the geologist, date of sampling) folds and fits into the

sample sack. Hole and sample No. are signed on the sample sack. Selected samples of the holes are delivered to the core storage, where sample weighing and sample group formations for the sample

preparation are performed.


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Photos 12 and 13. Core prepared for sampling and sample group formation

Selected samples are weighed on the scale of the РН-10Ц134 make, with a scale division of 0.05 g and

are entered in the sample registration log.

Photos 14 and 15 . Sample weighing and selecting batch of samples

Here, samples are distributed under orders to carry out analytical work.

The second halves of the core are neatly stacked in core boxes. Boxes for the drill holes with the

remaining core are closed and transported to the core storehouse.

Later, if necessary, laboratory-technological samples would be selected out of the core remnants for

the study of mineral and technological ore properties.

Samples selected by the program of exploration works have been processed in crushing mill workshop of Saga Creek Gold Company’s gold processing plant in Stepnogorsk since 2009. Sample processing

is performed using the Richard-Chechett' formula Q = kd2 at k = 0,5. Experimental validation of the

coefficient of uneven mineralization was carried out on the field in 2005. Previously, samples were processed in other companies. In 2002-2004, the sample processing was carried out in the Reagent

LLP (Stepnogorsk), in 2005-2009 – in Quartz Chemical Analysis Laboratory . Over the entire period

of gold mine exploration grinding of the gold samples was carried out at K equal to 0.5. Scheme of the core sample processing is shown in the figure. Subsequent to the drying samples are sent to the first

stage of the jaw crusher, where sample material was grinded up to 7 mm. Following the first stage

of grinding crushed material is sent to roller crusher, where it is grinded up to 1 mm.


129

Figure 2. Sample processing scheme

Sample rescreening is performed following the roll crusher through the sieve with a cell of 1 mm.

Material not passed through the sieve returns to regrinding in the roll crusher.


130

Photos 16 and 17. Jaw crusher of the VIBROTECHNIK company and disc eraser of Russian

production After the second phase of fragmentation and reduction by the Jones index the bulk of crushed to 1

mm samples enters the geological sample storage as a geological sample duplicate,

while sample itself weighing about 0.75 kg is directed to reduce to a particle of size 200 mesh (disk eraser).

Batches for the production of atomic absorption and assays are selected from the sample material, and

so are the batches for the internal and external control production. Remains of analytical sample are sent to the sample storage as analytical duplicates. Analytical samples themselves are

grouped in separate orders and are sent to the Stewart Group laboratory (Kara-Balta, Kyrgyzstan).

All instruments used for crushing and sample reduction are equipped with special instructions for

operators. The workshop of the sample preparation is kept clean. After every sample preparation all appliances and countertops are cleaned and blown by compressed air without fail. Cleaning each

batch of samples with inert material (granite, glass or any other dead rock) is not performed after

crushing.

4.4.Storage of the geological sample and core duplicates

Following sawing and hole sampling boxes with core are sent to the core storage. Core holes not

penetrated the ore body are kept close to the core storage. Sample duplicates are stored in a specially designed sample storage building of the gold processing plant.

Photos 18 and 19. Open platform of the core storage


131

All duplicate samples are strictly controlled. Geological sample duplicates are packed in sacks, sacks

are signed, stacked by holes, and are kept on special shelves. All samples are numbered and easily accessible.

Photos 20 and 21. Storing geological sample and core duplicates inside core storage.


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CONCLUSION AND RECOMMENDATIONS

The Dombraly Deposit was studied with surface ditches, pipes, pits and deep exploratory holes and

underground mine workings, passed in the horizon 30 m below the surface. The upper part of the

oxidation zone of an open pit has been processed to a depth of 60 m. The pit is partially reclaimed. Material composition and technological properties of oxidized and gold-sulfide ore deposits as well as

technogenic structures of Dombraly II deposit were studied by laboratory-technological samples and

samples for technological mapping. Hydrogeological and geotechnical conditions of deposit

development are simple. Reserve estimation of oxidized and gold-containing ores in mineral resources, gold technogenic

mineral formations and dump recultivation rock in the open pit path was calculated. Delineation of

oxide ore was produced in the bowels of evaluative condition, approved by the SRC RK (Minutes №

496-06-K on 21/3/2006 ), gold-sulfide ore - contingent condition, taken by analogy with an estimated body condition for sulfide ores for underground mining of the Uzboy deposit. Stocks of recultivated

mass in the open pit path and TMO are estimated without cut-off grade of gold. The total reserves in

the bowels of the oxidized ore, recultivated rock mass in the open pit path and technogenic deposits on the blade on С2 category were: 4254.1 thousand tons of ore, 6478.8 kg of gold, the average gold

content of 1.52 g / t. Stocks gold-sulfide ore to a depth of 300 meters on C 2 category were: 575.6

thousand tons of ore, 3119.6 kg of gold, the average gold content of 5.42 g / t. Forecast resources of category P 1 to a depth of 500 m: 510.1 thousand tons of ore, 3072.6 kg of gold,

the average gold content of 6.02 g / t.

In general, for the deposit of С2+Р1: 5339.8 tons of ore, 12671.0 kg of gold, the average gold content

of 2.37 g / t. Shirotnaia deposit is studied mainly within the development of oxidation zone. It established two

zones of mineralization, including some of the ore bodies and lenses. Exploration network of traversed

holes in the deposit will evaluate the deposit according to the C 2 + P 1 category. The deposit is very kindly, as evidenced by the estimated valuation of resources in the zone of oxidation by category

C 2 and P 2, performed by the Geological Survey of Saga Creek Gold Company LLP. The deposit

requires exploration to be continued in the upper levels for category C1 stock for oxidized ores and

exploration of primary gold-sulfide ores. Exploration works are carried out by qualified personnel in accordance with existing procedures.

Additionally, recommendations are as follows:

Keep the mouth of the exploratory open pits by leaving a pipe of 0.7-1.0m length above the surface with a designated hole number and the date of drilling hole;

Conduct geological hole documentation in electronic form (Word), without giving up the log records

documentation; Make a design for hole photography at 45

0;

Cut out ¼ part of the core of each 20th ore sample for quality control of samples and determine the

probable non-uniformity of mineralization (the direction of a reference laboratory);

Apply the "single, obviously empty" samples, along with certified samples.

Author: Evgeniy Mikhailovich Zhuravlyov,

Senior Geologist, Micromine Kazakhstan LLP


133

APPENDIX 2. DOMBRALY PROJECT

DATABASE LISTING AND

VALIDATION REPORTS


134

2011 Resource Estimation Input Data Listing

P18785 October 2011 Dombraly Resource Estimate Data Listing - 15/09/2011

Datatype Source File Name MM File Name Records Comments QAQC

Drill Data Alhambra_Dombraly_23032011.pro

ALL_COLLAR_01052008 ALL_COLLAR_01052008

744 (compilation of all pre-2006 exploration data); • 10 C* DDH holes (2005) • 110 W* RAB holes (2002) • 49 #* RAB holes (2002) • 47 ER* RAB holes (2002) • 8 EW-* low grade stockpile trenches (2005) • 2 EW* low grade stockpile trenches (2005) • 17 Nkv* old crosscut trenches/channels • 17 Skv* old crosscut trenches/channels • 14 #* old DDH holes • 1 C * old DDH holes • 4 C* old DDH holes • 187 C-* old DDH holes • 24 D-* old DDH holes • 19 k-* old trenches • 5 k* old trenches • 212 P* old trenches • 18 P * old trenches

Y - Dombraly QC Fire Assay - Partial for 47 ER* RAB, 8 EW-* and 2 EW* low grade stockpile

ALL_SURVEY_01052008 ALL_SURVEY_01052008 1742

ALL_ASSAY_01052008 ALL_ASSAY_01052008 17331

ALL_GEOLOGY_06052006 ALL_GEOLOGY_06052006 121 10 C* DDH lithology (english long hand)

Dombraly DDH DB Weathering.xls ALL_WEATH 16 Weathering data for 10 C* holes

Alhambra_Dombraly_23032011.pro

2003_WP_PP_pit samples_coordinates 2003_WP_PP_pit samples_coordinates

96 Saga Creek 2003 low grade stockpile and pit infill sample pits

N

2003_WP_PP_pit samples_survey 2003_WP_PP_pit samples_survey 96

2003_WP_PP_pit samples_assay 2003_WP_PP_pit samples_assay 96

Lithology?

Weathering?


135


2006 RAB coordinates 2006 RAB coordinates 105 Saga Creek 2006 pit and low grade stockpile RAB N

2006 RAB survey 2006 RAB survey 210

2006 RAB assay 2006 RAB assay 676

Lithology?

Weathering?


2010_RC_Collar 2010_RC_Collar

37 Sage Creek 2010 pit and low grade stockpile shallow RC

Y- Dombraly 2010 RC QAQC

2010_RC_Survey

2010_RC_Assay 2010_RC_Assay 880

Lithology?

Weathering?

Dombraly additional Data.zip (9 sep 2011)

2010_DDH_Collar.xls 2010_DDH_Collar.dat 13 Saga Creek 2010 DDD* diamond holes Y - Dombraly 2010 DDH QAQC

2010_DDH_Survey.xls 2010_DDH_Survey.dat 358

2010_DDH_Assay.xls 2010_DDH_Assay.dat 2877

Dombraly 2010 DDH logging.xls DDD lithology in Russian/Kazakh

Dombraly DDH DB Weathering.xls 2010_DDH_Weath.dat 30 Weathering data for 12 DDD* holes (1 hole in all oxide zone)

Trench Data Alhambra_Dombraly_23032011.pro

ALL_COLLAR_01052008 ALL_COLLAR_01052008

744 (compilation of all pre-2006 exploration data); • 10 C* DDH holes (2005) • 110 W* RAB holes (2002) • 49 #* RAB holes (2002) • 47 ER* RAB holes (2002) • 8 EW-* low grade stockpile trenches (2005) • 2 EW* low grade stockpile trenches (2005) • 17 Nkv* old crosscut trenches/channels

Partial - for 47 ER* RAB, 8 EW-* and 2 EW* low grade stockpile Y - trenches - Dombraly QC Fire

Assay


136

• 17 Skv* old crosscut trenches/channels • 14 #* old DDH holes • 1 C * old DDH holes • 4 C* old DDH holes • 187 C-* old DDH holes • 24 D-* old DDH holes • 19 k-* old trenches • 5 k* old trenches • 212 P* old trenches • 18 P * old trenches

ALL_SURVEY_01052008 ALL_SURVEY_01052008 1742

ALL_ASSAY_01052008 ALL_ASSAY_01052008 17331

ALL_GEOLOGY_06052006 ALL_GEOLOGY_06052006 121 10 C* DDH lithology (english long hand)

Weathering?

Bulk Density

SG Core.xls ALL_C_SG_Core.dat 55 Old C* DDH SG data

New DDH SG Core? No density for new 2010 DDD* holes

QAQC QA-QC.zip

Shirotnaia & Dombraly QC Fire Assay.xls Dombraly QC Fire Assay 96 Old (2004-2005) Dombraly field duplicate and external QC data (4 for ER* RAB, 92 for EW-* and

EW* low grade stockpile trenches)

4 for ER* RAB, 92 for EW-* and EW* low grade stockpile

trenches

Dombraly 2010 RC QAQC New Dombraly RC QC data (RCW* and RCD* holes) 39 blank, 41 standard, 40 field dup, 151 lab dup

Dombraly 2010 DDH QAQC New Dombraly DDH QC data (DDD* holes) 128 blank, 126 standard, 132 field dup, 403 lab dup

GIS Alhambra_Dombraly_23032011.pro

Dombraly_Topo_07 Dombraly_Topo_07 Topo DTM wireframe with pit infill

Dombraly_Topo_OLD Dombraly_Topo_OLD Topo DTM wireframe without pit infill

Dombraly additional Data.zip (9 sep 2011)

Contours.tab Contours_100911.str 0.5m topo and pit contours

Measurement points.tab Measurement points_100911.dat topo survey points


137

Measurement results.tab Measurement results_100911.dat sames as points file above

Situatuon contours.tab Situation contours_100911.str clay pits, water areas polygons

Situatuon lines.tab Situation lines_100911.str Same as situation countour file

Slope.tab Slope_100911.str Pit, low grade stockpile(?) interp layer (no RL)

Steeps.tab Steeps_100911.str Pit, low grade stockpile(?) detail/gully interp layer

(no RL)

Triangulation stations.tab Triangulation stations_100911.dat Survey station location layer

Interps Waste Dump_Estimate Waste Dump_Estimate Waste Dump_Estimate solid Wireframe

ALL_DDH_AU_0_18.str ALL_DDH_AU_0_18.str Pit area in situ and backfill Au min polygons x section interp from DDH

TRENCH_CROSS_AU_0_25.str TRENCH_CROSS_AU_0_25.str Pit area in situ and backfill Au min polygons plan interp from crosscut data

2011 Dombraly Validation Reports

Dombraly DHDB and Trench DB Validation Report Tables Sep 2011

File Record Hole ID From To Warning Action Comment Resolved? Dombraly_150508.dhdb

ALL_SURVEY_01052008.DAT various various Hole deviation > 0.10 degrees/metre Check ‐ ok? may be valid change OK

ALL_GEOLOGY_06052006.DAT various various Various Holes not defined Check ‐ ok? No Geology for these holes OK

2006_RAB_230311.dhdb

No errors detected

2010_RC_300811.dhdb

2010_RC_Assay.DAT 824 RCD1405 71 74 Missing interval check Input if possible Y ‐ total depth 71m corrected in collar file

2010_DDH_100911.dhdb


138

2010 DDH Survey.DAT 102 DDD0703 Hole deviation > 0.10 degrees/metre check OK







2010 DDH Assay.DAT 1662 DDD4803 341.7 343 Overlapping intervals check error Input error - corrected to 1m interval 341.7-342.7


139

APPENDIX 3. Dombraly do main top cut

stats and graphs


140

Mineralised low grade stockpile – Lower Top Cut @ 10g/t Au

Mineralised low grade stockpile – UpperNo top cut


141

Pit Infill Zone Top cut @ 5g/t Au

In Situ Domain A Top Cut @ 10g/t Au


142

In situ Domain B No top cut

In situ domain C No top cut


143

In situ domain D No top cut

In situ domain E No top cut In situ domain F Top cut @ 6g/t Au

In situ domain G No top cut

In situ domain X No top cut

In situ domain Y No top cut

In situ domain Z No top cut


144

In situ domain Z No top cut


145

APPENDIX 4. Dombraly Validation Cross

Sect ions


146

Mineralised low grade stockpile – Section 2


147

Mineralised low grade stockpile – Section 6


148

Pit Infill Zone – Section 14


149

In Situ Zone – Section 0


150

In Situ – Section -2


151

APPENDIX 5. Dombraly IDW Resource Nov

2011


152

Dombraly IDW Resource Nov 2011

In Situ Total Inferred Resources

Total

From SG Cum Vol Cum Tonnes Cum Au Au Metal (g) Au Metal (oz) Material

5.0000 2.8176 31100.00 87626.88 5.7775100 506,265 16,277 Total

2.0000 2.6075 396590.00 1040624.92 3.1351900 3,262,557 104,894 Total

1.0000 2.7019 1018748.00 2721602.74 1.9921800 5,421,923 174,319 Total

0.9000 2.6144 1072516.00 2862172.80 1.9404700 5,553,960 178,564 Total

0.8000 2.6532 1262560.00 3366397.98 1.7773300 5,983,200 192,364 Total

0.7000 2.6881 1399964.00 3735758.60 1.6759200 6,260,833 201,290 Total

0.6000 2.6538 1571722.00 4191571.76 1.5643300 6,557,001 210,812 Total

0.5000 2.6449 1732398.00 4616551.66 1.4716200 6,793,810 218,426 Total

0.4000 2.6692 1916432.00 5107777.66 1.3737400 7,016,758 225,594 Total

0.3000 2.5930 2302724.00 6109422.44 1.2028200 7,348,535 236,261 Total

0.2000 2.6607 2540620.00 6742402.10 1.1131100 7,505,035 241,292 Total

0.1000 2.6375 2807222.00 7445552.52 1.0208300 7,600,643 244,366 Total

0.0000 2.6719 2812728.00 7460264.00 1.0188300 7,600,741 244,369 Total

Oxide

Total


5 2.60 2,581 6,715 6.70 44,977 1,446 Oxide

2 2.60 149,291 388,419 2.96 1,148,142 36,914 Oxide

1 2.61 362,614 945,445 2.04 1,931,857 62,111 Oxide

0.9 2.63 402,845 1,051,428 1.93 2,030,654 65,287 Oxide

0.8 2.60 434,148 1,132,952 1.85 2,100,992 67,548 Oxide

0.7 2.61 469,575 1,225,353 1.77 2,169,746 69,759 Oxide

0.6 2.62 527,271 1,376,308 1.65 2,266,972 72,885 Oxide

0.5 2.62 598,925 1,564,166 1.52 2,373,090 76,297 Oxide

0.4 2.62 665,616 1,738,682 1.41 2,451,576 78,820 Oxide

0.3 2.60 808,883 2,111,880 1.22 2,577,486 82,868 Oxide

0.2 2.63 928,069 2,425,326 1.10 2,656,266 85,401 Oxide

0.1 2.63 1,043,408 2,728,155 0.99 2,699,646 86,796 Oxide

0 2.60 1,045,198 2,732,809 0.99 2,699,688 86,797 Oxide

Transitional

Total


5 2.60 1,156 3,006 5.94 17,840 574 Transitional

2 2.57 48,098 123,496 3.01 371,380 11,940 Transitional

1 2.64 108,223 282,114 2.06 582,469 18,727 Transitional

0.9 2.59 110,168 287,146 2.05 587,268 18,881 Transitional

0.8 2.56 115,956 301,990 1.99 599,943 19,289 Transitional


153

0.7 2.65 123,658 322,367 1.91 615,257 19,781 Transitional

0.6 2.61 146,456 381,979 1.71 653,146 20,999 Transitional

0.5 2.61 160,161 417,789 1.61 673,234 21,645 Transitional

0.4 2.64 178,670 466,579 1.49 695,408 22,358 Transitional

0.3 2.55 226,246 587,792 1.25 735,004 23,631 Transitional

0.2 2.61 248,694 646,384 1.16 749,449 24,095 Transitional

0.1 2.61 274,891 714,691 1.06 757,901 24,367 Transitional

0 2.64 276,401 718,681 1.05 757,936 24,368 Transitional

Primary

Total


5 2.85 27,363 77,906 5.69 443,448 14,257 Primary

2 2.62 199,202 528,710 3.30 1,743,029 56,040 Primary

1 2.77 547,912 1,494,043 1.95 2,907,588 93,481 Primary

0.9 2.55 559,504 1,523,599 1.93 2,936,051 94,396 Primary

0.8 2.67 712,457 1,931,455 1.70 3,282,257 105,527 Primary

0.7 2.72 806,732 2,188,038 1.59 3,475,830 111,750 Primary

0.6 2.69 897,995 2,433,285 1.49 3,636,860 116,928 Primary

0.5 2.67 973,313 2,634,596 1.42 3,747,476 120,484 Primary

0.4 2.71 1,072,147 2,902,517 1.33 3,869,781 124,416 Primary

0.3 2.60 1,267,596 3,409,751 1.18 4,036,054 129,762 Primary

0.2 2.71 1,363,857 3,670,692 1.12 4,099,355 131,797 Primary

0.1 2.65 1,488,923 4,002,706 1.04 4,143,041 133,202 Primary

0 2.75 1,491,129 4,008,773 1.03 4,143,107 133,204 Primary

Oxide Domain Inferred Resources

Total

From SG Cum Vol Cum Tonnes Cum Au

Au Metal (g)

Au Metal (oz) Material

5.0000 0.0000 0.00 0.00 0.0000000 0 0 Total

2.0000 2.6499 13060.00 34607.80 2.3250100 80,463 2,587 Total

1.0000 2.6500 55222.00 146336.70 1.5286900 223,703 7,192 Total

0.9000 2.6494 78096.00 206938.60 1.3533500 280,060 9,004 Total

0.8000 2.6497 90404.00 239551.20 1.2875600 308,437 9,916 Total

0.7000 2.6465 94536.00 250486.60 1.2644600 316,730 10,183 Total

0.6000 2.6500 96772.00 256412.00 1.2501500 320,553 10,306 Total

0.5000 2.6500 100100.00 265231.20 1.2271200 325,471 10,464 Total

0.4000 2.6462 103962.00 275450.70 1.1980400 330,001 10,610 Total

0.3000 2.6481 109772.00 290836.20 1.1526200 335,224 10,778 Total

0.2000 2.6500 114914.00 304462.50 1.1118900 338,529 10,884 Total

0.1000 2.6500 118860.00 314919.40 1.0806000 340,302 10,941 Total

0.0000 0.0000 118860.00 314919.40 1.0806000 340,302 10,941 Total


154

Oxide

Total


Au Metal (g)


5 0.00 0 0 0.00 0 0 Oxide

2 2.65 13,060 34,608 2.33 80,463 2,587 Oxide

1 2.65 55,222 146,337 1.53 223,703 7,192 Oxide

0.9 2.65 78,096 206,939 1.35 280,060 9,004 Oxide

0.8 2.65 90,404 239,551 1.29 308,437 9,916 Oxide

0.7 2.65 94,536 250,487 1.26 316,730 10,183 Oxide

0.6 2.65 96,772 256,412 1.25 320,553 10,306 Oxide

0.5 2.65 100,100 265,231 1.23 325,471 10,464 Oxide

0.4 2.65 103,962 275,451 1.20 330,001 10,610 Oxide

0.3 2.65 109,772 290,836 1.15 335,224 10,778 Oxide

0.2 2.65 114,914 304,463 1.11 338,529 10,884 Oxide

0.1 2.65 118,860 314,919 1.08 340,302 10,941 Oxide

0 0.00 118,860 314,919 1.08 340,302 10,941 Oxide

In Situ Domain A Inferred

Total

From SG Cum Vol Cum Tonnes

Cum Au

Au Metal (g)


5 2.60 4,996 12,990 6.47 84,107 2,704 Total

2 2.60 124,420 323,492 3.15 1,019,205 32,768 Total

1 2.60 181,108 470,881 2.60 1,224,621 39,372 Total

0.9 2.60 181,508 471,921 2.60 1,225,577 39,403 Total

0.8 2.60 183,000 475,800 2.58 1,228,852 39,508 Total

0.7 2.60 187,674 487,952 2.54 1,238,053 39,804 Total

0.6 2.60 188,674 490,552 2.53 1,239,774 39,860 Total

0.5 2.60 192,792 501,259 2.48 1,245,501 40,044 Total

0.4 2.60 194,832 506,563 2.46 1,248,073 40,126 Total

0.3 2.60 196,578 511,103 2.45 1,249,696 40,179 Total

0.2 2.60 197,932 514,623 2.43 1,250,509 40,205 Total

0.1 2.60 372,444 968,354 1.35 1,305,801 41,982 Total

0 2.60 375,892 977,319 1.34 1,305,866 41,985 Total

Oxide

Total


Cum Au

Au Metal (g)


5 2.60 2,485 6,461 6.76 43,690 1,405 Oxide

2 2.60 35,455 92,183 3.25 299,491 9,629 Oxide

1 2.60 63,800 165,880 2.43 402,561 12,943 Oxide

0.9 2.60 64,200 166,920 2.42 403,517 12,973 Oxide

0.8 2.60 64,786 168,444 2.40 404,807 13,015 Oxide

0.7 2.60 66,434 172,727 2.36 408,072 13,120 Oxide


155

0.6 2.60 66,507 172,917 2.36 408,194 13,124 Oxide

0.5 2.60 69,710 181,245 2.28 412,641 13,267 Oxide

0.4 2.60 70,628 183,632 2.25 413,770 13,303 Oxide

0.3 2.60 70,848 184,204 2.25 413,970 13,309 Oxide

0.2 2.60 70,988 184,568 2.24 414,057 13,312 Oxide

0.1 2.60 126,181 328,071 1.31 431,226 13,864 Oxide

0 2.60 127,971 332,725 1.30 431,259 13,865 Oxide

Transitional

Total


Cum Au

Au Metal (g)


5 2.60 1,156 3,006 5.94 17,840 574 Transitional

2 2.60 22,209 57,743 3.45 198,990 6,398 Transitional

1 2.60 40,861 106,237 2.50 266,015 8,553 Transitional

0.9 0.00 40,861 106,237 2.50 266,015 8,553 Transitional

0.8 2.60 41,767 108,593 2.47 268,000 8,616 Transitional

0.7 2.60 43,573 113,290 2.40 271,576 8,731 Transitional

0.6 2.60 44,398 115,435 2.36 272,993 8,777 Transitional

0.5 2.60 44,620 116,012 2.36 273,307 8,787 Transitional

0.4 2.60 45,742 118,929 2.31 274,751 8,833 Transitional

0.3 2.60 47,042 122,309 2.26 275,960 8,872 Transitional

0.2 2.60 47,588 123,729 2.23 276,295 8,883 Transitional

0.1 2.60 70,949 184,468 1.54 283,558 9,117 Transitional

0 2.60 72,211 187,749 1.51 283,582 9,117 Transitional

Primary

Total


Cum Au

Au Metal (g)


5 2.60 1,355 3,523 6.41 22,577 726 Primary

2 2.60 66,756 173,566 3.00 520,725 16,742 Primary

1 2.60 76,448 198,764 2.80 556,044 17,877 Primary

0.9 0.00 76,448 198,764 2.80 556,044 17,877 Primary

0.8 0.00 76,448 198,764 2.80 556,044 17,877 Primary

0.7 2.60 77,668 201,936 2.77 558,405 17,953 Primary

0.6 2.60 77,770 202,201 2.76 558,587 17,959 Primary

0.5 2.60 78,463 204,003 2.74 559,553 17,990 Primary

0.4 0.00 78,463 204,003 2.74 559,553 17,990 Primary

0.3 2.60 78,689 204,590 2.74 559,766 17,997 Primary

0.2 2.60 79,357 206,327 2.71 560,156 18,009 Primary

0.1 2.60 175,314 455,816 1.30 591,018 19,002 Primary

0 2.60 175,710 456,845 1.29 591,025 19,002 Primary

In Situ Domain B Inferred


156

Total


Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Total

2 2.55 23,762 60,593 4.01 242,810 7,807 Total

1 2.55 71,108 181,325 2.30 416,285 13,384 Total

0.9 2.55 82,144 209,467 2.12 443,411 14,256 Total

0.8 2.55 114,598 292,225 1.76 514,854 16,553 Total

0.7 2.55 137,430 350,447 1.59 558,454 17,955 Total

0.6 2.55 157,872 402,574 1.47 592,681 19,055 Total

0.5 2.55 169,612 432,511 1.41 609,334 19,591 Total

0.4 2.55 210,012 535,531 1.23 656,320 21,101 Total

0.3 2.55 229,578 585,424 1.15 673,911 21,667 Total

0.2 2.55 248,628 634,001 1.08 686,630 22,076 Total

0.1 2.55 249,158 635,353 1.08 686,823 22,082 Total

0 0.00 249,158 635,353 1.08 686,823 22,082 Total

Oxide

Total


Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Oxide

2 2.55 14,982 38,203 3.97 151,673 4,876 Oxide

1 2.55 44,472 113,402 2.36 267,374 8,596 Oxide

0.9 2.55 49,788 126,958 2.21 280,194 9,008 Oxide

0.8 2.55 58,962 150,352 2.00 300,302 9,655 Oxide

0.7 2.55 70,486 179,738 1.79 322,193 10,359 Oxide

0.6 2.55 80,675 205,720 1.65 339,027 10,900 Oxide

0.5 2.55 85,309 217,537 1.59 345,709 11,115 Oxide

0.4 2.55 106,085 270,517 1.37 369,975 11,895 Oxide

0.3 2.55 118,411 301,947 1.26 381,012 12,250 Oxide

0.2 2.55 126,079 321,500 1.20 386,176 12,416 Oxide

0.1 0.00 126,079 321,500 1.20 386,176 12,416 Oxide

0 0.00 126,079 321,500 1.20 386,176 12,416 Oxide

Transitional

Total


Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Transitional

2 2.55 5,037 12,843 3.89 49,995 1,607 Transitional

1 2.55 8,564 21,838 2.84 61,939 1,991 Transitional

0.9 2.55 9,788 24,959 2.60 64,963 2,089 Transitional

0.8 2.55 13,218 33,705 2.15 72,486 2,330 Transitional

0.7 2.55 14,961 38,151 1.99 75,812 2,437 Transitional

0.6 2.55 18,192 46,388 1.75 81,161 2,609 Transitional

0.5 2.55 20,924 53,355 1.59 84,997 2,733 Transitional


157

0.4 2.55 27,126 69,170 1.33 92,252 2,966 Transitional

0.3 2.55 28,619 72,977 1.28 93,665 3,011 Transitional

0.2 2.55 31,496 80,314 1.19 95,593 3,073 Transitional

0.1 2.55 31,618 80,625 1.19 95,635 3,075 Transitional

0 0.00 31,618 80,625 1.19 95,635 3,075 Transitional

Primary

Total


Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Primary

2 2.55 3,744 9,547 4.31 41,143 1,323 Primary

1 2.55 18,073 46,085 1.89 86,971 2,796 Primary

0.9 2.55 22,569 57,550 1.71 98,254 3,159 Primary

0.8 2.55 42,419 108,168 1.31 142,065 4,567 Primary

0.7 2.55 51,984 132,558 1.21 160,449 5,159 Primary

0.6 2.55 59,006 150,465 1.15 172,492 5,546 Primary

0.5 2.55 63,380 161,619 1.11 178,626 5,743 Primary

0.4 2.55 76,802 195,844 0.99 194,089 6,240 Primary

0.3 2.55 82,549 210,500 0.95 199,232 6,405 Primary

0.2 2.55 91,054 232,188 0.88 204,864 6,587 Primary

0.1 2.55 91,462 233,228 0.88 205,014 6,591 Primary

0 0.00 91,462 233,228 0.88 205,014 6,591 Primary

In Situ Domain C Inferred

Total


Au Metal (g)


5 0.00 0 0 0.00 0 0 Total

2 2.53 120,464 304,774 2.33 711,248 22,867 Total

1 2.53 134,982 341,504 2.27 774,143 24,889 Total

0.9 0.00 134,982 341,504 2.27 774,143 24,889 Total

0.8 2.53 135,004 341,560 2.27 774,190 24,891 Total

0.7 0.00 135,004 341,560 2.27 774,190 24,891 Total

0.6 0.00 135,004 341,560 2.27 774,190 24,891 Total

0.5 2.53 135,894 343,812 2.26 775,499 24,933 Total

0.4 2.53 136,932 346,438 2.24 776,655 24,970 Total

0.3 2.53 142,454 360,409 2.17 781,359 25,121 Total

0.2 2.53 182,274 461,153 1.75 804,929 25,879 Total

0.1 2.53 182,898 462,732 1.74 805,204 25,888 Total

0 0.00 182,898 462,732 1.74 805,204 25,888 Total

Oxide

Total

From SG Cum Vol Cum Cum Au Au Metal Au Metal Material


158

Tonnes (g) (oz)

5 0.00 0 0 0.00 0 0 Oxide

2 2.53 32,742 82,836 2.39 198,363 6,378 Oxide

1 2.53 35,191 89,032 2.35 208,818 6,714 Oxide

0.9 0.00 35,191 89,032 2.35 208,818 6,714 Oxide

0.8 0.00 35,191 89,032 2.35 208,818 6,714 Oxide

0.7 0.00 35,191 89,032 2.35 208,818 6,714 Oxide

0.6 0.00 35,191 89,032 2.35 208,818 6,714 Oxide

0.5 2.53 35,687 90,287 2.32 209,536 6,737 Oxide

0.4 2.53 36,331 91,916 2.29 210,221 6,759 Oxide

0.3 2.53 40,340 102,059 2.09 213,644 6,869 Oxide

0.2 2.53 53,130 134,419 1.65 221,360 7,117 Oxide

0.1 2.53 53,628 135,679 1.63 221,580 7,124 Oxide

0 0.00 53,628 135,679 1.63 221,580 7,124 Oxide

Transitional

Total


Au Metal (g)


5 0.00 0 0 0.00 0 0 Transitional

2 2.53 19,571 49,515 2.31 114,163 3,670 Transitional

1 2.53 22,347 56,537 2.25 126,925 4,081 Transitional

0.9 0.00 22,347 56,537 2.25 126,925 4,081 Transitional

0.8 0.00 22,347 56,537 2.25 126,925 4,081 Transitional

0.7 0.00 22,347 56,537 2.25 126,925 4,081 Transitional

0.6 0.00 22,347 56,537 2.25 126,925 4,081 Transitional

0.5 0.00 22,347 56,537 2.25 126,925 4,081 Transitional

0.4 0.00 22,347 56,537 2.25 126,925 4,081 Transitional

0.3 2.53 22,394 56,656 2.24 126,963 4,082 Transitional

0.2 2.53 27,611 69,856 1.86 130,042 4,181 Transitional

0.1 2.53 27,623 69,886 1.86 130,048 4,181 Transitional

0 0.00 27,623 69,886 1.86 130,048 4,181 Transitional

Primary

Total


Au Metal (g)


5 0.00 0 0 0.00 0 0 Primary

2 2.53 68,152 172,423 2.31 398,722 12,819 Primary

1 2.53 77,445 195,936 2.24 438,401 14,095 Primary

0.9 0.00 77,445 195,936 2.24 438,401 14,095 Primary

0.8 2.53 77,467 195,992 2.24 438,447 14,096 Primary

0.7 0.00 77,467 195,992 2.24 438,447 14,096 Primary

0.6 0.00 77,467 195,992 2.24 438,447 14,096 Primary

0.5 2.53 77,861 196,988 2.23 439,038 14,115 Primary

0.4 2.53 78,255 197,985 2.22 439,509 14,131 Primary

0.3 2.53 79,721 201,694 2.19 440,752 14,170 Primary


159

0.2 2.53 101,533 256,878 1.77 453,527 14,581 Primary

0.1 2.53 101,647 257,167 1.76 453,578 14,583 Primary

0 0.00 101,647 257,167 1.76 453,578 14,583 Primary

In Situ Domain D Inferred

Total

From SG Cum Vol

Cum Tonnes

Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Total

2 0.00 0 0 0.00 0 0 Total

1 2.62 114,392 299,707 1.26 377,631 12,141 Total

0.9 2.62 115,376 302,285 1.26 380,000 12,217 Total

0.8 2.62 181,606 475,808 1.11 530,435 17,054 Total

0.7 2.62 203,454 533,049 1.08 573,172 18,428 Total

0.6 2.62 272,562 714,112 0.96 688,304 22,129 Total

0.5 2.62 358,406 939,024 0.87 814,800 26,196 Total

0.4 2.62 376,164 985,550 0.85 835,845 26,873 Total

0.3 2.62 377,958 990,250 0.85 837,494 26,926 Total

0.2 2.62 378,300 991,146 0.85 837,736 26,934 Total

0.1 2.62 379,494 994,274 0.84 838,163 26,948 Total

0 2.62 380,074 995,794 0.84 838,170 26,948 Total

Oxide

Total

From SG Cum Vol

Cum Tonnes

Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Oxide

2 0.00 0 0 0.00 0 0 Oxide

1 2.62 22,271 58,350 1.26 73,521 2,364 Oxide

0.9 0.00 22,271 58,350 1.26 73,521 2,364 Oxide

0.8 0.00 22,271 58,350 1.26 73,521 2,364 Oxide

0.7 2.62 22,991 60,235 1.24 74,887 2,408 Oxide

0.6 2.62 48,912 128,149 0.92 118,060 3,796 Oxide

0.5 2.62 88,078 230,763 0.77 177,501 5,707 Oxide

0.4 0.00 88,078 230,763 0.77 177,501 5,707 Oxide

0.3 0.00 88,078 230,763 0.77 177,501 5,707 Oxide

0.2 0.00 88,078 230,763 0.77 177,501 5,707 Oxide

0.1 0.00 88,078 230,763 0.77 177,501 5,707 Oxide

0 0.00 88,078 230,763 0.77 177,501 5,707 Oxide

Transitional

Total

From SG Cum Vol

Cum Tonnes

Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Transitional


160

2 0.00 0 0 0.00 0 0 Transitional

1 2.62 18,653 48,871 1.26 61,577 1,980 Transitional

0.9 0.00 18,653 48,871 1.26 61,577 1,980 Transitional

0.8 0.00 18,653 48,871 1.26 61,577 1,980 Transitional

0.7 2.62 20,022 52,456 1.23 64,300 2,067 Transitional

0.6 2.62 35,980 94,267 0.96 90,570 2,912 Transitional

0.5 2.62 43,957 115,167 0.89 102,634 3,300 Transitional

0.4 0.00 43,957 115,167 0.89 102,634 3,300 Transitional

0.3 0.00 43,957 115,167 0.89 102,634 3,300 Transitional

0.2 0.00 43,957 115,167 0.89 102,634 3,300 Transitional

0.1 0.00 43,957 115,167 0.89 102,634 3,300 Transitional

0 0.00 43,957 115,167 0.89 102,634 3,300 Transitional

Primary

Total

From SG Cum Vol

Cum Tonnes

Cum Au

Au Metal (g)


5 0.00 0 0 0.00 0 0 Primary

2 0.00 0 0 0.00 0 0 Primary

1 2.62 73,468 192,486 1.26 242,533 7,798 Primary

0.9 2.62 74,452 195,064 1.26 244,899 7,874 Primary

0.8 2.62 140,682 368,587 1.07 395,335 12,710 Primary

0.7 2.62 160,442 420,358 1.03 433,986 13,953 Primary

0.6 2.62 187,670 491,696 0.98 479,674 15,422 Primary

0.5 2.62 226,372 593,094 0.90 534,668 17,190 Primary

0.4 2.62 244,130 639,620 0.87 555,715 17,867 Primary

0.3 2.62 245,924 644,320 0.87 557,356 17,919 Primary

0.2 2.62 246,266 645,216 0.86 557,596 17,927 Primary

0.1 2.62 247,460 648,345 0.86 558,024 17,941 Primary

0 2.62 248,040 649,864 0.86 558,032 17,941 Primary

Documents

A C A HOWE INTERNATIONAL LIMITED MICROMINE ...alhambraresources.com/news/releases/ACA Howe_MCS_Dombraly...MICROMINE CONSULTING SERVICES James Hogg, BSc, MSc, MAIG Marta Sostre, BSc,